San Ramon Feasibility Study

AMENDED NI 43-101 TECHNICAL REPORT

FEASIBILITY STUDY OF THE SANTA ROSA GOLD

PROJECT

Prepared for:

Red Eagle Mining Corporation

Issue Date: October 27, 2014

PREPARED BY THE FOLLOWING QUALIFIED PERSONS:

Mr. Stefan Gueorguiev, P. Eng., Lycopodium Minerals Canada Ltd.

Mr. Thomas L. Dyer, P.E., Mine Development Associates

Mr. Michael Lindholm, C.P.G., Mine Development Associates

Mr. W. Joseph Schlitt, Ph.D., P. Eng., Hydrometal Inc.

Mr. Terry Eldridge, P. Eng., Golder Associates South America Ltd.

Page ii

RED EAGLE MINING CORPORATION

FEASIBILITY STUDY OF THE SANTA ROSA GOLD PROJECT


AMENDED NI 43-101 TECHNICAL REPORT Lycopodium Report No.: Job No File Location: 16.04

Rev: A

This report is effective as of the 6th day of October, 2014

[Signed]____________________________________________________ October 27, 2014

Mr. Stefan Gueorguiev, P. Eng., Lycopodium Minerals Canada Ltd. Date

[Signed]____________________________________________________ October 27, 2014

Mr. Thomas L. Dyer, P.E., Mine Development Associates Date

[Signed]____________________________________________________ October 27, 2014

Mr. Michael Lindholm, C.P.G., Mine Development Associates Date

[Signed]____________________________________________________ October 27, 2014

Mr. W. Joseph Schlitt, Ph.D., P. Eng., Hydrometal Inc. Date

[Signed]____________________________________________________ October 27, 2014

Mr. Terry Eldridge, P. Eng., Golder Associates South America Ltd. Date

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AMENDED NI 43-101 TECHNICAL REPORT TABLE OF CONTENTS

1.0 SUMMARY 1.16 1.1 Property Description and Ownership 1.17 1.2 Exploration and Mining History 1.18 1.3 Geology and Mineralization 1.19 1.4 Mineral Resource Estimate 1.20 1.5 Mineral Reserve Estimate 1.21 1.6 Mining Methods 1.23 1.7 Mineral Processing and Metallurgical Testwork 1.25 1.8 Recovery Methods and Process Plant Design 1.26 1.9 Project Infrastructure 1.28 1.10 Capital Cost Estimate 1.30

1.10.1 Overall Capital Costs 1.30 1.10.2 Mine Capital Costs 1.31 1.10.3 Plant and Site Infrastructure Capital Costs 1.33

1.11 Operating Cost Estimate 1.34 1.11.1 Mine Operating Costs 1.34 1.11.2 Plant Operating Costs 1.36

1.12 Economic Analysis 1.36 1.13 Conclusions and Recommendations 1.37

1.13.1 Geology and Mineral Resource 1.37 1.13.2 Mining and Mineral Reserves 1.38 1.13.3 Metallurgical testwork and Recovery Methods 1.39 1.13.4 Project Infrastructure 1.40 1.13.5 Environmental Studies, Permitting, and Social and

Community Impact 1.41 1.13.6 Main Conclusion 1.41

2.0 INTRODUCTION 2.1 2.1 General 2.1 2.2 Sources of Information 2.2 2.3 Qualified Persons 2.2 2.4 Site Visits 2.6

3.0 RELIANCE ON OTHER EXPERTS 3.1

4.0 PROPERTY DESCRIPTION AND LOCATION 4.1 4.1 Property Location 4.1 4.2 Colombian Mining Law Regarding Concession Contracts 4.1 4.3 Land Area 4.2 4.4 Agreements and Encumbrances 4.1

4.4.1 Purchase Agreement for Five Original Project Concessions 4.2 4.4.2 Purchase Agreement for LKA-08004 4.2 4.4.3 Purchase Agreement with Bullet 4.2 4.4.4 Purchase Agreement with AngloGold Ashanti Colombia S.A 4.3 4.4.5 Agreement Regarding Artisanal Mining Operations 4.3 4.4.6 Royalties 4.3

4.5 Environmental Liabilities 4.3

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND

PHYSIOGRAPHY 5.1

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AMENDED NI 43-101 TECHNICAL REPORT 5.1 Access to Property 5.1 5.2 Climate 5.2 5.3 Physiography 5.2 5.4 Infrastructure and Local Resources 5.2 5.5 Regional Seismicity 5.2

6.0 HISTORY 6.1 6.1 Exploration History 6.1

7.0 GEOLOGIC SETTING AND MINERALIZATION 7.1 7.1 Geologic Setting 7.1

7.1.1 Regional Geology 7.1 7.1.2 Local Geology 7.5 7.1.3 Property Geology 7.6

7.2 Mineralization 7.6

8.0 DEPOSIT TYPES 8.1

9.0 EXPLORATION 9.1 9.1 Geologic Mapping 9.1 9.2 Geochemistry 9.1 9.3 Geophysics 9.3 9.4 Topography 9.5

10.0 DRILLING 10.1 10.1 Summary 10.1 10.2 Drilling by Red Eagle Mining 10.1 10.3 Drill-Hole Collar Surveys 10.2 10.4 Down-Hole Surveys 10.2

11.0 SAMPLE PREPARATION, ANALYSIS, AND SECURITY 11.1 11.1 Sampling Procedures 11.1

11.1.1 Surface and Adit Sampling 11.1 11.1.2 Drill Sampling 11.1

11.2 Sample Preparation and Analysis 11.2 11.2.1 Surface and Adit Sampling 11.2 11.2.2 Drill Samples 11.2

11.3 Sample Security 11.3 11.3.1 Surface and Adit Sampling 11.3 11.3.2 Drilling Samples 11.3

11.4 Quality Assurance / Quality Control 11.3 11.4.1 Surface and Adit Sampling 11.3 11.4.2 Drill Sampling 11.3

11.5 Bulk Density and Specific Gravity 11.4 11.6 Summary Statement 11.5

12.0 DATA VERIFICATION 12.1 12.1 Database Audit 12.1

12.1.1 Drill-Collar Audit 12.1 12.1.2 Down-Hole Survey Audit 12.1 12.1.3 Geological Data Audit 12.2 12.1.4 Assay Database Audit 12.3

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AMENDED NI 43-101 TECHNICAL REPORT 12.1.5 Specific Gravity Data 12.4 12.1.6 Geotechnical Data 12.4

12.2 Quality Control and Quality Assurance 12.5 12.2.1 Standards Assays 12.5 12.2.2 Field and Preparation Duplicate Sample Assays 12.8 12.2.3 Resource Update (SR-140 to SR-233) 12.11 12.2.4 Blanks 12.14 12.2.5 Check Assays 12.16

12.3 MDA Independent Verification of Drill-Collars 12.21 12.3.1 Initial Resource Estimate (SR-001 to SR-139) 12.21

12.4 MDA Independent Verification of Mineralization 12.22 12.5 Summary Statement on Data Verification 12.23

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING 13.1 13.1 Acme Metallurgical Limited Test Work 13.1 13.2 Kappes, Cassiday & Associates Test Work 13.2

13.2.1 Sample Description 13.3 13.2.2 Whole-ore Leaching 13.4 13.2.3 Flotation with Concentrate Regrinding and Leaching 13.5 13.2.4 Gravity Concentration 13.8 13.2.5 McClelland Laboratories, Inc. Confirmatory Test Work 13.8 13.2.6 McClelland Laboratories, Inc. Test Work on High Grade Ore 13.11 13.2.7 Comminution 13.31 13.2.8 Ore and Concentrate Mineralogy 13.35 13.2.9 Metallurgical QA/QC Program 13.41

13.3 Summary of Metallurgical Test Work 13.43 13.3.1 Metallurgical Results 13.43 13.3.2 Risks and Opportunities 13.46

14.0 MINERAL RESOURCE ESTIMATES 14.1 14.1 Introduction 14.1

14.1.1 Mineral Resource 14.1 14.1.2 Inferred Mineral Resource 14.2 14.1.3 Indicated Mineral Resource 14.2 14.1.4 Measured Mineral Resource 14.3 14.1.5 Modifying Factors 14.3

14.2 Database 14.4 14.3 Underground Workings 14.5 14.4 Mineral Domains 14.5 14.5 Density 14.10 14.6 Sample and Composites Descriptive Statistics 14.11 14.7 Estimation 14.13 14.8 Mineral Resources 14.16 14.9 Discussion of Resources 14.22

15.0 MINERAL RESERVE ESTIMATES 15.1 15.1 Mineral Reserves 15.1 15.2 Economic Parameters 15.6 15.3 Dilution and Ore Loss 15.6 15.4 Reserve Comparison with the Preliminary Economic Assessment 15.7

16.0 MINING METHODS 16.1

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AMENDED NI 43-101 TECHNICAL REPORT 16.1 Underground Development 16.1 16.1.2 Portal Construction 16.5 16.1.3 Primary Development 16.12 16.1.4 Haulage Drifts 16.12 16.1.5 Attack Ramps 16.13 16.1.6 Ventilation Shafts, Raises and Drifts 16.13 16.1.7 Other Development 16.13

16.2 Stoping Methods 16.13 16.2.2 Attack Ramp Development 16.16 16.2.3 Development in Ore 16.17 16.2.4 Rib Mining 16.17 16.2.5 Mining Additional Lifts 16.17 16.2.6 Mucking 16.18 16.2.7 Backfill 16.18 16.2.8 Overall Stope Geometry 16.19 16.2.9 Modified MSDF for Wide Stopes 16.20 16.2.10 Modified MSDF for Multiple Veins 16.20 16.2.11 Benefits of MSDF 16.20

16.3 Ventilation 16.21 16.3.2 Equipment Loads 16.26 16.3.3 Airflow Velocity 16.26 16.3.4 Surface Fans 16.26

16.4 Geotechnical Studies 16.27 16.5 Hydrological Studies 16.30 16.6 Development and Production Scheduling 16.35

16.6.1 Mine Development Schedule 16.35 16.6.2 Mine Production Schedule 16.36

16.7 Mine Equipment 16.37 16.8 Manpower 16.39

17.0 RECOVERY METHODS 17.1 17.1 Process Design 17.1

17.1.1 Summary 17.1 17.1.2 Design Philosophy 17.1 17.1.3 Key Process Design Criteria 17.1 17.1.4 Projected Plant Recovery 17.3 17.1.5 Selected Process Flowsheet 17.4

17.2 Process Description 17.7 17.2.1 Run-of-Mine (ROM) Pad 17.7 17.2.2 Crushing Circuit 17.7 17.2.3 Grinding and Classification Circuit 17.8 17.2.4 Flotation and Regrind 17.10 17.2.5 Pre-Leach Thickening 17.11 17.2.6 Leach and Carbon Adsorption 17.11 17.2.7 Acid Wash 17.12 17.2.8 AARL Elution 17.13 17.2.9 Electrowinning and Doré Smelting 17.13 17.2.10 Carbon Regeneration 17.14 17.2.11 Cyanide Destruction Circuit 17.14 17.2.12 Tailings Disposal 17.14 17.2.13 Future Expansion 17.15

17.3 Reagents and Consumables 17.15

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AMENDED NI 43-101 TECHNICAL REPORT 17.4 Services 17.16 17.4.1 Raw Water 17.16 17.4.2 Fire Water 17.17 17.4.3 Potable Water 17.17 17.4.4 Process Water 17.17 17.4.5 Plant, Instrument, Flotation, and Air Supply 17.18

17.5 Instrumentation and Control 17.18 17.6 Metallurgical Accounting 17.18

18.0 PROJECT INFRASTRUCTURE 18.1 18.1 Project Infrastructure 18.1

18.1.1 Infrastructure Scope 18.1 18.1.2 Site Access 18.2 18.1.3 Site Plan 18.2 18.1.4 Site Roads 18.4

18.2 Plant Area Buildings 18.6 18.2.1 Buildings General 18.6 18.2.2 Plant Workshop and Main Warehouse 18.6 18.2.3 Reagents Permanent Storage 18.6 18.2.4 Gold Room 18.6 18.2.5 Tails Filtration Building 18.6 18.2.6 Plant Air Compressors/ Blower Shed 18.7 18.2.7 Filter Area Compressors Shed 18.7 18.2.8 Guard House 18.7 18.2.9 Plant Administration Building 18.7 18.2.10 Assay and Metallurgical Laboratory 18.7 18.2.11 Plant Control Room 18.8 18.2.12 Crusher Control Room 18.8 18.2.13 Main Switch Room 18.8 18.2.14 Reagents Switch Room 18.8 18.2.15 Filter Plant Switch Room 18.8

18.3 Mine Area Buildings 18.9 18.3.1 Truck Shop 18.9 18.3.2 Warehouse 18.9 18.3.3 Explosives Storage 18.9

18.4 Building Fire Protection Systems 18.9 18.5 Sewage Treatment 18.10 18.6 Security System and CCTV Monitoring 18.10 18.7 Site Services 18.10

18.7.1 First Aid 18.10 18.7.2 Communication Systems 18.10 18.7.3 Underground Communications 18.11 18.7.4 Transportation 18.11 18.7.5 Solid Waste Disposal 18.11

18.8 Fuel Delivery and Storage 18.11 18.9 Site Fencing 18.11 18.10 Power Supply and Distribution 18.12

18.10.1 Power Supply – Incoming 18.12 18.10.2 Main Transformer and Medium Voltage Switchgear 18.12 18.10.3 Power Distribution 18.13 18.10.4 Power Demand 18.13 18.10.5 Process Plant Power Demand 18.13

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AMENDED NI 43-101 TECHNICAL REPORT 18.10.6 Mine Area Power Requirements 18.14 18.10.7 Voltage Selection 18.16 18.10.8 Voltage Drops Limit 18.17 18.10.9 Supply Voltage 18.17 18.10.10 Emergency Power Supply - Process Plant 18.17 18.10.11 Emergency Power Supply – Mine Area 18.18 18.10.12 Construction Power 18.18 18.10.13 Power Quality 18.18 18.10.14 Lightning Protection System (LPS) 18.18 18.10.15 Electrical Switch Rooms 18.19

18.11 Transportation and Logistics 18.19 18.11.1 General 18.19 18.11.2 Port of Barranquilla 18.20 18.11.3 Port of Buenaventura 18.21 18.11.4 Containerized Cargo and Legal Load Trucking 18.21

18.12 Waste Rock and Tailings Facility 18.22 18.12.1 Design Concept 18.22 18.12.2 Geotechnical Investigation 18.23 18.12.3 Dry Waste Management Facility Design 18.30 18.12.4 DWMF Slope Stability 18.32 18.12.5 Deposition Plan 18.33 18.12.6 DWMF Water Management 18.36 18.12.7 Soil and Topsoil Storage 18.39 18.12.8 Preliminary Acid Drainage Considerations 18.40

19.0 MARKET STUDIES AND CONTRACTS 19.1 19.1 Marketing Studies 19.1

20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY

IMPACT 20.1 20.1 Introduction 20.1

20.1.1 Requirement For EIA 20.1 20.1.2 Objectives of the EIA 20.2

20.2 Environmental Licensing 20.2 20.2.1 Process Stages of the Santa Rosa Gold Project 20.3 20.2.2 Community Participation 20.4

20.3 The Company 20.5 20.4 The Santa Rosa Gold Project 20.5 20.5 Baseline Studies 20.7

20.5.1 Baseline Study Consultants 20.7 20.5.2 Areas of Influence 20.7 20.5.3 Abiotic Environment 20.11 20.5.4 Biotic Environment 20.22 20.5.5 Socio-Economic Environment 20.25

20.6 Corporate Social Responsibility 20.33 20.6.1 Introduction 20.33 20.6.2 Relations with Stakeholders 20.33 20.6.3 Consultation 20.35 20.6.4 Social Management Plan 20.37

20.7 Natural Resource Requirements 20.38 20.8 Environmental Assessment 20.41

20.8.1 Assessment Process 20.41

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AMENDED NI 43-101 TECHNICAL REPORT 20.9 Impact Assessment 20.44 20.10 Environmental Zoning 20.45 20.11 Environmental Management Plan 20.46

20.11.1 Introduction 20.46 20.11.2 Management Programs 20.46

20.12 Emergency Plans 20.61 20.13 Closure Plan 20.63

20.13.1 Partial Closure 20.63 20.13.2 Gradual Closure 20.64 20.13.3 Temporary Closure 20.64 20.13.4 Final Closure and Post-closure 20.64

20.14 Investment Plan 20.67 20.14.1 Introduction 20.67 20.14.2 Investment Formula 20.67 20.14.3 Investment Proposals 20.67 20.14.4 Additional Investments 20.67

20.15 Feasibility Study, September, 2014 Update 20.68 20.15.1 Technical Differences 20.68

21.0 CAPITAL AND OPERATING COSTS 21.1 21.1 Mine Capital Costs 21.1

21.1.1 Mine Capital Cost 21.1 21.2 Mine Operating Cost Estimate 21.8

21.2.1 Mine Operating Cost 21.8 21.3 Process Plant Capital Costs 21.32

21.3.1 Introduction 21.32 21.3.2 Summary 21.32 21.3.3 Estimate Support Documents 21.33 21.3.4 Procurement 21.35 21.3.5 Earthworks 21.35 21.3.6 Concrete 21.36 21.3.7 Steelwork 21.36 21.3.8 Plate work and Tankage 21.36 21.3.9 Mechanical Equipment 21.37 21.3.10 Plant Pipework and Valves 21.37 21.3.11 Overland Piping 21.37 21.3.12 Electrical and Instrumentation 21.37 21.3.13 Buildings 21.38 21.3.14 Labour Rates and Crew Rates 21.38 21.3.15 Contractors’ Indirects 21.39 21.3.16 Productivity 21.39 21.3.17 EPCM Services 21.40 21.3.18 Working Capital 21.40 21.3.19 Vendor Commissioning 21.40 21.3.20 Spares 21.40 21.3.21 First Fills Inventory 21.41 21.3.22 Exchange Rates 21.41 21.3.23 Freight 21.41 21.3.24 Owners’ Costs 21.41 21.3.25 Contingency 21.42 21.3.26 Qualifications for Capital Cost Estimate 21.42

21.4 Process Plant Operating Cost Estimate 21.43

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AMENDED NI 43-101 TECHNICAL REPORT 21.4.1 Introduction 21.43 21.4.2 Plant Design Parameters 21.44 21.4.3 Qualifications and Exclusions 21.44 21.4.4 Operating Cost Accuracy 21.44 21.4.5 Cost Categories 21.44 21.4.6 Process Plant Labour 21.45 21.4.7 Consumables 21.48 21.4.8 Power 21.51 21.4.9 Maintenance 21.51 21.4.10 General and Administration Costs 21.52 21.4.11 Plant Operating Cost Summary 21.54

22.0 ECONOMIC ANALYSIS 22.1 22.1 Introduction 22.1 22.2 Principal Assumptions 22.2

22.2.1 Life-of-Mine Process Plant Feed Schedule and Recovery 22.2 22.2.2 Gold Selling Price, Exchange Rate and Escalation 22.2 22.2.3 Revenue Deductions - Refining Costs and Royalties 22.2 22.2.4 Land Reclamation Cost 22.3 22.2.5 Working Capital 22.3 22.2.6 Taxes 22.3 22.2.7 Depreciation 22.4 22.2.8 Cash-flow Model 22.4 22.2.9 Financial Results 22.6 22.2.10 Economic Sensitivity 22.6

23.0 ADJACENT PROPERTIES 23.1

24.0 OTHER RELEVANT DATA 24.1 24.1 Project Execution Strategy – Mining 24.1 24.2 Project Execution Strategy – Plant and Infrastructure 24.6 24.3 Detailed Engineering and Design 24.7 24.4 Procurement Management 24.8 24.5 Contracts Management 24.9

24.5.1 Strategy 24.9 24.5.2 Bulk Earthworks 24.12 24.5.3 Concrete Supply and Installation 24.12 24.5.4 Field Erected Tankage 24.13 24.5.5 Structural Steel Supply and Fabrication 24.13 24.5.6 Structural, Mechanical and Piping (SMP) Installation Works 24.13 24.5.7 Electrical and Instrumentation (E&I) Installation Works 24.14

24.6 Construction Management 24.14 24.7 Commissioning 24.15

24.7.1 General 24.15 24.7.2 Pre-commissioning and Testing 24.16 24.7.3 Mechanical Completion 24.16 24.7.4 Wet Commissioning 24.16 24.7.5 Process Commissioning 24.17

24.8 Project Close-out and Handover 24.17 24.9 Project Execution Schedule 24.17

25.0 INTERPRETATION AND CONCLUSIONS 25.1

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AMENDED NI 43-101 TECHNICAL REPORT 25.1 Geology and Mineral Resource 25.1 25.2 Mining and Mineable Reserves 25.1 25.3 Metallurgical Testing and Recovery Methods 25.2 25.4 Project Infrastructure 25.2 25.5 Capital Cost Estimate 25.3 25.6 Operating Cost Estimate 25.3 25.7 Economic Analysis 25.4 25.8 Environmental Studies, Permitting, and Social and Community Impact 25.4 25.9 Risks and Opportunities 25.5

26.0 RECOMMENDATIONS 26.1 26.1 Geology and Mineral Resources 26.1 26.2 Mining and Mineral Reserves 26.1 26.3 Metallurgical Testing and Recovery Methods 26.2 26.4 Project Infrastructure 26.2 26.5 Environmental Studies, Permitting, and Social and Community Impact 26.3 26.6 Main Recommendation 26.3

27.0 REFERENCES 27.1

TABLES

Table 1.1 Mineral Resources 1.21 Table 1.2 Total Proven and Probable Reserves by Material Type 1.22 Table 1.3 Changes from PEA Material Processed to Reserves 1.22 Table 1.4 Mine Development Summary 1.24 Table 1.6 Santa Rosa Gold Project Overall Capital Cost Estimate ($M) 1.31 Table 1.7 Total Mine Capital ($M) 1.32 Table 1.9 Mine Operating Cost 1.35 Table 1.10 Summary of Process Operating Cost Estimate at nominal duty of

360,000 tpa 1.36 Table 1.11 Financial Performance Indicators 1.37 Table 2.1 Technical Report Section List of Responsibility 2.4 Table 4.1 Legal Status of Red Eagle Mining’s Mining Concession Contracts and

Applications 4.3 Table 4.2 Processes Required for Environmental Activities 4.4 Table 12.1 Summary of Gold Results for Standards (SR-001 to SR-139) 12.6 Table 12.2 Comparison of Means for Field Duplicates (SR-001 to SR-139) 12.9 Table 12.3 Comparison of Means for Preparation Duplicates (SR-001 to SR-139) 12.10 Table 12.4 Comparison of Means for Field Duplicates (SR-140 to SR-233) 12.12 Table 12.6 Comparison of Means for Pulp Check Assays, Mean of Pairs = 0.01 to

0.4g Au/t (SR-001 to SR-139) 12.17 Table 12.7 Comparison of Means for Pulp Check Assays, Mean >0.4g Au/t

(SR-001 to SR-139) 12.17 Table 12.8 Comparison of Means for Pulp Check Assays, Mean of Pairs > 0.10g

Au/t (SR-140 to SR-233) 12.19 Table 12.9 Core Recoveries for by Lithologic, Weathering and Oxidation Types 12.20 Table 12.10 MDA Confirmation Samples from Core, Outcrop, and Adits 12.23 Table 13.3 Results of Direct Leaching on All Ore Types and Variability Samples 13.5 Table 13.4 Summary of Optimum Rougher Flotation Results 13.6 Table 13.5 Summary of Concentrate Leach Tests 13.7 Table 13.6 Summary of Gold Recovery from Sulfides with Gravity Concentration 13.8

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AMENDED NI 43-101 TECHNICAL REPORT Figure 13.1 Gold Leach Rate Profiles, Bottle Roll Test 13.9 Table 13.7 Comparable Gold Extractions in the MLI and KCA Tests 13.10 Table 13.8 Comparable Cyanide Consumptions in the MLI and KCA Tests 13.10 Table 13.9 MLI Gold Assays by Size Fraction in the Leach Residues 13.10 Table 13.10 Partial Chemical Composition of the Six Test work Underground

Composites 13.14 Table 13.11 Summary of Results of the Dissolved Oxygen Tests 13.16 Table 13.12 Tail Screen Analysis for Direct Leach Test CY-23 (Composite 4

Ground to a Nominal P80 of 53 µm) 13.19 Table 13.13 Summary results of the E-GRG Tests for Gold 13.21 Table 13.14 Summary of the Gravity / Cyanidation Tests 13.23 Table 13.15 Results of the Optimization Tests Using Composite 6 13.25 Table 13.16 Results of the Float-Combined Leach Tests on the Grade Composites

(with 11 min. concentrate regrind) 13.28 Table 13.17 Comminution Test Results 13.31 Table 13.18 Description of Crusher Work Index Samples 13.32 Table 13.19 Bond Low Impact Crusher Work Index Results 13.32 Table 13.20 Bond Ball Mill and Abrasion Index Results 13.33 Table 13.21 Head Assays on CWi Samples 13.33 Table 13.22 Selected SMC Test Results 13.35 Table 13.23 Cyclosizer Particle Sizing on Unreground Composite 6 Rougher

Concentrate 13.40 Table 14.1 Exploration and Resource Database Descriptive Statistics 14.5 Table 14.2 Density Measurements and Values Applied to the Block Model 14.10 Table 14.3 Descriptive Statistics of Coded Samples 14.12 Table 14.4 Descriptive Statistics of Coded Composites 14.13 Table 14.5 Estimation Parameters 14.15 Table 14.6 Classification Criteria 14.17 Table 14.7 San Ramon Block-Diluted Gold Resources – Measured and Indicated 14.18 Table 14.8 San Ramon Block-Diluted Gold Resources – Inferred 14.19 Table 15.1 Total Proven and Probable Reserves by Material Type 15.3 Table 15.2 Total Proven and Probable Reserves and Dilution 15.4 Table 15.3 Economic Parameters 15.6 Table 15.4 Changes from PEA Material Processed to Reserves 15.8 Table 16.1 Equipment Loads 16.26 Table 16.2 Recommended Maximum Airflow Velocities 16.26 Table 16.3 Estimated Fan Operating Points (Calculated at Ventilation Raise

Collar) 16.27 Table 16.4 Hole Collar Locations 16.32 Table 16.5 Yearly Development Schedule 16.36 Table 16.6 Yearly Mine Production 16.37 Table 16.7 Equipment Requirements and Underground Facilities 16.38 Table 16.8 Mine Personnel Requirements 16.40 Table 17.1 Summary of Key Process Design Criteria 17.2 Table 17.2 Summary of Metallurgical and Anticipated Gold Recoveries 17.3 Table 17.3 Reagent Summary 17.16 Table 18.1 Process Plant Power Requirements 18.14 Table 18.2 Maximum Mine Area Power Demand 18.15 Table 18.3 Mine Area Electrical Equipment in Operation by Year 18.15 Table 18.4 Voltage Levels 18.16 Table 18.5 Supply Voltage 18.17 Table 18.6 Summary of geotechnical investigation 18.23

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AMENDED NI 43-101 TECHNICAL REPORT Table 18.7 Shear strength estimates for saprolite soil 18.28 Table 18.8 Summary of triaxial strength tests on 'undisturbed' samples 18.29 Table 18.9 Summary of direct shear strength tests on remoulded samples (to be

used as compacted fill) 18.29 Table 18.10 Summary of strength values for in-situ and compacted saprolite 18.29 Table 18.11 Results of slope stability analyses (Golder, 2014b) 18.32 Table 18.12 San Ramón project waste production plan 18.33 Table 21.1 Santa Rosa Gold Project Overall Capital Cost Estimate ($M) 21.1 Table 21.2 Mine Development Capital Cost 21.3 Table 21.3 Contractor Quotation for Development ($/m) 21.4 Table 21.4 Owner’s Cost - Mine Departmental Capital 21.6 Table 21.5 Total Mine Capital 21.7 Table 21.6 Mine Operating Cost 21.9 Table 21.7 Mine Operating Cost per Tonne 21.10 Table 21.8 Annual Drilling Costs 21.12 Table 21.9 Annual Blasting Costs 21.13 Table 21.10 Annual Loading Costs 21.15 Table 21.11 Annual Haulage Costs 21.17 Table 21.12 Annual Ground Support Costs 21.19 Table 21.13 Annual Backfill Costs 21.21 Table 21.14 Annual Mine Support Costs 21.23 Table 21.15 Annual Stope Delineation Costs 21.25 Table 21.16 Annual Ventilation Costs 21.27 Table 21.17 Annual Dewatering Costs 21.29 Table 21.18 Annual Mine General Services 21.31 Table 21.19 Process Plant Estimated Total Installed Costs ($M) 21.33 Table 21.20 Key Documents Level of Development 21.34 Table 21.21 List of Equipment and Materials (Budget Price/Estimation) 21.35 Table 21.22 Crew Rates 21.39 Table 21.23 Productivity Factor (PF) 21.40 Table 21.24 Exchange Rates 21.41 Table 21.25 Wages and Salaries 21.46 Table 21.26 Summary of Major Consumables 21.50 Table 21.27 Summary of Process Plant Power Costs 21.51 Table 21.28 Summary of Maintenance Costs 21.52 Table 21.29 Summary of G&A Costs 21.53 Table 21.30 Summary of Process Operating Cost Estimate (nominal 360,000 tpa

throughput) 21.54 Table 21.31 Area Breakdown of Process Operating Cost Estimate (nominal

360,000 tpa throughput) 21.55 Table 22.1 Financial Performance Indicators 22.1 Table 22.2 LOM Process Plant Feed Schedule and Metal Production 22.2 Table 22.3 Santa Rosa Cash-Flow Model 22.5 Table 22.4 Financial Performance Indicators 22.6 Table 22.5 Economic Sensitivity to Operating Cost 22.7 Table 22.6 Economic Sensitivity to Capital Cost 22.7 Table 22.7 Economic Sensitivity to Gold Selling Price 22.8 Table 24.1 Procurement Long Lead Items 24.9 Table 24.2 Proposed Major Contracts 24.11 Table 24.3 Major Project Milestones 24.18 Figure 24.1 Project execution schedule 24.19 Table 26.1 Recommended Additional Work Cost Estimate 26.3

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FIGURES

Figure 1.1 Location of the Santa Rosa Gold Project 1.17

Figure 4.1 Location of the Santa Rosa Project 4.1

Figure 4.2 Concessions and Applications Forming the Santa Rosa Gold Project 4.4

Figure 4.3 Concessions and Applications in the Immediate Vicinity of the San

Ramon Deposit 4.1

Figure 5.1 Access to the Santa Rosa Project 5.1

Figure 5.2 Maximum expected magnitude (ML) for Antioquia (Project area inside

red square) From Sánchez (2009) 5.3

Figure 5.3 Peak ground acceleration map from Norma Sismo Resistente

Colombiana (NSR-10, 2010) 5.4

Figure 7.1 Volcano-plutonic Magmatic Arcs and Major Intrusions of the Northern

Andes 7.2

Figure 7.2 Santa Rosa Gold Project Regional Geology 7.4

Figure 7.3 Local Geology in the Vicinity of the Santa Rosa Gold Project 7.5

Figure 7.4 Oxidized Saprolite Exposed in a “Batición” 7.6

Figure 7.5 Water-Powered Stamp Mill 7.7

Figure 7.6 Adits and Baticiones on the Santa Rosa Gold Project 7.8

Figure 9.1 MMI Geochemistry of the Santa Rosa Concessions Showing Gold

Anomalies 9.2

Figure 9.2 First Vertical Derivative of Total Magnetic Intensity 9.3

Figure 9.3 Total Potassium Radiometric Plot 9.4

Figure 10.1 Drill-Hole Location Map in the San Ramon Deposit Area 10.3

Figure 12.1 Control Chart for Gold in Standard #3 (SR-001 to SR-139) 12.5

Figure 12.2 Control Chart for Gold in Standard #7 (SR-140 to SR-233) 12.7

Figure 12.3 Relative Percent Differences for Field Duplicates (SR-001 to SR-139) 12.9

Figure 12.4 Relative Percent Differences for Preparation Duplicates

(SR-001 to SR-139) 12.10

Figure 12.5 Relative Percent Differences for Field Duplicates (SR-140 to SR-233) 12.12

Figure 12.6 Relative Percent Differences for Preparation Duplicates

(SR-140 to SR-233) 12.13

Figure 12.7 Gold in Pulp Blanks vs. Preceding Sample (SR-001 to SR-139) 12.15

Figure 12.8 Gold in Pulp Blanks vs. Preceding Sample (SR-140 to SR-233) 12.15

Figure 12.9 Relative Percent Differences for Pulp Check Assays (SR-001 to SR-

139) 12.16

Figure 12.10 Relative Percent Differences for Pulp Check Assays (SR-140 to SR-

233) 12.18

Figure 12.11 Core Recovery versus Gold Grade 12.20

Figure 13.2 The Location of All Drill Intervals with Grades >2 g Au/t Used in the

Metallurgical Composites 13.13

Figure 13.3 The Gold-Silver Head Grade Relationship 13.15

Figure 13.4 The Relationship between Grind Size and Gold Recovery in a 72-Hr.

Direct Leach 13.18

Figure 13.5 Gravity Recoverable Gold and Silver vs. Grind Size (Composite 3) 13.20

Figure 13.6 Gravity Recoverable Gold by Size Fraction (Composite 3) 13.21

Figure 13.7 Total Gravity Recoverable Gold vs. Head Grade 13.22

Figure 13.8 Concentrate Particle Size Distributions as a Function of Regrind

Conditions 13.26

Figure 13.9 Residue Assays by Size Fraction with and without Regrinding 13.27

Figure 13.10A A Relationship between Gold Head Grade and Recovery 13.29

Page 1.13




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.10B Adjusted Relationship between Gold Head Grade and Recovery 13.29

Figure 13.11 Relationship between Head Grade and BMWi Values 13.34

Figure 13.12 Pyrite Grain Cut by Gold (arrow) Veinlet (Credit EGC) 13.36

Figure 13.13 A Polished Section of Sample SR-50051 Showing Crushed Pyrite

Cemented by Sphalerite, Galena-Gold, Quartz and Late-Stage

Carbonate Veinlets (Credit EGC) 13.37

Figure 13.14 Crushed Arsenopyrite (left)-Pyrite (right) Contact with Gold (arrow)

Along the Contact and Extending Along the Fractures (Credit EGC) 13.38

Figure 14.1 Gold Domains and Geology – Section 856500E 14.8

Figure 14.2 Gold Domains and Geology – Section 857700E 14.9

Figure 14.3 Gold Block Model Section 856500E 14.20

Figure 14.4 Gold Block Model Section 857700E 14.21

Figure 15.1 Mineral Reserves 15.5

Figure 15.2 Changes to Tonnage – PEA vs Reserves 15.8

Figure 15.3 Changes to Ounces of Gold – PEA vs Reserves 15.8

Figure 16.1 Underground Development – Long View 16.3

Figure 16.2 Underground Development – Plan View 16.4

Figure 16.3 Portal and Ramp Area Geology 16.5

Figure 16.4 Portal Pad Area 16.6

Figure 16.5 Portal Area Excavation Long Section 16.7

Figure 16.6 Cross-section of Portal Excavation Area 16.8

Figure 16.7 Subsurface Portion of the Ramp 16.9

Figure 16.8 Long Section of the Initial 200 m of the Ramp 16.10

Figure 16.9 Ramp Centerline Projection to Topography 16.11

Figure 16.10 Typical Cross-Section 16.12

Figure 16.11 Mechanized Shrinkage with Delayed Fill: Development in Ore 16.14

Figure 16.12 Mechanized Shrinkage with Delayed Fill: First Lift 16.14

Figure 16.13 Mechanized Shrinkage with Delayed Fill: Second Lift 16.14

Figure 16.14 Mechanized Shrinkage with Delayed Fill: Subsequent Lifts 16.15

Figure 16.15 Mechanized Shrinkage with Delayed Fill: Mucking 16.15

Figure 16.16 Mechanized Shrinkage with Delayed Fill: Backfill 16.15

Figure 16.17 Mechanized Shrinkage with Delayed Fill: Completed Stope 16.16

Figure 16.18 Ventilation Conceptual Design – Phase I 16.23

Figure 16.19 Ventilation Conceptual Design – Phase II 16.24

Figure 16.20 Ventilation Conceptual Design – End of Mine Life 16.25

Figure 16.21 Barton Q Characterization for San Ramon RQD 16.28

Figure 16.22 Bieniawski RMR Classification for San Ramon RQD 16.29

Figure 16.23 Bieniawski Span Tolerances for San Ramon 16.30

Figure 16.24 Geotechnical and Hydrological Drill Program Locations 16.31

Figure 16.25 Drilling of Piezometer Pole PZ-8 (distant) and Water Well W-2

(foreground) 16.32

Figure 16.26 Drilling of Water Well 15 16.33

Figure 16.27 Geological Section Through Hole 15 16.34

Figure 17.1 Projected Gold Recovery vs Mill Head Grade 17.4

Figure 17.2 Santa Rosa Gold Project Simplified Process Flowsheet 17.6

Figure 17.3 Crushing Circuit Layout 17.8

Figure 17.4 Grinding Circuit Layout 17.10

Figure 17.5 Leach and Carbon Adsorption Circuit Layout 17.12

Figure 18.1 Overall Site 3D View 18.2

Figure 18.2 Overall Site Layout 18.5

Figure 18.3 44 kV Overhead Power Line from EPM Substation 18.12

Figure 18.4 Santa Rosa Mine Site Location 18.19

Page 1.14




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.5 Port of Barranquilla 18.20

Figure 18.6 Port of Buenaventura 18.21

Figure 18.7 General Configuration of the DWMF 18.22

Figure 18.8 Geotechnical Investigation Carried Out in 2014 (Golder, 2014a) 18.24

Figure 18.9 Saprolite Thickness in Project Area (CRA, 2013) 18.26

Figure 18.10 Near-Vertical Cuts in Saprolite 18.27

Figure 18.11 Standard Proctor Test Results (GeoLogic Associates, 2014) 18.31

Figure 18.12 Tailings Particle Size Distribution. Data from GeoLogics Associates,

2014 18.31

Figure 18.13 Critical slip surface for static conditions (Golder, 2014b) 18.32

Figure 18.14 DWMF Deposit Configuration, Pre-mining and Year 1 18.34

Figure 18.15 DWMF Deposit Configuration, Years 1 and 2 18.34




Figure 18.19 Water Management for the DWMF 18.37

Figure 18.20 Runoff discharges for wet, average and dry years (Golder, 2013) 18.38

Figure 18.21 Water Flows for Average Year Conditions 18.39

Figure 18.22 Spoil Saprolite Storage Area 18.40

Figure 21.1 Overall Operating Cost Distribution (nominal 360,000 tpa throughput) 21.54

Figure 21.2 Consumables Cost Distribution 21.56

Figure 21.3 Power Cost Distribution 21.56

Figure 22.1 Project Pre-Tax NPV @ 0% (Undiscounted C.F.) Sensitivity 22.9

Figure 22.2 Project Pre-Tax NPV @ 5% Sensitivity 22.9

Figure 22.3 Project Pre-Tax NPV @ 8% Sensitivity 22.10

Figure 22.4 Project Pre-Tax IRR Sensitivity 22.10

Figure 22.5 Project Post-Tax NPV @ 0% (Undiscounted C.F.) Sensitivity 22.11

Figure 22.6 Project Post-Tax NPV @ 5% Sensitivity 22.11

Figure 22.7 Project Post-Tax NPV @ 8% Sensitivity 22.12

Figure 22.8 Project Post-Tax IRR Sensitivity 22.12

FREQUENTLY USED ACRONYMS AND ABBREVIATIONS

AA atomic absorption spectrometry

ABA acid-base accounting

Ag silver

ARD acid rock drainage

Au gold

CIL carbon-in-leach processing method

CIM Canadian Institute of Mining, Metallurgical, and Petroleum

CIP carbon-in-pulp processing method

cm centimeters

CN cyanide

core diamond core-drilling method

oC degrees centigrade

DWMF dry waste management facility

Page 1.15




AMENDED NI 43-101 TECHNICAL REPORT g/t grams per tonne

ha hectares

ICP inductively coupled plasma geochemical analysis

kg kilograms

km Kilometers

kN/m3 kilonewtons per cubic meter

kPa kilopascal

kV kilovolt

kW kilowatt

kWh/year kilowatt-hour per year

kWh/t kilowatt-hour per tonne

L/s liters per second

LHD load-haul dump unit

LNG liquefied natural gas

LOM life of mine

m meters

m3/h cubic meters per hour

µm micron

mm millimeters

MSDF mechanised shrinkage with delayed fill

MMI mobile metal ion geochemical technique

MW megawatt

NaCN sodium cyanide

Nm3 normal cubic meters. Measure of gas volume

oz troy ounce (12 oz to 1 pound)

QA/QC quality assurance and quality control

RC reverse-circulation drilling method

RQD rock-quality designation

t tonnes

TDS total dissolved solids

tpd tonnes per day

tph tonnes per hour

V volts

WAD weak acid dissociable (used for cyanide concentration)

Page 1.16




AMENDED NI 43-101 TECHNICAL REPORT 1.0 SUMMARY

The following Technical Report was compiled by Lycopodium Minerals Canada Ltd. (Lycopodium) and

presents the results of the Feasibility Study for the Santa Rosa Gold Project in the Department of

Antioquia, northern Colombia. The Technical Report was prepared at the request of Red Eagle Mining

Corporation (Red Eagle Mining), a British Columbia corporation. Red Eagle Mining is listed on the TSX

Venture exchange (RD) and the OTCQX (RDEMF).

This Feasibility Study was completed by:

Lycopodium Minerals Canada Ltd. (“Lycopodium”) for the process design, plant

infrastructure, capital and operating cost estimates, and economic analysis;

Mine Development Associates (“MDA”) for geology, resource, mining and reserves, as

well as mining capital and operating costs;

Hydrometal Inc. (“Hydrometal”) for metallurgical testwork management, largely

undertaken at McClelland Laboratory Services;

Golder Associates South America Ltd. (“Golder”) for the design of the dry waste

management facility, and geotechnical and hydrogeological aspects of the project; and

Unless otherwise denoted, all costs referred to in this Technical Report are quoted in current 3Q 2014

United States dollars.

The effective date of the mineral resource estimate is August 5, 2013. The effective date of the Feasibility

Study is October 6, 2014. The effective date of this report is October 6, 2014.

Page 1.17




AMENDED NI 43-101 TECHNICAL REPORT Figure 1.1 Location of the Santa Rosa Gold Project

1.1 Property Description and Ownership

The Santa Rosa Gold Project is located approximately 20 km southeast of the town of Santa Rosa de

Osos, in the municipality of the same name, in the Department of Antioquia, 73 km northeast of the

department capital Medellín in northern Colombia (Figure 4.1). The centre of the resource is located at

approximately latitude 6° 36' 57 N and longitude 75° 22' 20 W. The San Ramon deposit lies in the

southeastern part of concession B7560005 held by Red Eagle Mining and described in more detail in

Section 4.3. The project is accessible from Medellín via a paved highway for about 65 km, then via an

unpaved road for approximately 8 km.

Santa Rosa Gold Project Site

Page 1.18




AMENDED NI 43-101 TECHNICAL REPORT Red Eagle Mining’s property covers a total area of approximately 33,000 hectares and consists of:

12 concession contracts; and

14 concession-contract applications for which the technical studies have been approved

and free areas have been declared.

Red Eagle Mining holds a 100% interest in the project, subject to completing payments of $1,790,000 to

the underlying owners and subject to royalty obligations described below. The San Ramon deposit is

located on concession contract B7560005.

There is a government-imposed royalty on gold and silver production that is effectively 3.2%.

Liberty Metals & Mining Holdings (“LMM”), a subsidiary of Liberty Mutual Insurance, Boston, holds a 3%

net smelter return royalty on four concession contracts numbered B7560005, B7171005, H5791005, and

H5790005 and two concession contract applications numbered LDM-08061 and LKA-08004. Red Eagle

Mining may buy back 1% of the royalty for $8,333,333 for a period of two years from the date of the first

gold production. This royalty applies to the concession in which the resource is located (B7560005).

Bullet Holding Corp. (“Bullet”) and AngloGold Ashanti (“AngloGold”) respectively hold non-overlapping

1.5% and 2% net smelter return royalties on those concession contracts and applications that Red Eagle

Mining acquired from them, and still holds. These royalties do not apply to the concession in which the

current resource is located.

1.2 Exploration and Mining History

The gold potential of the Santa Rosa area was recognized around the year 1600, with gold production

from upland placers reaching a peak in the 18th century and mining from high-grade oxide ores peaking in

the 1940s. Based on the existence of the extensive abandoned mine workings, past gold production

appears to have been significant, but actual historic gold production is unknown. Although small-scale

mining by artisanal miners continues some distance outside the San Ramon deposit area, the first known

modern mineral exploration activity was begun by Red Eagle Mining in July 2010. Since that time, Red

Eagle Mining has undertaken the following exploration activities in the San Ramon deposit area:

Mapping and sampling historic workings and adits;

Grab, channel, and panel sampling of quartz veins and diorite-granodiorite wall rock;

Auger saprolite sampling;

Soil sampling with mobile-metal-ion (“MMI”) analysis, covering over 20 square

kilometers;

Helicopter-borne magnetic and radiometric surveying;

Page 1.19




AMENDED NI 43-101 TECHNICAL REPORT Induced-polarization surveying;

Drilling of 238 exploration core holes totalling 45,609 m between September, 2011 and

May, 2013; and

Drilling of 11 geotechnical and hydrogeological holes in 2014.

Red Eagle Mining undertook stream-sediment, rock-chip, trench, and MMI geochemical sampling in

addition to mapping of geology and the distribution of adits in the concessions beyond the area of the San

Ramon deposit with the objective of identifying potential targets for mineralization that may become future

satellite feed to the Santa Rosa Gold Project.

1.3 Geology and Mineralization

The north-western margin of South America, including the Santa Rosa Gold Project area, is comprised of

a mosaic of Paleozoic and younger autochthonous and allochthonous terranes that were accreted to the

South American continent. Subduction-related magmatic arcs now represented by plutonic batholiths,

sub-volcanic intrusions, and associated volcanic rocks, were superimposed on these terranes during the

Jurassic, Cretaceous, Eocene to late Miocene, and Late Miocene to Recent periods.

The Santa Rosa Gold Project is located within the Cajamarca-Valdivia terrane that includes a

metamorphic basement complex and the Antioquia Batholith. Lower greenschist- to lower amphibolite-

grade metasedimentary units and oceanic ophiolitic volcanic and intrusive rocks were accreted to the

continental margin in the Ordovician-Silurian periods. These rocks were onlapped by volcanic-

sedimentary units during the Mesozoic. During the Cretaceous period, both the Paleozoic basement

rocks and Mesozoic volcanic-sedimentary units were intruded by the Antioquia quartz-diorite, diorite, and

granodiorite batholith.

Monotonous diorite and quartz diorite (granodiorite) dominate the geology of the project area, with isolated

roof pendants of amphibolites and metasedimentary rocks, as well as dikes of microdiorite and dacite

porphyries. Red-brown saprolite is widespread and often deep (up to 50 m). Soils are generally about

50 cm thick and rarely up to 2 m. Schistose fault-zone mylonite was observed at several locations in

outcrop and adits.

Hypogene gold mineralization within the Santa Rosa Gold Project area is generally associated with shear

zones developed in homogeneous diorite country rock, with higher grades occurring in sulfide-mineralized

quartz veins. Exploration by Red Eagle Mining has identified a strongly mineralized shear zone (San

Ramon deposit shear zone) trending about east-west and containing mineralized quartz veins, sheeted

veins, and anastomosing vein networks. Besides gold, the mineralization contains some silver, significant

quantities of sphalerite, and minor amounts of galena. Sulfides range from 1% to 5% but can reach 10%,

and there appears to be a direct correlation between sulfide content and gold grade. The upper portions

of the mineralized shear zone are oxidized and heavily weathered to saprolite. Historically, gold has been

extracted on the property from these saprolitic, as well as colluvial, gold deposits, by artisanal miners from

underground workings, and from hydraulically mined areas known locally as “baticiones.”

Page 1.20




AMENDED NI 43-101 TECHNICAL REPORT 1.4 Mineral Resource Estimate

The two resource estimates performed for Red Eagle Mining at the San Ramon deposit have defined an

east-west-trending, shear zone-hosted gold deposit that extends over approximately 2,000 m of strike

length and that predominantly dips 70° to the north and is open at depth. A significant outcome of Red

Eagle Mining’s most recent work is the further development of a comprehensive geologic model, now

based on 238 core holes totalling 45,609 m, which provided the basis of the most recent resource

estimate update (two additional geotechnical / hydrogeological holes encountered mineralization but are

not included in the estimate because they were drilled after the effective date of the resource estimate).

Approximately 75% of the resource is Measured and Indicated, with the remainder Inferred; classification

is primarily a function of drill-spacing. The upgraded classification in the current resource is the result of

increased drill density with substantial infill drilling, as well as improvements in QA/QC procedures,

additional metallurgical testwork, and more comprehensive density and geotechnical data. The reported

estimate, presented in Table 1.1, was performed using inverse distance to the third and fourth power for

estimation of low and higher grade domains, respectively. MDA reports resources at cut-offs that are

reasonable for deposits of this nature given anticipated mining methods and plant processing costs, while

also considering economic conditions, because of the regulatory requirements that a resource exists “in

such form and quantity and of such a grade or quality that it has reasonable prospects for economic

extraction.”

The initial resource in 2012 (reported in January 2013) was reported at a single cut-off grade of 0.3g Au/t.

A subsequent resource in 2013 (reported in September 2013) was reported as a combination of open-pit-

mineable material at a cut-off grade of 0.3g Au/t and underground-mineable material at a cut-off of 1.2g

Au/t. The current resource estimate on which this study is based is the same as the one reported in

September 2013 but is reported at a cut-off grade for only underground mining of 1.2g Au/t. The current

resource estimate also includes a small addition on land not formerly controlled by Red Eagle Mining. The

change in reporting cut-off was appropriate because the mining scenario has been determined to be by

underground methods. Because of the change in cut-off, the resource tabulation summarized in Table 1.1

cannot be directly compared to the initial 2012 resource estimate or the subsequent 2013 resource

estimate.

Page 1.21




AMENDED NI 43-101 TECHNICAL REPORT Table 1.1 Mineral Resources

Total Measured

Cut-off

g Au/t Tonnes g Au/t oz Au

1.20 678,000 4.270 93,000

Total Indicated

Cut-off


1.20 3,475,000 3.46 386,000

Total Measured & Indicated

Cut-off


1.20 4,153,000 3.59 479,000

Total Inferred

Cut-off


1.20 1,524,000 2.72 133,000

1.5 Mineral Reserve Estimate

Mineral reserves were developed using the resource modelled high-grade domains along with undiluted

grade estimates. The high-grade domains were used as a basis for stope designs. To build up the

mineable reserves, the following steps were performed:

Expand the high-grade resource estimation domain polygons to a minimum mining width

of 2.5 m resulting in mid-block, level-plan polygons that represent potential stopes;

Estimate diluted stope grades using only Measured and Indicated undiluted grade

estimates from both high and low-grade estimation domains;

Apply economic parameters to calculate net value for each block;

Revise stope polygons to remove negative value blocks where possible;

Resources inside of each stope polygon were summarized and the economic value for

each stope was calculated;

Stope polygons with negative values were either modified or eliminated;

Centreline development was refined to access each stoping area;

Stope economics were reassessed to include an allocation of development costs

required to mine each stope;

Final refinement or elimination of stope polygons was completed; and

Page 1.22




AMENDED NI 43-101 TECHNICAL REPORT Solids of the final stope polygons were created and resources inside of each solid were

summarized.

Measured and Indicated resources above and below the economic cut-off grade (oxide 1.96g Au/t,

transition 2.14g Au/t, and sulfide 2.00g Au/t) were summarized within each stope solid. The average

minimum width is 3.0 meters. The mineral reserves are shown in Table 1.2 below:

Table 1.2 Total Proven and Probable Reserves by Material Type

Dilution by Measured and Indicated blocks included material below the mining cut-off

grade, but above the resource cut-off grades. This dilution totals 188,000 tonnes of

material grading 1.60 g Au/t and is part of the Proven and Probable reserves.

The stope solids include material that is classified as Inferred resource or was not

estimated, and while the tonnage of this material is included as dilution, no additional

metal content is included. This dilution totals 334,000 tonnes of material at zero grade

and is not part of the Proven and Probable reserves as presented above in Table 1.2.

On an overall tonnage basis, the total dilution included is approximately 23%.

The diluted Proven and Probable reserves are based on the same resources model that was used in the

Preliminary Economic Assessment of October, 2013 (PEA), with minor changes due to subsequent

additional land to the east of concession B7560005. The changes consider the original diluted material

processed from the PEA, removal of Inferred material, changes to design, and expansion of the minimum

mining width to include external dilution. The comparison is shown in Table 1.3.

Table 1.3 Changes from PEA Material Processed to Reserves

Oxide Mixed Sulfide Total Proven and Probable

K Tonnes g Au/t K Ozs Au K Tonnes g Au/t K Ozs Au K Tonnes g Au/t K Ozs Au K Tonnes g Au/t K Ozs Au

Proven 11 4.60 2 3 3.12 0 415 5.99 80 429 5.93 82

Probable 63 3.91 8 17 4.37 2 1,915 5.08 313 1,995 5.04 323

Proven & Probable 74 4.01 10 20 4.20 3 2,331 5.24 393 2,425 5.20 405

Tonnes g Au/t Ozs Au

PEA Material Processed 3,642,000 4.76 557,300

Remove Inferred (832,000) 4.16 (111,200)

Change due to Design (540,000) 4.26 (73,900)

Add External Dilution 489,000 2.09 32,900

Final Diluted Reserve 2,759,000 4.57 405,100

Page 1.23




AMENDED NI 43-101 TECHNICAL REPORT 1.6 Mining Methods

The San Ramon deposit has been planned as an underground mining operation.

Due to the experience and capabilities found in prospective Colombian contract miners, Red Eagle Mining

elected to use a contract mining group. A contract will be issued for a five year term, with an option to

extend, and with Red Eagle Mining retaining the right to take over the contract at any time should it prefer.

Development will include construction of a decline (which will serve as the access), main haulage drifts,

and attack ramps. Underground ventilation will require ventilation shafts to the surface along with raises

between levels.

The mining method selected is Mechanized Shrinkage with Delayed Fill (“MSDF”). The method is similar

to mechanized cut and fill but uses breast blasting of the back between lifts. Then, rather than mucking of

ore and backfilling immediately, the MSDF method leaves the ore in place in the stope, only removing

enough material from the stope to remove swell (similar to shrinkage stoping but removing the swell from

the top instead of the bottom). Access to the stope is provided by establishing an attack ramp. For each

lift, the attack ramp back is blasted to establish access for the subsequent lift. Support is placed in the

stope as needed during the drilling and blasting cycle of each lift. Once enough lifts have been drilled and

blasted, the ore will be mucked out completely. The last cycle of mining for MSDF stopes is backfilling of

the stope from the bottom up, progressively backfilling the attack ramp as well.

Backfill is to be used as required, but not all stopes will be fully backfilled. The backfill will consist of

filtered tailings and development waste rock. Tailings will be placed in a single lift, and once the lift of

tailings is placed, then a lift of waste rock from development will be placed on top of the tailings. The

amount of waste placed on top of the tailings will be sufficient to allow equipment to drive on top of the

backfill. It has been assumed that 60% tailings and 40% development waste rock will be a suitable

mixture of tailings to waste to stabilize the fill.

The benefits to using MSDF instead of cut-and-fill mining are:

Because ore is being slashed from the stope back and dropped on ore, dilution and ore

loss during mucking are greatly reduced;

Because MSDF slashing techniques are used for the bulk of the stope, the powder

factor required is reduced in comparison to cut-and-fill mining;

Leaving mined ore in the stope helps to maintain stability of the hanging wall and

footwall until the ore is mucked out;

If development of MSDF stopes is maintained ahead of ore requirements, underground

stockpiles of ore will be available that can be delivered as required to the mill;

Most of the fill only requires enough strength for equipment to be operated on top of it to

deliver more fill, which reduces cement requirements and costs; and

Page 1.24




AMENDED NI 43-101 TECHNICAL REPORT The method can easily be converted to cut-and-fill techniques where ground conditions

become weak.

Geotechnical

A geotechnical analysis of the San Ramon deposit and country rock characteristics was undertaken by

Golder. The conclusions to this study were that the quality for the hanging and foot walls is expected to be

good to very good with minimal support requirements for 20 m spans or less. Occasional support in the

form of bolting and shotcrete may be needed, particularly when the openings encounter a dyke or weak

zone. Conversely, for development along the poor quality shear zone, bolting and reinforcement will be

required at all stages of development.

Hydrological

Two water well drill and pump tests were undertaken to investigate water presence on both the

granodiorite country rock (within an identified fault zone), and the San Ramon deposit shear zone. The

information gained from the two water wells confirmed that the granodiorite country rock yields

insignificant flows and has low conductivity, and the shear zone similarly yields low flows and has low

conductivity, with inflow rates ranging from 0.2 to 2 L/sec. The mine design for pumping has been

conservatively sized for an initial inflow of 5 L/sec, rising to 10 L/sec as the operations increase with depth.

Development and Production Schedules

A significant amount of detail was incorporated into the scheduling of both the mine development and

stope production. Each area (totalling 106) and the development necessary to access the areas, was

incorporated into a monthly schedule for the mine life.

Table 1.4 shows the development schedule over the life of the mine.

Table 1.4 Mine Development Summary

Table 1.4 shows a total of almost 28 km of development drifts over the mine life, and averaging 3.5 km per

year. The main decline will eventually be 4.4 km in length at a 14% gradient.

Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Portal (Portal costs are in Capital) m 30 - - - - - - - - 30

Sublevel Ramp m 484 1,810 2,557 576 795 1,954 370 820 1,738 11,104

Haulage Drift m - 302 155 - 704 173 153 109 218 1,813

Main Ramp m 1,106 640 365 365 365 366 730 444 - 4,381

Ventilation Drift m 8 53 106 - 96 24 8 43 46 383

Ventilation Raise m - 171 200 - 147 - 63 57 177 815

Attack Ramp m - 938 239 2,696 1,634 1,098 721 610 1,184 9,120

Total m 1,627 3,914 3,622 3,637 3,742 3,614 2,044 2,084 3,362 27,646

Sublevel Ramp Tonnes 24,494 91,643 129,465 29,143 40,247 98,918 18,753 41,489 87,999 562,152

Haulage Drift Tonnes - 15,290 7,823 - 35,633 8,739 7,746 5,538 11,030 91,798

Main Ramp Tonnes 57,497 32,400 18,478 18,478 18,478 18,529 36,956 22,486 - 223,303

Ventilation Drift Tonnes 392 2,677 5,373 - 4,878 1,209 392 2,178 2,308 19,406

Ventilation Raise Tonnes - 3,018 3,529 - 2,604 - 1,106 1,014 3,130 14,400

Attack Ramp Tonnes - 14,658 3,740 42,125 25,536 17,150 11,260 9,534 18,494 142,497

Total Tonnes 82,384 159,686 168,408 89,746 127,376 144,545 76,213 82,239 122,960 1,053,556

Page 1.25





Table 1.5 shows the production schedule over the life of the mine.

Table 1.5 Production Schedule

Metallurgical recoveries and the resulting gold production from the process plant are also shown in this

Table 1.5. A total of 143,000 ounces are produced in the first 2 years.

1.7 Mineral Processing and Metallurgical Testwork

In 2014 the metallurgical program for this Feasibility Study was oriented towards samples that reflected

the range of head grades expected in the underground operation, namely 2 to 9 g Au/t. This program has

been conducted by McClelland Laboratories, Inc. in Sparks, Nevada. Three process options have been

tested in detail:

Whole ore cyanide leaching of finely ground ore in an agitated leach circuit;

Initial gravity concentration of the coarse gold, followed by grinding and agitated cyanide

leaching of the combined gravity tailings; and

A fairly coarse primary grind (P80 of 125 microns), followed by flotation of the sulfides.

The resulting rougher concentrate is then reground to a P80 of 20 microns, blended back

into the flotation tailings, and the combined product is cyanide leached in an agitated

circuit.

Mine Production Schedule Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Oxide K Tonnes 0 15 - 0 1 21 22 8 8 75

g Au/t 2.60 5.12 - 1.75 3.47 4.07 4.20 2.68 2.29 3.96

K ozs Au 0 2 - 0 0 3 3 1 1 10

Mixed K Tonnes - 6 0 1 1 1 1 9 0 20

g Au/t - 6.55 2.40 3.63 3.59 3.58 3.40 2.86 2.13 4.16

K ozs Au - 1 0 0 0 0 0 1 0 3

Sulfide K Tonnes 9 298 320 321 318 295 293 304 193 2,351

g Au/t 3.65 6.96 7.58 5.03 4.63 5.41 3.64 3.12 5.11 5.20

K ozs Au 1.0 66.8 78.0 51.9 47.3 51.2 34.3 30.5 31.8 393

Total K Tonnes 9 319 320 322 320 317 317 321 201 2,446

g Au/t 3.63 6.87 7.58 5.02 4.62 5.31 3.68 3.10 5.00 5.15

K ozs Au 1.0 70.5 78.0 52.0 47.5 54.1 37.5 32.1 32.3 405

Internal Waste K Tonnes 1 40 40 38 40 44 43 39 27 312

Total Diluted Material K Tonnes 10 359 360 360 360 361 360 360 228 2,759

Mined g Au/t 3.13 6.11 6.74 4.49 4.11 4.67 3.24 2.77 4.42 4.57

K ozs Au 1.0 70.5 78.0 52.0 47.5 54.1 37.5 32.1 32.3 405

Total K Tonnes - 352 360 360 360 361 360 360 245 2,759

g Au/t - 6.21 6.74 4.49 4.11 4.67 3.24 2.77 4.25 4.57

K ozs Au - 70 78 52 48 54 37 32 34 405

K Ozs Rec - 68 75 50 45 52 36 30 32 388

Net Rec 0% 96% 97% 95% 95% 96% 95% 95% 96% 96%

Page 1.26




AMENDED NI 43-101 TECHNICAL REPORT All three routes were capable of recovering between 93 to 96% of the gold. However, the flotation-

combined leach route has been selected as the preferred option, achieving optimum gold recovery. This

approach requires the least fine grinding (lowest grinding costs) and will recover an average of 96% of the

gold over the life of the operation. An estimated 69 % of the contained silver will also be recovered,

contributing to project revenue, though not included in this economic analysis.

Reagent consumptions in the float-combined leach process are as low as or lower than they are in the

other processes. Cyanide consumption averages 0.39 kilograms of sodium cyanide per tonne of ore and

lime consumption averages 1.6 kilograms per tonne of ore.

Additional testwork is still underway. This includes cyanide destruction (detox) tests using SO2 / air, initial

liquid / solid separation tests and carbon loading tests.

1.8 Recovery Methods and Process Plant Design

Based on the metallurgical testwork Lycopodium selected an overall process plant flowsheet which

includes grinding and flotation followed by concentrate regrinding. The flotation tailings and reground

concentrate are together leached in a CIL circuit. Cyanide in the CIL tailings will be detoxified using the

SO2 / Air process prior to the tailings being filtered. Part of the filtered tailings will be dry stacked in a dry

waste management facility (DWMF), the balance will be used as backfill in the mine. Filtrate will be

recycled back to the process plant to minimise the raw water requirement.

The process plant for the Santa Rosa Gold Project is based on a robust metallurgical flowsheet designed

for optimum recovery with minimum operating costs. The flowsheet is based upon unit operations that are

well proven in industry.

The key project and ore specific criteria that the plant design must meet are:

The plant is designed for an initial throughput of 360,000 tpa with provision for future

expansion to at least 720,000 tpa;

Testwork shows that the ore is of medium hardness with average head grades over the

life of the project of 4.57 g/t gold and 8.5 g/t silver;

Mechanical availability for the process plant of 91.3%;

A level of automation to reduce the technical complexity of the plant with manual

operation where practical;

Equipment selection for reliability and ease of maintenance; and

Layout for ease of access to all equipment for operating and maintenance requirements

whilst maintaining a compact footprint that will minimise construction costs.

The process plant design incorporates the following unit process operations:

Page 1.27




AMENDED NI 43-101 TECHNICAL REPORT Primary crushing with a single toggle jaw crusher (1,100 x 700 mm) to produce a

crushed product size of 80% passing (P80) 100 - 120 mm;

A crushed ore surge bin (30 m3) with a nominal capacity of 1 hour process plant feed of

45 t;

Single stage SAG mill (5.0 diameter x 3.5 m – 1,200 kW) in closed circuit with cyclones

to produce a P80 grind size of 125 micron;

Rougher scavenger flotation (6 x 8 m3 conventional cells) to produce a sulphides / gold

concentrate;

Tower mill (150 kW) for regrind of the concentrate to a P80 grind size of 15 - 20 µm;

Pre-leach thickener (10.3 m) to minimise carbon in leach (CIL) tankage and reduce

overall reagent consumption;

A hybrid CIL circuit incorporating one leach tank and six adsorption tanks (430 m3 each)

with 48 h total residence time;

A 2 tonne AARL elution circuit with electrowinning and smelting to produce doré bars;

Cyanide destruction using the SO2 / Air process (120 min retention); and

Tailings pressure filtration (445 m2 filter area) to 16% moisture content.

The level of instrumentation and control has been selected to provide a basic regulatory control to

maintain steady operation with minimal process excursions. Following industry practice for similar size

plants, a supervisory control and data acquisition (SCADA) and programmable logic controller (PLC)

architecture was selected for the plant wide process control system. It is a reliable and low cost approach.

Raw water for the Project is diverted from the spring source and catchment area of the La Veta creek to a

three day capacity storage pond located east of the plant site. Additionally, seasonal precipitation plus

surplus water from the La Veta creek system is collected in the seepage collection pond, which will be

pumped back to the process water storage tank at the plant. The seepage collection pond will have a

constant overflow to sustain the minimum flow rate requirement for the La Veta creek (7.2 m3/h). Raw

water is used to feed the potable and stripping water treatment plants as well as reagent mixing, and

gland water.

For a year with an average rainfall the runoff entering the project site will be on average between 22 to 44

m3/h. The estimated plant raw water make-up requirement will be approximately 14 m3/h. The

underground mining operation is assumed self-sustainable from recycled underground inflows.

Page 1.28




AMENDED NI 43-101 TECHNICAL REPORT 1.9 Project Infrastructure

The Santa Rosa Gold Project consists of an underground decline accessed mine, processing plant, and

plant infrastructure. As part of this Feasibility Study, a site plan was developed for the project site. Major

plant infrastructure consists of the following:

Main access road from the concession gate to portal pad;

44 kV power transmission line from EPM substation (8.9 km long);

44 kV switch yard;

Main switch room;

Reagents switch room;

Mine services and filter plant switch room;

DWMF;

Sedimentation pond;

Seepage pond;

Monitoring pond;

Event pond;

Reagent storage building;

Plant administration building;

Assay and metallurgical laboratory;

Plant workshop and main warehouse building; and

Mine truck shop/ warehouse buildings.

The proposed process plant site is bounded by the underground mining area to the south, the La Veta

creek valley to the east, and the ridge line separating the San Francisco valley from the La Veta creek

valley to the north and west. The topographic relief in the project area is moderate with gentle to relatively

steep-sided valleys and hills. Elevations range between 2,300 m and 2,500 m above sea level.

The following factors were considered when developing the site plan layout:

Page 1.29




AMENDED NI 43-101 TECHNICAL REPORT Minimizing the environmental impact and visibility;

Reduced haul truck travel distance from the mine portal area to the run of mine (ROM)

pad / primary crusher;

Usage of existing access roads;

Equalizing cut and fill material to avoid fill import or excessive cut to waste;

Ensuring all heavy equipment foundations are supported on native undisturbed soil with

sufficient load bearing capacity;

Optimized ore flow through the process plant;

Compact plant layout to reduce overall footprint, which minimizes electrical cabling,

piping and service roads thus reducing capital costs; and

Locating the filter presses as close as possible to the mine portal and the DWMF.

The major building structures will be made of structural steel with uninsulated roof and wall cladding. The

building foundations will consist of cast in-situ conventional reinforced concrete footings. Secondary

buildings will be of pre-engineered, pre-fabricated, or portable module type, where applicable. The

ancillary buildings will require varying degrees of air conditioning and ventilation. The process plant facility

will be entirely outdoors, and only the main control room, the electrical switch rooms, and the laboratory

building will be air conditioned. The gold room will be totally enclosed and ventilated. Fire protection,

lightning protection and smoke detection have been considered for various buildings.

The Santa Rosa Gold Project process plant will be supplied by a 44 kV overhead power line coming from

an existing EPM substation located approximately 9 km south of the plant site at Rio Grande. Figure 18.3

shows the EPM power substation and plant site with the proposed route for the new 44 kV power line.

The overhead power line and associated infrastructure will have sufficient capacity for the planned

expansion (10 MW).

The stepped down 4.16 kV power will be distributed to the plant switch rooms (load centres) using 5 kV

cables installed as a combination of cable trays and direct buried. The main switch room will supply power

for equipment located in the milling, feed preparation, and flotation areas. The reagents switch room will

supply power to leaching, desorption, goldroom, and reagents areas. The plant administration building and

the plant workshop and main warehouse will be fed from separate 4.16 kV lines and their dedicated 4,160

/ 480 V transformers.

Page 1.30




AMENDED NI 43-101 TECHNICAL REPORT A 4.16 kV direct buried cable will supply power to the 4.16 kV mine services switchgear, which will feed

the mine area operation, ventilation fans, and the filter plant switch room. The mine area operation and

ventilation fans will be fed via underground 4.16 kV lines. Each line will be protected by a 4.16 kV feeder

breaker. The step down transformers and their protection at the destination will be supplied by the mining

contractor and ventilation supplier. A dedicated feeder breaker will supply power via an oil filled outdoor

4,160 / 480 V, 500 kVA transformer to the filter plant switch room.

Colombia is serviced directly, or through trans-shipment, by ample carriers of all modes, and it has the

supporting infrastructure to receive major project cargo from offshore. Highway No. 25 from Medellin

through Santa Rosa de Osos is a major transport route to Caribbean ports.

The Santa Rosa Gold Project involves the delivery of major equipment from various parts of the world,

including Asia, North America and Europe. For this project, marine and truck transportation services will

be utilized, individually or in combination. Off-site laydown areas, marshalling areas and project

warehousing have been identified in the Medellin and Santa Rosa de Osos areas.

There is adequate road access to the site from the ports of Barranquilla and Buenaventura.

1.10 Capital Cost Estimate

1.10.1 Overall Capital Costs

The capital cost estimate includes all the direct and indirect costs and appropriate project estimating

contingencies for all the facilities required to bring the Santa Rosa Gold Project into production, as defined

by this Feasibility Study.

Mine capital has been minimized by employing a mining contractor for all mining activity. The largest

portion of the capital cost estimate is attributed to development costs, which have been based on

contractor quotations. Ventilation equipment quotations have been received from vendors. Other minor

equipment capital costs have been assumed based on InfoMine estimation guides.

In the process plant capital, all equipment and material are assumed to be new. The labour rate build up is

based on the statutory laws governing benefits to workers in effect in Colombia at the time of the estimate.

Colombian import tariffs have been applied. The estimate does not include any allowances for escalation,

exchange rate fluctuations or project risks. The capital cost estimate has a predicted accuracy of +/- 15%.

The total estimated cost of the overall project (mine plus process plant) is $ 69.90 million. This total has

been compiled as shown in Table 1.6.

Page 1.31




AMENDED NI 43-101 TECHNICAL REPORT Table 1.6 Santa Rosa Gold Project Overall Capital Cost Estimate ($M)

Main Area Supply Cost Installation Cost +Freight & Taxes

Total Cost

Construction Indirects 2.12 0.79 2.91

Treatment Plant 15.51 8.68 24.19

Reagents & Plant Services 3.22 1.56 4.78

Infrastructure 2.07 0.64 2.71

Mining 9.64 0.04 9.68

Construction Management Costs 5.70 0.41 6.11

Owner’s Costs 8.50 0.00 8.50

Working Capital 4.02 0.00 4.02

Subtotal 62.90

Contingencies 7.00

Grand Total 69.90

1.10.2 Mine Capital Costs

Mining capital has been minimized by employing a mining contractor for all mining activity. The mining

contractor will be required to provide the mining equipment, and the cost of the equipment will be

amortized into the mining cost. Mining capital includes development capital, pre-production mining costs,

and other mine capital that is comprised of portal collar work, contractor mobilization, and mine surface

facilities.

Most of the equipment and supplies that supports the mine will be purchased and installed by the

contractor with reimbursement by Red Eagle Mining. The exception is some equipment that Red Eagle

Mining will purchase from other vendors. This includes primary and auxiliary ventilation fans, dewatering

pumps, compressors, substation, and electrical switching gear.

The total capital is shown in Table 1.7. Pre-production capital and pre-production expensed costs are

included in the initial capital. Total initial mine capital is $9.44 million, which includes pre-production costs

of ore mined and stockpiled prior to the start of production. Mining costs that are normally expensed are

included in the initial mine capital and include development cost for mining attack ramps to stoping areas

during the pre-production period along with delineation drilling. Development metres have a contingency

of 10% applied to the design to allow for unaccounted for (miscellaneous) development.

Table 1.7 shows the mining capital costs.

Page 1.32





Table 1.7 Total Mine Capital ($M)

Item Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Mine Capital Summary

Mine Development 3.50 5.26 5.98 1.70 3.75 4.42 2.33 2.59 3.90 33.45

Initial Delineation Drilling 0.43 0 0 0 0 0 0 0 0 0.43

Pre-Production 2.97 0 0 0 0 0 0 0 0 2.96

Contractor Capital 0.78 0 0 0 0 0 0 0 0.02 0.80

Departmental Capital 1.77 1.21 0.34 0.22 0.53 0.34 0.24 0.04 0.01 4.68

Total Mining Capital 9.44 6.47 6.32 1.92 4.28 4.76 2.57 2.63 3.93 42.32

Total Mining Capital

Initial Capital 9.44 0 0 0 0 0 0 0 0 9.44

Sustaining Capital 0 6.47 6.32 1.92 4.28 4.76 2.57 2.63 3.93 32.88

Total Capital 9.44 6.47 6.32 1.92 4.28 4.76 2.57 2.63 3.93 42.32

Page 1.33




AMENDED NI 43-101 TECHNICAL REPORT 1.10.3 Plant and Site Infrastructure Capital Costs

Major plant and site infrastructure included in the capital costs consists of the following:

Process plant;

44 kV power line and switch yard;

Access road;

Ponds;

Ancillary buildings; and

DWMF.

The total estimated cost of the process plant is $59.25 million as shown in Table 1.8. This total excludes

the initial mining capital investment of $9.44 million.

Table 1.8 Process Plant Estimated Total Installed Costs ($M)

Main Area Supply Cost Installation Cost + Freight and Taxes

Total Cost






Owner’s Costs 8.50 0.00 8.50


Subtotal 53.22

Contingencies 6.03

Grand Total 59.25

Contingency is an integral part of an estimate and has been applied to all parts of the estimate, direct

costs, indirect costs, owners’ costs, etc. The average contingency for the process plant capital costs is

11.2%.

Page 1.34




AMENDED NI 43-101 TECHNICAL REPORT 1.11 Operating Cost Estimate

1.11.1 Mine Operating Costs

Mine operating costs have been estimated based on first principle operating parameters and costing

parameters using hourly equipment and personnel rates provided by contractor quotations. Note that

operating costs estimated during pre-production have been included in the capital costs, but have been

estimated in the same manner. Mine operating costs are shown in Table 1.9.

Electrical and fuel consumption rates were determined using InfoMine estimation guides. Owner operated

equipment cost estimates included ventilation and dewatering costs, which were estimated using hourly

rates from InfoMine. Electrical and fuel costs have been assumed to be $0.11/kWH and $1.10/L,

respectively. This is based on Red Eagle Mining’s input and research.

The overall average mine operating cost is estimated to be $37.36 per tonne, or $265.61 per ounce of

gold produced.

The accuracy of the mine operating cost estimate is expected to be within ± 15%.

Page 1.35





Table 1.9 Mine Operating Cost

Mine Production Costs Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Drilling K USD 960 962 962 962 965 962 962 609 7,346

Blasting K USD 1,615 1,619 1,619 1,619 1,623 1,619 1,619 1,038 12,369

LHD Haulage to Muck Bays K USD 788 813 835 783 782 745 715 453 5,915

LHD Loading K USD 170 170 170 170 171 170 170 141 1,332

Haulage to Surface K USD 673 915 977 893 890 877 1,028 938 7,192

Ground Support K USD 1,196 1,199 1,199 1,199 1,202 1,199 1,199 759 9,153

Backfill K USD 796 914 930 909 898 924 889 685 6,945

Expensed Development K USD 1,636 418 4,703 2,851 1,915 1,257 1,064 2,065 15,908

Mine Support Services K USD 358 357 357 357 358 357 357 298 2,801

Delineation Drilling Sampling K USD 1,483 992 726 727 726 726 726 309 6,414

Ventilation K USD 579 527 533 526 528 525 518 402 4,138

Dewatering K USD 80 86 85 104 104 104 102 59 725

Mine General Services K USD 982 982 982 982 982 982 982 818 7,688

Subtotal Mine Cost K USD 11,318 9,953 14,078 12,082 11,144 10,448 10,331 8,573 87,927

Contractor Fixed Cost K USD 1,931 1,931 1,931 1,931 1,931 1,931 1,931 1,609 15,128

Total Mine Cost K USD 13,249 11,884 16,010 14,013 13,075 12,379 12,263 10,182 103,055

Total Mine Cost $/t 36.70 33.01 44.47 38.93 36.22 34.39 34.06 43.94 37.36

Page 1.36




AMENDED NI 43-101 TECHNICAL REPORT 1.11.2 Plant Operating Costs

The operating cost estimate includes five major categories as defined below:

Process plant labour;

Operating consumables;

Power;

Maintenance; and

General and administration.

Operating costs have been developed using the plant parameters specified in the process design criteria.

The operating cost estimate includes all the cost items relevant to processing the ore by crushing and

grinding, flotation, CIL, electrowinning and smelting to produce gold doré. The operating costs listed by

major category are presented in Table 1.10.

The process plant operating cost was determined to be $24.11 per tonne of ore processed or $171.48 per

ounce of gold produced. With inclusion of G&A costs of $9.68 per tonne, the total annual operating cost is

$33.80 per tonne of ore processed or $240.35 per ounce of gold produced.

Table 1.10 Summary of Process Operating Cost Estimate at nominal duty of 360,000 tpa

Cost Category Total Cost Distribution

% $/y $/t ore $/oz Au

Process Plant

Process Plant Labour 2,122,822 5.90 41.94 17%

Operating Consumables 2,958,244 8.22 58.44 24%

Power 2,654,988 7.37 52.45 22%

Maintenance 944,027 2.62 18.65 8%

Subtotal - Process Plant 8,680,082 24.11 171.48 71%

General & Administration 3,486,559 9.68 68.88 29%

Total 12,166,641 33.80 240.35 100%

1.12 Economic Analysis

An economic analysis was developed for the Santa Rosa Gold Project using the production schedule

along with capital and operating costs as described in Section 21. The analysis uses fully diluted Proven

and Probable reserves. The generated cash-flow model was carried out on a pre-tax and post-tax basis.

Page 1.37




AMENDED NI 43-101 TECHNICAL REPORT Net gross revenues were estimated using a $1,300 per ounce gold price with revenue deductions for

royalties, refining, transportation and insurance costs. Cash cost calculations use the total operating cost

plus royalties divided by the payable gold ounces. Total costs are calculated using the total costs

(operating and capital) plus royalties and taxes divided by the payable gold ounces.

The cash-flow model calculates the Net Present Value (NPV) based on a discounted rate of 0%

(undiscounted), 5% and 8%. The base case considers the NPV at 5%. The Internal Rate of Return (IRR)

on total investment and the payback period were also calculated.

A sensitivity analysis was also conducted on parameters that are deemed to have the biggest impact on

the Project financial performance (capital cost, operating cost and gold selling price). The financial results

are summarised in Table 1.11.

Table 1.11 Financial Performance Indicators

Item Units Pre-Tax Post-Tax

NPV @ 0% K USD 171,763 131,501

NPV @ 5% K USD 136,895 103,678

NPV @ 8% K USD 119,921 90,027

IRR % 64.4% 52.6%

Payback Years 1.3 1.3

Cash Costs US$/t ore 83.78

Total Costs US$/t ore 94.24 107.27

Cash Costs US$/oz Au 596.12

Total Costs US$/oz Au 670.56 763.30

1.13 Conclusions and Recommendations

1.13.1 Geology and Mineral Resource

The Santa Rosa Gold Project is located within the Cajamarca-Valdivia terrane that includes a

metamorphic basement complex and the Antioquia batholith. Typically, structures within the Antioquia

batholith trend northwest-southeast. These structures are interpreted to be the cause of the formation of

San Ramon deposit shear zone which, under dilation has received quartz, precious metals and sulphide

mineralisation. Although mineralisation in the shear zone is pervasive, there are distinct high grade

domains that can be followed consistently along the known east-west strike length of the deposit.

The current resource estimate on which this Feasibility Study is based was reported in August 2013 but is

now reported at a different cut-off grade 1.2 g Au/t. The change in reporting cut-off was appropriate

because the mining scenario has been determined to be by underground methods. The current resource

estimate also includes a small addition on land not formerly controlled by Red Eagle Mining.

Page 1.38




AMENDED NI 43-101 TECHNICAL REPORT The following recommendations have been proposed:

Further exploration drilling is required to test the current Inferred resource, and to test

the expansion of the deposit with depth and plunge to the east; and

Detailed delineation drilling will be necessary for ore zone definition and stope design

prior to mining. The delineation drill pattern should be a closely spaced as practically

possible, and initially, no greater than a 10 m by 10 m pattern.

1.13.2 Mining and Mineral Reserves

Mineral reserves were developed using the resource modelled high-grade domains along with undiluted

grade estimates. The high-grade domains were used as a basis for stope designs. Measured and

Indicated resources above and below the economic cut-off grades (oxide 1.96 g Au/t, transition 2.14 g

Au/t, and sulfide 2.00 g Au/t) were determined for the stope solids. Initial stope designs were created with

a 2.0 m minimum width, but then expanded to include a 2.5 m width to account for external wall dilution or

selvedge. Dilution with Measured and Indicated material was added at modelled grades; waste and

Inferred dilution material was added at zero grade. The average minimum width is 3 meters.

A mining method was developed that was considered suitable to the geometry of the stope solids, and the

geotechnical characteristics of the shear zone containing the high grade domains. The method developed

was Mechanized Shrinkage with Delayed Fill (MSDF). This method has the advantages of:

dilution and ore loss during mucking are greatly reduced;

the powder factor required is reduced in comparison to the similar method of cut-and-fill

mining;

leaving mined ore in the stope maintains wall stability until the ore is mucked out;

underground stockpiles of ore will be available that can be delivered as required to the

mill; and

easy to convert to cut-and-fill techniques where ground conditions become weak.

Life of mine production schedules were created based on required development and the designed stopes.

The schedules were done on a monthly basis and include productivity from contractor quotations. These

schedules were used for first principle cost evaluation and economic analysis.

Reserves have been classified in order of increasing confidence into Proven and Probable categories to

be in compliance with the “CIM Definition Standards - For Mineral Resources and Mineral Reserves”

(2014) and therefore Canadian National Instrument 43-101.

Page 1.39




AMENDED NI 43-101 TECHNICAL REPORT The following recommendations are proposed:

Maximum support design for the decline portal and the upper section of the decline in

saprolite and schist is recommended. The decline is the only access into the deposit and

must remain stable for the life of the mine. Special support attention may also be

necessary when advancing from the schistose rock through to granodiorite;

The Feasibility Study assumes mining contractors will be used to offset mining capital.

Proposals were received as part of the mining study, and several discussions have been

underway with multiple mining contractors. These negotiations need to be completed

with the goal of developing a working partnership with the contractor that will promote

productivity and safety while maintaining costs at a minimum;

Mining contractors in the area are very experienced in underground construction and

development, but are less experienced in mine production. Red Eagle Mining has

decided to bring in a team of experienced mining experts to help train the contract

miners. To make this strategy successful, these mining experts need to be sourced;

Management of dilution and ore loss will be the key to the success of the underground

operations. This will require planning and implementation of a delineation program and

the procedures to be used. Red Eagle Mining should determine the procedures and

work flow required to ensure a successful program of delineation for stopes. This

includes how drill planning is done, how drilling results are tracked, establishment of

quality control protocols and procedures, modelling of results, inclusion into detailed

stope designs, and reconciliation of planned and actual mining; and

Explosives are supplied by the Colombian Government and are long lead orders.

Negotiations with the contractor need to be completed and orders placed to ensure that

explosive products are available when required.

1.13.3 Metallurgical testwork and Recovery Methods

Metallurgical laboratory testwork achieved the desired quality and demonstrated that by using the

designed process flowsheet it is possible to economically recover gold from the San Ramon deposit.



concentrate are leached in a CIL circuit. Cyanide in the CIL tailings will be detoxified using the SO2 / Air

process prior to the tailings being filtered. Part of the filtered tailings will be stacked in a dry waste

management facility; the balance will be used as backfill in the mine. Filtrate will be recycled back to the

process plant to minimise the raw water requirement.




Page 1.40




AMENDED NI 43-101 TECHNICAL REPORT To compliment the results of the Feasibility Study, additional metallurgical testwork is recommended as

follows:

Bulk flotation testwork to generate 10 to 20 kg of flotation concentrate using a 100 to

200 kg sample. It is recommended that the sample is extracted from the existing Hilo

Azul underground cross cut, which intersects the full width of the shear zone in

sulphides;

Regrind testwork should be undertaken by the selected equipment vendor using the

flotation concentrate. This testwork will confirm the size of the regrind equipment, refine

the operating cost estimate and will facilitate process guarantees upon purchase of the

equipment;

The bulk flotation tail and reground product should be collected after testing. The

samples should be sent for vendor thickening and filtration testwork to facilitate process

guarantees upon purchase of the equipment;

Given the lower cyanide consumption and lower final cyanide concentration in the coast

down tests, further testing is recommended to optimize the cyanide concentration and

dosage schedule during the leach step; and

A single test with separate leaching of the reground concentrate and flotation tails

increased gold recovery by more than 1% over the combined leach. Therefore it is

recommended that the test be repeated to confirm the improved recovery. If this is

confirmed, then a value added exercise should be undertaken to determine if the

separate leach approach offers a significant improvement in the process.

In addition to metallurgical testwork the following modelling and value engineering assessments are

recommended:

Value engineering assessment of the inclusion of a tail thickener to recover cyanide

should be undertaken. The current design minimises capital rather than operating cost;

and

Given the closed circuit nature of the flowsheet it is recommended that a more detailed

mass and chemical modelling of the circuit be undertaken including both the plant and

site water system to better define the chemistry of circulating loads and the final effluent.

1.13.4 Project Infrastructure

Site infrastructure facilities in support of the mining and processing of the San Ramon deposit have been

developed to take into consideration the local topographic features, water courses and access. The level

of the detail and planning is commensurate with that normally associated with Feasibility Study level.

Page 1.41




AMENDED NI 43-101 TECHNICAL REPORT There is adequate space for the intended facilities and there are no known adverse conditions that could

affect the design and construction of the required project infrastructure. The layouts of equipment and

building sizes including auxiliary buildings, maintenance facilities, and power supply infrastructure will be

compatible with other similar size projects.

The current level of infrastructure design incorporates all necessary facilities to protect the environment

including water treatment, sewage disposal, surface water catchment, retention ponds and recirculation

systems.

Due to the reduction in the project site footprint and process plant relocation late in the Feasibility Study,

an extension to the completed geotechnical program needs to be undertaken for the equipment and

building foundations in the current location of the plant site to supplement the drilling done as part of this

Feasibility Study. The program must be completed prior to the start of detailed engineering.

1.13.5 Environmental Studies, Permitting, and Social and Community Impact

Red Eagle Mining needs to obtain the necessary Environmental License in order to start construction

activities at the Project site. The social program of information and workshops that have been running for

over two years should continue through the duration of operations. The corporate policy of Red Eagle

Mining is to remain transparent to all communities and stakeholders in the region.

Red Eagle Mining has a number of sustainable programs planned that will positively develop the social

environment. This will be achieved by the existing commitment to maximize the recruitment of local

personnel during both construction and operations. It is estimated that at least 150 people will benefit from

direct employment and at least another 500 will benefit indirectly. Local suppliers of goods and support

services to the mine will be able to develop their businesses through increased opportunities. Education

and training programs will raise the skills level of the workers in the area. Culturally the local communities

are willing to accept change and have a high potential and desire to improve and benefit themselves.

The project footprint lies within an area owned by a single landowner and therefore there is limited effect

on land or property tenure.

There are no impacts related to displacement of the local population and any impacts on their way of life

are minimal and management plans are in place to mitigate these. The remote location of the project in

relation to the local communities and population assists in minimizing any potential impacts.

1.13.6 Main Conclusion

The Feasibility Study for the Santa Rosa Gold Project has been completed in sufficient detail to refine the

economics to a +/-15% level of accuracy and outline the issues facing the project going forward. The

Project economics are sufficiently robust to warrant moving to the next phase of detailed engineering and

construction. The estimated cost for the entire mine development and process plant and infrastructure

construction up to, and including start-up and commissioning is $69.90 million (excluding VAT to be paid,

but reclaimed in the first year of production).

Page 2.1





2.0 INTRODUCTION

2.1 General

The following Technical Report was compiled by Lycopodium Minerals Canada Ltd. (Lycopodium) and

presents the results of the Feasibility Study for the Santa Rosa Gold Project in the Department of

Antioquia, northern Colombia. The Technical Report was prepared at the request of Red Eagle Mining

Corporation (Red Eagle Mining), a British Columbia corporation. Red Eagle Mining is listed on the TSX

Venture exchange (RD) and the OTCQX (RDEMF).

The purpose of this Technical Report is to support the public disclosure of a Feasibility Study for the Santa

Rosa Gold Project. The resource estimate, on which this Feasibility Study is based, was previously

described in a Technical Report by Mine Development Associates (Lindholm and Schlitt, 2013b). This

Technical Report, the resource and reserve estimate, and the Feasibility Study have been prepared in

compliance with the disclosure and reporting requirements set forth in the Canadian Securities

Administrators’ National Instrument 43-101 (“NI43-101”), Companion Policy 43-101CP, and Form

43-101F1, as well as with the Canadian Institute of Mining, Metallurgy and Petroleum’s “CIM Definition

Standards - For Mineral Resources and Reserves, Definitions and Guidelines” (“CIM Standards”) adopted

by the CIM Council on May 10, 2014.

This Feasibility Study was completed by:

Lycopodium Minerals Canada Ltd. (“Lycopodium”) for the process design, plant

infrastructure, Capital and Operating cost estimates, and Economic analysis;

Mine Development Associates (“MDA”) for mining and geology;

Hydrometal Inc. (Hydrometal”) for metallurgical testwork management, largely

undertaken at McClelland Laboratory Services;

Golder Associates South America Ltd. (“Golder”) for the design of the dry waste

management facility, and geotechnical and hydrogeological aspects of the project; and

Unless otherwise denoted, all costs referred to in this Feasibility Study are quoted in current 3Q United

States dollars.

The effective date of the mineral resource estimate is August 5, 2013. The effective date of the Feasibility

Study is October 6, 2014. The effective date of this report is October 6, 2014.

Page 2.2




AMENDED NI 43-101 TECHNICAL REPORT 2.2 Sources of Information

Lycopodium has based its interpretation on the data and information provided by Red Eagle Mining for the

completion of this report including prior Technical Reports by MDA. Sections 6, 7 and 9 are based on

information provided by Red Eagle Mining, including the Technical Report by Jemielita (2011b).

Observations by MDA are also included in Section 7. Section 8 was condensed from Jemielita (2001b)

and includes additional information provided by Red Eagle Mining. The information provided by Red Eagle

Mining and other references are listed in Section 27.0 'References'.

2.3 Qualified Persons

The following Qualified Persons (QP), as the term is defined by NI43-101, have contributed to the writing

of this Report and have provided the requisite QP certificates included at the end of this Technical Report.

The information contained in the certificates pertains to each section of this Technical Report that the

designated QP is responsible for.

Mr. Stefan Gueorguiev P. Eng. Lycopodium Minerals Canada Ltd

Mr Gueorguiev is the Manager of Projects at Lycopodium, and is the QP responsible for matters relating to

the process plant and infrastructure sections of the report together with the sections related to capital and

operating costs and the economic analysis as well as the general sections of the document. Mr

Gueorguiev was also responsible for the coordination, consolidation and review of Sections 1, 2, 3, 4, 5,

19, 23, 24, 25, 26 and 27.

Mr. Stefan Gueorguiev is also the QP responsible for Section 20 of the Report and in his supervision of

the preparation of that section relied entirely on the Environmental Impact Study dated February 20, 2014

provided to the Issuer and prepared by Alejandra Maz Courrau, of the firm Tetra Tech Colombia S.A.S.

Mr. Thomas L. Dyer P. E. Mine Development Associates

Mr. Thomas L. Dyer, P.E., is a Senior Engineer with MDA and is the QP for mineral reserves, mining

methods, and the mining portion of capital and operating costs sections of this report.

Mr. Michael Lindholm C.P.G. Mine Development Associates

Mr, Lindholm is a Senior Geologist at MDA and is the QP responsible for matters related to geology and

resources. This includes history, geological setting and mineralization, deposit types, exploration, drilling,

sample preparation, analyses and security, data verification, and mineral resource estimates

Mr. W. Joseph Schlitt Ph.D., P. Eng Hydrometal Inc.

Mr. Schlitt is the President of Hydrometal and is the QP responsible for the mineral processing and

metallurgical testing sections of this report.

Page 2.3




AMENDED NI 43-101 TECHNICAL REPORT Mr. Terry Eldridge P.Eng Golder Associates South America Ltd.

Mr Eldridge is the South America mine waste management leader of Golder Associates and is the QP

responsible for matters relating to the dry waste management facility section of the report.

Table 2.1 outlines responsibility for the various sections of the Report.

Page 2.4





Table 2.1 Technical Report Section List of Responsibility

Section

Number Section Title Responsible QP Co-author

Other

Experts

1 Summary Lycopodium (Stefan Gueorguiev)/ MDA (Thomas L. Dyer, Michael Lindholm)/ Joe Schlitt

/ Golder (Terry Eldridge) REM

2 Introduction Lycopodium (Stefan Gueorguiev)

3 Reliance on Other experts Lycopodium (Stefan Gueorguiev)

4 Property Description and location Lycopodium (Stefan Gueorguiev) REM REM

5 Accessibility, Climate, Local Resources, Infr. and Physiography Lycopodium (Stefan Gueorguiev) REM, Golder REM

6 History MDA (Michael Lindholm) REM

7 Geological Setting and Mineralization MDA (Michael Lindholm) REM

8 Deposit Types MDA (Michael Lindholm) REM

9 Exploration MDA (Michael Lindholm) REM

10 Drilling MDA (Michael Lindholm) REM

11 Sample Preparation, Analyses and Security MDA (Michael Lindholm) REM

12 Data Verification MDA (Michael Lindholm)

13 Mineral Processing and Metallurgical Testing Hydrometal Inc. (Joe Schlitt) REM

14 Mineral Resource Estimates MDA (Michael Lindholm) REM

15 Mineral Reserve Estimates MDA ( Thomas L. Dyer )

16 Mining Methods MDA ( Thomas L. Dyer )

17 Recovery Methods Lycopodium (Stefan Gueorguiev)

18 Project Infrastructure Lycopodium (Stefan Gueorguiev)/ Golder (Terry Eldridge) REM

19 Market Studies and Contracts Lycopodium (Stefan Gueorguiev) REM

20 Environmental Studies, Permits and Social or Community Impact Lycopodium (Stefan Gueorguiev) REM

21 Capital and Operating Costs Lycopodium (Stefan Gueorguiev)/ MDA ( Thomas L. Dyer )/ Golder (Terry Eldridge REM

22 Economic Analysis Lycopodium (Stefan Gueorguiev) REM

23 Adjacent Properties Lycopodium (Stefan Gueorguiev) REM

24 Other Relevant Data and Information Lycopodium (Stefan Gueorguiev)

25 Interpretation and Conclusions Lycopodium (Stefan Gueorguiev)/ MDA ( Thomas L. Dyer; Michael

Lindholm)/Hydrometal Inc. (Joe Schlitt)/ Golder (Terry Eldridge) REM

Page 2.5





Section

Number Section Title Responsible QP Co-author

Other

Experts

26 Recommendations Lycopodium (Stefan Gueorguiev)/ MDA ( Thomas L. Dyer ; Michael Lindholm)/

Hydrometal Inc. (Joe Schlitt)/ Golder (Terry Eldridge) REM

27 References Lycopodium (Stefan Gueorguiev)/ MDA ( Thomas L. Dyer ; Michael Lindholm)/

Hydrometal Inc. (Joe Schlitt)/ Golder (Terry Eldridge) REM

Page 2.6




AMENDED NI 43-101 TECHNICAL REPORT 2.4 Site Visits

Mr. Dyer and Mr. Lindholm made a site visit to the Santa Rosa project on September 25 through 29, 2012.

MDA reviewed current drill programs, drilling/logging/sampling procedures, pertinent geology, and data

availability. In addition, MDA reviewed drill core from San Ramon, collected independent confirmation

samples of core and coarse rejects, spot-checked collar locations, and visited potential mining facility

locations.

Mr. Dyer made a second visit to the site along with Mr. Schlitt and Mr. Eldridge from February 19 to 22,

2013, touring potential infrastructure locations, viewing core, and discussing development strategies. In

addition, on several occasions Mr. Schlitt has visited all the metallurgical laboratories that performed a

significant amount of testwork on samples from San Ramon deposit.

Mr. Lindholm made a second site visit to the project on May 14 through 17, 2013, during which he

examined high-grade intercepts in core, spot-checked new collar locations, and gathered the newest

drilling data.

Mr. Gueorguiev visited the Santa Rosa Project site on February 19, 2014. The primary objective of the site

visit was to verify the site conditions, existing infrastructure, and utilities in the area surrounding the mine

site.

Page 3.1




AMENDED NI 43-101 TECHNICAL REPORT 3.0 RELIANCE ON OTHER EXPERTS

The authors are not experts in legal matters, such as the assessment of the legal validity of mining

concessions, private lands, mineral rights, and property agreements. The authors did not conduct any

investigations of the environmental or social-economic issues associated with the Santa Rosa Gold

Project, and the authors are not experts with respect to these issues.

The authors rely on information provided by Red Eagle Mining as to the legal status of Red Eagle Mining

and related companies, as well as surface rights comprising the Santa Rosa Gold Project, material terms

of all property agreements, and the existence of applicable royalty obligations. Lycopodium has relied on

Red Eagle Mining and their independent Colombian legal counsel, Lloreda Camacho & Co, for current

legal title of the concessions. Section 4 is based on information provided by Red Eagle Mining and

Lloreda Camacho & Co, and the authors offer no professional opinions regarding the provided information.

Lycopodium has relied on Red Eagle Mining and their financial staff and advisors to determine appropriate

tax implications in the financial analysis for this Technical Report. Lycopodium is not an expert on

Colombian tax issues.

The Environmental Impact Assessment (EIA), which has been summarised in this Feasibility Study

(Section 20) was prepared by Tetra Tech, along with the full support of the Universidad de Antioquia

(University of Antioquia) and Fundación Universitaria Católica del Norte (Foundation Catholic University of

the North) in compliance with requirements and specifications provided by the Environmental Authority.

Page 4.1




AMENDED NI 43-101 TECHNICAL REPORT 4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 Property Location

The Santa Rosa Gold Project is located approximately 20 km southeast of the town of Santa Rosa de

Osos, in the municipality of the same name, in the Department of Antioquia, 73 km northeast of the

department capital Medellín in northern Colombia (Figure 4.1). The centre of the resource is located at

approximately latitude 6° 36' 57 N and longitude 75° 22' 20 W. The San Ramon deposit lies in the south

eastern part of concession B7560005 held by Red Eagle Mining and described in more detail in

Section 4.3.

Figure 4.1 Location of the Santa Rosa Project

4.2 Colombian Mining Law Regarding Concession Contracts

The following information is provided by Red Eagle Mining; Lycopodium has no expertise in Colombian

law and has not confirmed this information.

Santa Rosa Gold Project Site

Page 4.2




AMENDED NI 43-101 TECHNICAL REPORT All mineral resources belong to the state and can be explored and exploited by means of concession

contracts granted by the state. Under Colombian mining law, a concession contract consists of

exploration, construction, and exploitation terms. Once the contract is registered, the exploration term is

three years, renewable for an additional eight years in increments of two years. The concession contract

is then convertible to the construction term, which is two years, renewable for an additional year. Finally

the concession contract is convertible to an exploitation term upon the granting of an Environmental

Licence by the Environmental Authority.

The total term for the concession contract (exploration and exploitation) is 30 years renewable for a further

30 years, unless the concession contract was signed under Law 1382 of 2010, in which case the

concession contract is renewable for a further 20 years.

Producing mines are subject to a federal royalty of 4% of the gross value of gold and silver production at

80% of the current London gold price which is in effect 3.2% (under a modification of mining law 685 of

2001).

Legal surveys are conducted to define the coordinates of the concession by the Secretaria de Minas de

Antioquia (Secreminas, or mines department) prior to granting the concession contract.

4.3 Land Area

The Santa Rosa Gold Project comprises a total area of approximately 33,000 hectares and consists of 12

concession contracts and 14 concession applications for which the technical studies have been approved.

Red Eagle Mining holds a 100% interest in the properties comprising the Santa Rosa Gold Project, subject

to completion of property payments described in Sections 4.4.1, 4.4.2, 4.4.3 and 4.4.4, and 4.4.5 to royalty

obligations described in Section 4.4.6. Figure 4.3 shows the concessions in the immediate vicinity of the

San Ramon deposit, located in the south eastern part of concession B7560005, which is the focus of this

Technical Report.

The annual holding costs for the concessions are approximately $12 per hectare.

Red Eagle Mining does not currently control any surface rights on the San Ramon deposit, which are held

by a single land owner. Red Eagle Mining has paid a monthly fee for surface access to the surface land

rights owner while conducting exploration activities during the past (almost) 4 years. A preliminary

footprint for potential mining operations has been determined and discussions with the surface rights

owner to extend the lease for the life of the mine are nearing conclusion.

Page 4.3




AMENDED NI 43-101 TECHNICAL REPORT Table 4.1 Legal Status of Red Eagle Mining’s Mining Concession Contracts and

Applications

LEGAL STATUS OF MINING CONCESSION CONTRACTS AND APPLICATIONS

ID Proposal Concession

Contract

Technical

Study

Area

(ha) Holder Contract

Currently

Lapsed

Remaining

Time Extension

1 H5790005 X 270 REMDC 30 Years 10 Years 20 Years 30 Years

2 H5791005 X 220 REMDC 30 Years 9 Years 21 Years 30 Years

3 B7560005* X 500 REMDC 30 Years 2 Years 28 Years 30 Years

4 B7171005 X 498 REMDC 30 Years 1 Year 29 Years 30 Years

5 LDM-08061 X 1,630 Perez

6 LKA-08004 X 52 Perez

7 OG2-081816 X 1,809 REMDC

8 OG2-08229 X 121 REMDC

9 JC3-08091 X 4 Grupo de Bullet 30 Years 2 Years 28 Years 30 Years

10 ICQ-0800643X X 164 Grupo de Bullet 30 Years 3 Months 30 Years 20 Years

11 JC3-08092X X 8 Grupo de Bullet 30 Years 1 Year 29 Years 30 Years

12 JI8-08071 X 6,113 Grupo de Bullet 30 Years 5 Months 30 Years 20 Years

13 JIT-08461 X 8,590 Grupo de Bullet 30 Years 1 Year 29 Years 20 Years

14 KGM-14151 X 3,962 Grupo de Bullet

15 KGM-14153X X 15 Grupo de Bullet

16 KGM-14152X X 506 Grupo de Bullet

17 KGM-14241 X 1,796 Grupo de Bullet

18 KGM-14242X X 4,969 Grupo de Bullet

19 KIG-11151 X 10 Grupo de Bullet

20 LIN-11551 X 366 Grupo de Bullet

21 7590B X 352 AngloGold

Ashanti 30 Years 3 Years 27 Years 30 Years

22 7591 X 281 AngloGold

Ashanti 30 Years 1 Year 29 Years 30 Years


Ashanti 30 Years 4 Years 26 Years 30 Years


Ashanti

25 7723 X 693 AngloGold

Ashanti

26 HBS-10501X X 30 AngloGold

Ashanti

TOTAL PROJECT AREA 33,224

*San Ramon Gold Deposit

Page 4.4





Figure 4.2 Concessions and Applications Forming the Santa Rosa Gold Project

Page 4.1




AMENDED NI 43-101 TECHNICAL REPORT Figure 4.3 Concessions and Applications in the Immediate Vicinity of the San Ramon

Deposit

4.4 Agreements and Encumbrances

In 2010 Red Eagle Mining initially acquired four concession contracts and one application that had passed

technical review by the Mining Authority (Secreminas) totalling 3,117.7 hectares. Shortly thereafter, it

acquired an additional application that had passed technical review by Secreminas. These six concession

contracts and applications comprised 3,169.9 hectares.

In October 2012, Red Eagle Mining acquired six concession contracts and eight applications from Bullet

and nine associated companies, totalling 30,097.0 hectares. Bullet has first rights to any tenements that

Red Eagle Mining should relinquish.

In July, 2013, Red Eagle Mining filed applications for two additional concessions.

Page 4.2




AMENDED NI 43-101 TECHNICAL REPORT In June, 2014 Red Eagle Mining acquired three concession contracts and three applications from

AngloGold Ashanti totalling 1,619.8 hectares.

One concession contract and two applications for which technical studies had been approved were

dropped by Red Eagle Mining in September, 2013, reducing the total tenement holding to its current area

of 33,224 hectares.

Red Eagle Mining holds 100% interest in all the tenements subject to completion of outstanding payments

of $1,790,000 to the underlying owners and subject to the royalty payments as described.

The San Ramon deposit and project site are located on concession contract – B7560005.

4.4.1 Purchase Agreement for Five Original Project Concessions

Five of the Santa Rosa Gold Project concessions are held by Red Eagle Mining under a sale and

purchase agreement. This agreement is to acquire 100% of five of the concession contracts: four

concession contracts numbered B7560005, B7171005, H5791005, and H5790005 registered in the name

of Red Eagle Mining, and concession contract application LDM-08061, for which the free area has been

declared that is awaiting transfer of title. The total purchase price was $8.2 million. $600,000 is

outstanding and is due upon title transfer of concession contract LDM-08061. One other concession

contract application (LKA-08004) was acquired under a separate purchase agreement described in

Section 4.4.2.

4.4.2 Purchase Agreement for LKA-08004

Red Eagle Mining can acquire LKA-08004 for a total cost of $180,000. The first payment of $40,000 has

been made. An additional payment of $70,000 is due ten days after the concession contract is registered

in the vendor’s name. The final payment of $70,000 is due ten days after the Secretary of Mines of

Antioquia confirms with a resolution the transfer from the vendor to Red Eagle Mining.

4.4.3 Purchase Agreement with Bullet

In October 2012, Red Eagle Mining entered into a Purchase Agreement with Bullet to purchase 100%

interest in one concession contract (at the time the Purchase Agreement was executed) and the mining

rights to 14 concession contract applications for which the technical study has been approved and free

areas have been declared that are awaiting transfer of title totalling about 30,097 hectares surrounding

Red Eagle Mining’s original concessions. In consideration for the property, Red Eagle Mining issued

905,000 common shares to Bullet and granted Bullet a royalty as discussed in Section 4.4.6 Bullet retains

a right of first refusal to acquire the parcels should Red Eagle Mining decide to let any of the mining rights

lapse.

Page 4.3




AMENDED NI 43-101 TECHNICAL REPORT 4.4.4 Purchase Agreement with AngloGold Ashanti Colombia S.A

Red Eagle Mining has entered into a Purchase Agreement with AngloGold Ashanti Colombia S.A.

pursuant to which Red Eagle Mining has agreed to acquire 100% of certain contiguous mineral

exploration rights totalling 1,619.8 hectares within Red Eagle Mining’s Santa Rosa Gold Project. In

consideration for the property, Red Eagle Mining has agreed to pay US$375,000 over two years from title

transfer (of which $25,000 has been paid to date) and grant a 2% net smelter return royalty on the

properties acquired. No other third party royalties exist on the property nor are they captured by existing

royalty arrangements.

This purchase covers a 1 km extension to the San Ramon deposit to the east, where there is potential for

additional mineralisation.

4.4.5 Agreement Regarding Artisanal Mining Operations

An artisanal gold mining operation in a 20 hectare area located within concession title B7560005 was shut

down in 2011 under an agreement between the vendors and Red Eagle Mining. This operation had

consisted of a single shaft into a narrow high-grade gold-bearing vein. Under this agreement, Red Eagle

Mining must pay the vendors $2,000,000 in instalments – US$1,300,000 has been paid, and US$700,000

is due by November 30, 2014. This agreement constitutes an additional part of the Purchase Agreement

for the five original project concessions described in Section 4.4.1.

4.4.6 Royalties

There is a government-imposed 4% royalty on gold and silver that is effectively 3.2% as described in

Section 0.

LMM purchased a 2% net smelter return royalty on the original six concessions of the Santa Rosa Gold

Project in October 2012 for $8,333,333. In December 2013 Red Eagle Mining sold a further 1% royalty to

LMM. Red Eagle Mining has the option to buy back 1% of the royalty for $8,333,333 for a period of two

years from the date of the first gold production. This royalty applies to the concession in which the

resource is located.

Bullet and AngloGold hold non overlapping 1.5% and 2% net smelter returns royalties respectively. These

royalties do not apply to the concession in which the resource is located.

4.5 Environmental Liabilities

There are no known environmental liabilities on the Santa Rosa Gold Project area at this time.

Red Eagle Mining’s activities to date have included exploration and inspections conducted by the

Environmental Authority on a regular basis. No major issues were raised and any minor issues were

remedied immediately to the satisfaction of the Environmental Authority.

Page 4.4




AMENDED NI 43-101 TECHNICAL REPORT The processes required for environmental related activities, as well as all commitments to date, are shown

in Table 4.2.

Table 4.2 Processes Required for Environmental Activities

Process Result

Submit a letter to the Antioquia Environmental

Authority (‘CORANTIOQUIA’) requesting Terms

of Reference for the EIA

Official Terms of reference for the area to be developed – Terms of

Reference were issued and accompanied by a Letter of

Confirmation from the Environmental Authority in July, 2013

Submit a letter to the Ministry of the Interior

(‘MININTERIOR’) requesting presence of ethnic

communities in the area

A certificate regarding ethnic communities presence/absence in the

area (black or indigenous communities) – Official Letter informing

the Company that there were no ethnic communities present in the

tenement areas issued in June, 2012

Submit a letter to the Institute of Rural

Development (‘INCODER’)

An official reply about ethnic community property in the area -

Official Letter informing the Company that there were no ethnic

communities present in the tenement areas issued in June, 2012

Submit a letter to the Institute of Anthropology

and History (‘ICANH’)

An official reply concerning any archaeological research

requirements in the area - Official Letter informing the Company

that the archaeological authorization N° 3425 is granted on May

22, 2013.

Official Letter informing the Company that the final report and the

Archaeological Management Plan have been approved in January,

2014.

Submit a Technical Study (PTO) describing the

technical aspects of the planned mining and

processing development and operations to the

Secretary of Mines (‘SECREMINAS’)

An official approval of the PTO, which initiates the environmental

approvals process – Submission of the official document in

November, 2013, with official approval of the PTO issued by

Secreminas in August, 2014.

Commence Environmental Base Line-EBL

studies

Preliminary baseline assessment – Work commenced under

contract with Universidad de Antioquia (UdeA), and Fundacion

Universitaria Catolica del Norte (FUCN) in early September 2012

and was completed in May 2013.

Consolidate Environmental Impact Assessment-

EIA

A final EIA document – Contract to complete the EIA was awarded

to Tetra Tech Colombia SAS in early December 2012, and was

completed and submitted to the Environmental Authority in

February, 2014.

Submit EIA to Environmental Authority

An official submission receipt on February 20, 2014 - An official

letter informing the Company that the Initial Act for the licensing

process is issued on May 6, 2014

Evaluate the final EIA and make resolution of

approval

An official Environmental License for the Company allowing it to

proceed, without any additional permits, to construction and

operations of the project.

After the Environmental License is issued and appropriate insurance against environmental impact has

been obtained, the concession is automatically authorized to start the construction phase and then is

converted to an exploitation concession; project construction can begin without any additional permitting

requirements.

Page 5.1





5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND

PHYSIOGRAPHY

5.1 Access to Property

The Santa Rosa Gold Project is accessible from the city of Medellín, which has an international airport, via

paved Highway 25 north-northeast through Copacabana and Don Matías for approximately 65km to a

turn-off located 12km south of Santa Rosa de Osos. From the turn-off to the east, it is approximately 8km

to the Red Eagle Mining camp on an unpaved road (Figure 5.1).

Figure 5.1 Access to the Santa Rosa Project

Santa Rosa Project Site

Page 5.2




AMENDED NI 43-101 TECHNICAL REPORT 5.2 Climate

The climate is mildly tropical with daytime temperatures around 24°C. Annual rainfall averages between

2,400 and 3,000 mm and falls mostly during rainy seasons from March to May and September to

December. There are no significant climatic restrictions on working. All activities, including exploration,

mining and Project construction can be conducted year round.

5.3 Physiography

Topographically, the Santa Rosa de Osos district consists of an irregular, dissected peneplain with gentle-

to steep-sided valleys and hills. Elevations range between about 2,300 m and 2,500 m. The region is

largely occupied by grass pasture and arable land with limited and often isolated areas of lush, low-growth

Andean forest, mostly located along drainages. Agriculture within the project area comprises cattle

farming in about 50% of the project area and tamarillo ("tree tomato") cultivation in approximately 10% of

the area, mostly in the southeast. There is also minor commercial forestry (pine). Tropically weathered

latosol profiles are ubiquitous and average 5m to 10m thick in undisturbed areas.

Because the topographic relief in the project area is moderate and the entire project is covered by

vegetation, surface exposures are limited to road cuts, cuts for drill-pad locations, and sluice areas, or

baticiones, denuded by historic hydraulic mining (see Figure 7.4).

The San Ramon deposit is located at an elevation of 2,453 m above sea level.

5.4 Infrastructure and Local Resources

The project area is located about 20 km by road southeast of the town of Santa Rosa de Osos (population

around 45,000 in the town and the same in the surrounding municipality), which is the nearest town for

supplies and labour. While most of the regional male population has had some involvement in

prospecting or small-scale mining, training would be necessary for a labour force to be used for mine

development. Security is provided in the district by a military base (Guadalupe) and troops (Batallón

Pedro Nel Ospina). Police patrol the Municipality of Santa Rosa de Osos.

Local resources and infrastructure in the Santa Rosa de Osos Municipality are excellent. A 44 kV power

line to the south and a 13.2 kV line to the north (both within 8 km) service the area. Water is abundant in

main rivers but is not potable.

5.5 Regional Seismicity

Most of the seismicity in the project area is considered ‘microseismicity’, as 79% of the recorded events

are shallow (< 30 km deep) (CRA, 2013). Figure 5.2 shows maximum event magnitudes for the

Department of Antioquia.

Page 5.3




AMENDED NI 43-101 TECHNICAL REPORT Expected ground accelerations according to the local building code (Norma Sismo-Resistente

Colombiana) reports a horizontal peak ground acceleration for the 475-year return period event of 0.15 g

(Figure 5.3). The project area is located in “Zone 3”, with an intermediate seismic classification. This value

is indicative of bedrock acceleration, which for the project area is covered by thick saprolite and residual

soils. A calculation of local seismicity and amplification potential has not been carried out to date.

After 1960 the national seismological network was established and about 30 monitoring stations are

working in the country.

Figure 5.2 Maximum expected magnitude (ML) for Antioquia (Project area inside red

square) From Sánchez (2009)

Page 5.4




AMENDED NI 43-101 TECHNICAL REPORT Figure 5.3 Peak ground acceleration map from Norma Sismo Resistente Colombiana (NSR-

10, 2010)

Red dot indicates project area

Page 6.1




AMENDED NI 43-101 TECHNICAL REPORT 6.0 HISTORY

6.1 Exploration History

Gold mining in the Department of Antioquia pre-dates, and continued during, Spanish colonial rule, mostly

exploiting alluvial deposits and oxidized veins (West, 1952). The Santa Rosa de Osos region was

discovered in 1541 by Captain Juan Francisco Vallejo, who named the area Valley of the Bears (de Osos)

because of the large number of the animals found in the region. Fifty years later, the gold potential of the

area was recognized, and hundreds of miners led by Captain Antonio de Espejo Serrano founded the

town of Santa Rosa de Osos in 1636. Gold production from the upland placers in the region reached its

peak in the 18th century. Mines within the Santa Rosa Gold Project area reportedly produced gold and

silver from bonanza-grade oxide ores during their heyday in the 1940s, but production declined as free

milling ores were exhausted. Small-scale gold mining continued intermittently and was recently in

operation at the Yaruma and Hilo Azul (Blue Vein) mines, which lie within the current resource area.

Exploration and small-scale gold mining by artisanal miners is ongoing about 10 km from the San Ramon

deposit within the Santa Rosa Gold Project area.

According to Jemielita (2011b), there was no known modern mineral exploration activity in the Santa Rosa

Gold Project area prior to the initial visit by Red Eagle Mining in July 2010. Since that date, Red Eagle

Mining has undertaken surveying and mapping of old mine workings, sampling and assaying in areas of

known mineralization, soil and saprolite grid geochemistry, airborne magnetic and radiometric geophysics,

and drilling of the San Ramon deposit shear zone. These activities are described in Section 9.0.

There are no known historical mineral resources or reserves estimates for the Santa Rosa Gold Project.

The first mineral resource estimate for the San Ramon deposit was reported in MDA’s Technical Report

for Red Eagle Mining, dated January 22, 2013 (Lindholm and Schlitt, 2013a). A subsequent Technical

Report (Lindholm and Schlitt, 2013b) described the updated resource estimate, which included a

combination of open-pit mineable material at a cut-off grade of 0.3g Au/t and underground mineable

material at a cut-off of 1.2 g Au/t. In 2014, a Preliminary Economic Assessment (Dyer et al., 2014) based

on this resource estimate was compiled. The September 2013 estimate is also used as the basis for the

Feasibility Study but the resource is now reported at a cut-off grade for only underground mining of 1.2 g

Au/t.

Based upon the existence of the extensive abandoned mine workings, past gold production appears to

have been significant, but actual historic gold production is unknown.

Page 7.1




AMENDED NI 43-101 TECHNICAL REPORT 7.0 GEOLOGIC SETTING AND MINERALIZATION

7.1 Geologic Setting

7.1.1 Regional Geology

The northwest margin of South America, comprising Colombia and adjacent Ecuador and western

Venezuela, is a complex elongated mosaic of Paleozoic and younger, autochthonous and allochthonous

terranes that were accreted to the South American continent (Guyana Shield) and subjected to

transpressional tectonics and subduction-related magmatism along a 2,000km-long segment of the Pacific

Rim (Aspden and Litherland, 1992, and Cediel at al., 2003, as cited by Jemielita, 2011a, 2011b).

Regional-scale structural geology is characterized by anastamosing, sub-parallel, dominantly north-

northeast-striking, major faults and shear zones, some of which are interpreted to be terrane boundary

sutures. Subduction-related magmatic arcs were superimposed on these terranes during the Jurassic,

Cretaceous, Eocene to late Miocene, and Late Miocene to Recent and produced porphyry copper

mineralization related to each period of arc magmatism (Figure 7.1). These arcs are characterized by

plutonic batholiths, sub-volcanic intrusions, and associated volcanic rocks.

The Andean Cordillera in Colombia is composed of three distinct mountain chains: the Western Cordillera

(Occidental), Central Cordillera (Central), and eastern Cordillera (Oriental) that are separated by broad

inter-Andean valleys. The Santa Rosa Gold Project is located in the Central Cordillera within the

Cajamarca-Valdivia terrane. This terrane dominates the geology of the northern portion of the Central

Cordillera and is a composite lithotectonic unit that includes a metamorphic basement complex and the

Antioquia Batholith (Cediel and Caceres, 2000, and Cediel et al., 2003, as cited by Jemielita, 2011b).

The Cajamarca-Valdivia basement terrane is made up of early Paleozoic metamorphic rocks, mostly lower

greenschist- to lower amphibolite-grade metasedimentary units and oceanic ophiolitic volcanic and

intrusive rocks. These comprise a para-autochthonous prism that was accreted to the continental margin

along the Palestina fault system in the Ordovician-Silurian and subsequently underwent regional

metamorphism (Cediel et al., 2003, as cited by Jemielita, 2011b). The Cajamarca–Valdivia terrane is

bounded on the west by the Romeral fault, a north-striking, dextral transcurrent suture that defines the

eastern limit of allochthonous oceanic terranes accreted to the northern Andean margin during the late

Mesozoic and Cenozoic.

During the Mesozoic, the metamorphic rocks of the Cajamarca-Valdivia terrane were onlapped by

volcano-sedimentary lithologies. Reactivation of the Palestina fault system and initiation of the Romeral

fault system occurred from the Aptian-Albian, together with a period of uplift and erosion of the Paleozoic

basement and Mesozoic volcano-sedimentary sequences (Cediel and Caceres, 2000, as cited by

Jemielita, 2011b). The basement and Mesozoic lithologies were intruded by the Antioquia Batholith during

the Cretaceous under a regional tectonic regime of dextral transpression.

Page 7.2




AMENDED NI 43-101 TECHNICAL REPORT Figure 7.1 Volcano-plutonic Magmatic Arcs and Major Intrusions of the Northern Andes

(From Jemielita, 2011b, citing PRODEMINCA, 2000)

Page 7.3




AMENDED NI 43-101 TECHNICAL REPORT The Antioquia Batholith is shown by a red arrow. It is located in northern Colombia and belongs to the

early- to mid-Cretaceous magmatic arc.

The Antioquia Batholith occupies more than 7,000 square Kilometers and is a subduction-related, multi-

phase, calc-alkaline, I-type plutonic-intrusive complex composed mostly of quartz-diorite and granodiorite

(Feininger and Botero, 1982, as cited by Jemielita, 2011b). Rock textures are holo-crystalline, medium to

coarse grained, and phaneritic. Mineralogy is dominated by plagioclase and quartz +\- potassium feldspar

and up to 20% mafic minerals ranging from biotite to hornblende-dominant. Trace minerals include

magnetite and titanite. Local- to regional-scale metasomatism is evident as chlorite after biotite and

epidote after hornblende. The batholith is cut in places by centimeter- to meter-scale dikes of porphyritic

diorite, quartz diorite, aplite, and granite pegmatite. Small plugs of bi-pyramidal quartz- and biotite-phyric

hypabyssal granodiorite porphyry have been identified at various localities. Radiometric age dates are

scattered from around 90 to 60 Ma but cluster around 70 Ma.

The Antioquia Batholith was intruded during dextral-oblique accretion and transpression of the Andes

during the late Cretaceous–Eocene. Major lineaments, especially in the east, strike west-northwest to

northwest and may be related to dextral reactivation of the Palestina fault system in the east and dextral

transpression along the Romeral fault system in the west.

Figure 7.2 shows the regional geology in the vicinity of the Santa Rosa Gold Project area.

Page 7.4




AMENDED NI 43-101 TECHNICAL REPORT Figure 7.2 Santa Rosa Gold Project Regional Geology

Page 7.5




AMENDED NI 43-101 TECHNICAL REPORT 7.1.2 Local Geology

The rocks in the vicinity of the Santa Rosa Gold Project are dominated by hornblende-biotite diorite and

quartz diorite typical of the Antioquia Batholith (Figure 7.3). Metamorphic rocks including amphibolites

and metasedimentary rocks occur as isolated roof pendants, primarily in the western half of the deposit

area. Pleistocene-Holocene volcanic ash cover is extensive. River valleys contain unconsolidated

alluvium.

Figure 7.3 Local Geology in the Vicinity of the Santa Rosa Gold Project

Page 7.6




AMENDED NI 43-101 TECHNICAL REPORT 7.1.3 Property Geology

The geology within the Santa Rosa Gold Project is characterized by relatively monotonous, coarse

grained, grey-colored, granular hornblende-biotite diorite and quartz diorite of the Antioquia Batholith

(Figure 7.3). Metamorphic rocks, including amphibolites (Valdivia & others Gr. on Figure 7.3) and

metasedimentary rocks (Valdivia Gr. on Figure 7.3), occur as isolated roof pendants, and dikes of micro-

diorite, dacite, felsite, aplite, and pegmatite are common. Red-brown saprolite is widespread and often

deep (up to 50 m). Deep weathering associated with saprolite formation has intensely altered the granites

to clay. Oxidized saprock continues below the saprolite and consists of more competent, albeit somewhat

decomposed, granitic rocks. Soils are generally about 50 cm thick and rarely up to 2 m. Schistose fault-

zone mylonite was observed at several locations in outcrop and adits.

7.2 Mineralization

Hypogene gold mineralization within the Santa Rosa Gold Project is generally associated with the shear

zones developed in homogeneous diorite country rock, with higher grades occurring in the associated

sulfide-mineralized quartz veins or as steep high-grade quartz-sulfide veins. There are also related

saprolitic gold deposits (Figure 7.4) and colluvial gold deposits, both of which have been mined by

artisanal miners underground and in hydraulically mined areas known locally as “baticiones.” The shear

zones and veins are best exposed in adits and baticiones.

Figure 7.4 Oxidized Saprolite Exposed in a “Batición”

Page 7.7




AMENDED NI 43-101 TECHNICAL REPORT Colluvial and saprolitic gold workings and adits are concentrated in two principal areas: one cutting

concession B7560005 from west to east across the center of the concession; and one to the west and

northwest in concession H5790005 and the northern portions of concessions B7171005 and H5791005.

These two areas contain large abandoned colluvial workings and both abandoned and some active

tunnels made by artisanal miners. At the time of the initial resource estimate, a total of 258 adits had been

surveyed on the Santa Rosa Gold Project area, totalling 5,886m; many more underground workings have

been discovered since then, bringing the total number of known adits to 1,150. These adits are nearly all

developed in oxidized saprolite. The northwestern area has more adits than the southeastern one. Gold

mineralization is processed in local water-powered California stamp mills (Figure 7.5) with mercury

amalgamation or, in the southeastern area, a jaw crusher and mercury amalgamation mills. Figure 7.6

shows the locations of adits, baticiones, and drill holes on the Santa Rosa Gold Project.

Figure 7.5 Water-Powered Stamp Mill

Page 7.8




AMENDED NI 43-101 TECHNICAL REPORT Figure 7.6 Adits and Baticiones on the Santa Rosa Gold Project

Exploration by Red Eagle Mining within concession B7560005 has identified a mineralized shear zone

(San Ramon deposit shear zone) containing sulfide-bearing quartz veins, sheeted veins, and

anastomosing vein networks. The shear zone is characterized by both ductile deformation in the form of

mylonite development in the most intense zones and brittle deformation in the form of breccia and gouge

zones. The structural zone strikes roughly east-west with a strike length of approximately 2.0 km,

predominantly dips about 70° to the north, and has an average width that ranges from about 8m in the

western half to 21 m in the eastern half. The dip of the shear zone steepens to near vertical close to the

surface and shallows to 50° to 60° at depth.

East-west-trending structures, including the San Ramon deposit shear zone, appear to be related to

northwest-trending regional structures. Sinistral movement along these structures may have created east-

west dilation zones, up to 60 m wide in the case of the San Ramon deposit shear zone, into which quartz

and quartz-carbonate veins and veinlets were emplaced. Most of the quartz veins and contained sulfides

have been brecciated by post-mineralization deformation, which suggests that the mineralization is syn-

deformational.

Page 7.9




AMENDED NI 43-101 TECHNICAL REPORT Numerous micro-diorite dikes and a few dacite dikes are generally intruded along shear-zone structures.

They are strongly fractured and brecciated and are commonly mineralized, which suggests the dikes are

pre-mineral and pre- to syn-deformational. Numerous dacite dikes also cross-cut the shear zone

structures. These may have been intruded along a fabric in the diorite country rock that is parallel to the

included schist bodies. These dacite dikes appear to be only weakly deformed by shearing and may be

predominantly post-mineral and post-deformational. However, assays of dike material commonly show

the presence of mineralization, so some of these dikes may have been emplaced during, but late in, the

mineralizing and deformational history of the deposit.

In the San Ramon deposit shear zone, mineralization occurs in fractured, brecciated, and ductily deformed

rock. The lowest levels of mineralization (~0.035g Au/t to ~0.1g Au/t) consist of very weak to moderate

ductile shearing that locally contains quartz veinlets and/or pyrite. At slightly higher grades (~0.1 to ~0.6 g

Au/t), mineralization is characterized by moderate to strong ductile deformation that contains scattered

quartz veinlets and sulfides, although the overall quantity of veinlets and sulfides is low. Some sericite,

weak to moderate brittle overprinting, and quartz veins greater than 12 mm in thickness that are mostly, if

not completely, barren of sulfides, may also be present. The boundary between the two grade ranges is

likely gradational, and there are numerous instances where quartz veins or pyrite are not apparent in

either. Shear zone intercepts containing these low grades of mineralization may be up to 80 m wide; the

maximum true width of the mineralized shear zone is about 60 m.

Higher-grade mineralization, >0.6 g Au/t, within the San Ramon deposit shear zone is characterized by

strong shearing, more intense sulfide mineralization, and quartz veins. At relatively lower grades within

this higher-grade mineralization (~0.6 to 5.0 g Au/t), strong ductile shearing with abundant sulfide minerals

and sericite is almost ubiquitous; quartz veins may or may not be present. Brittle overprinting is moderate

to strong and may be genetically related to the high-grade mineralization. At grades in excess of ~5.0 g

Au/t, massive and coarse-grained pyrite, pyrite stringers, medium-grained sphalerite, fine-grained galena,

and traces of chalcopyrite are present in quartz veins and quartz vein fragments, although some massive

pyrite and pyrite stringers occur independently of quartz veins. In the highest-grade intercepts (>50 g

Au/t), relatively thick (≥2 cm) massive pyrite veins are intermixed with quartz veins that contain coarse-

grained and massive pyrite, coarse-grained sphalerite, fine-grained galena, and traces of chalcopyrite.

The predominant distinguishing characteristic between low and higher-grade mineralization and for the

strength of gold mineralization in general, is the quantity of sulfide minerals. Both domains may appear

exactly alike in terms of strong ductile deformation and the presence of quartz veins, but the sulfide-

mineral content is the factor that distinguishes higher-grade mineralization in most cases.

The zones of strongest shearing have some demonstrable continuity between drill holes on individual

sections and between sections. The San Ramon deposit shear zone contains a pair of higher-grade

zones between 1 and 2 m in width that are present over a significant length of the deposit. Mineralization

at more extreme grades associated with coarse-grained sulfides in quartz veins lacks this continuity within

the shear zone. Red Eagle Mining staff has observed that one high-grade vein in the Hilo Azul workings is

approximately 30m in total length along strike. The vein appears to pinch out, then reappears after some

distance. Nevertheless, mineralization at ~2 g Au/t, the approximate economic cut-off grade, is

continuous over distances of hundreds of meters.

Page 7.10




AMENDED NI 43-101 TECHNICAL REPORT The pair of mineralized zones described above are parallel to the walls of the shear zone. The stronger

and better mineralized zone lies roughly along the footwall of the shear zone and has a fairly sharp

footwall contact but a less regular hanging-wall contact. The second mineralized zone, which lies toward

the hanging wall of the shear zone, is less persistent and less frequently mineralized than the footwall

mineralized zone. The fabric in the two mineralized zones is parallel to the trend of the San Ramon

deposit shear zone. The orientation and degree of mineralization of fabric in other parts of the shear zone

can vary. Locally, the shear zone may range from shallow-dipping to sub-vertical, may be cross-cutting to

the overall east-west shear zone fabric, and is variably mineralized.

Vein-quartz textures are mostly massive to ribbon-textured and, in places, medium- to coarse-grained

crystalline and often containing late calcite-infilled tensional features. Sulfides range from 1% to 5% but

can reach 10%, and there appears to be a direct correlation between sulfide and gold grade. Sulfides are

dominated by fine- to coarse-grained pyrite with subordinate sphalerite and galena and traces of

chalcopyrite and pyrrhotite. Oxides include hematite, goethite, and limonite. Black oxides are a

weathering effect caused by the migration of manganese away from organic-rich soil into fractures. There

is no preferential association of manganese oxides with gold mineralization. Gold and minor amounts of

silver occur as inclusions in sulfides, mostly in sphalerite and less commonly pyrite and galena. In

addition to a mixed oxide / sulfide transition zone, unoxidized sulfide minerals are commonly present

within the oxidized and saprolitized rocks within the shear zone.

There are calcite veins of various ages within the San Ramon deposit. Shear-zone controlled weak

propylitic alteration of regional extent pre-dates mineralization and consists of calcite-chlorite-prehnite

veins with over 80 % calcite. These veins are low in sulfides and do not contain gold. Calcite and calcite

veining are common in the main stage of the shear-zone mineralization; the veins are often brecciated,

suggesting they were an earlier part of the mineralizing event. This calcite is characteristically iron-rich

and weathers to a brown color. It is associated with sulfides, euhedral quartz, and the best gold

mineralization. There are also low-temperature veins of pure calcite that are late and cut all other

features. These calcite veins are regionally distributed. They contain only low levels of iron but no

sulfides, gold, or silicate minerals.

In the San Ramon deposit shear zone, the overall silver content is fairly low, but the silver content of the

gold grains is unusually high, in some cases >75% silver. Drill core assay and metallurgical testwork data

indicates a gold to silver ratio of approximately 1 to 1.85. On the basis of mineralogical studies, most of

the silver is alloyed with the gold rather than incorporated in sulfides. There can be significant quantities

of zinc, which is accompanied by minor quantities of lead. The presence of copper is insignificant. Silver

has not been geologically modelled.

Page 8.1




AMENDED NI 43-101 TECHNICAL REPORT 8.0 DEPOSIT TYPES

Gold mineralization in the Santa Rosa Gold Project has characteristics in common with mesothermal or

orogenic and intrusion-hosted gold veins. Mineralization of these types consists commonly of quartz and

quartz-carbonate veins located in moderately to steeply dipping, brittle-ductile shear zones and locally in

shallow-dipping extension fractures. Veins commonly extend along strike and down dip over very

significant distances and occur alone or, typically, in complex vein networks and shear zones. Vein

minerals are mostly quartz and carbonates with minor native gold, pyrite, and base metals sulfides. Veins

are usually massive or ribbon-textured, but vein breccias and drusy, crystalline quartz can also occur.

Wall-rock alteration is zoned and consists of carbonate (often ankerite), sericite, and pyrite.

Gold at the Santa Rosa Gold Project was historically extracted from saprolite (and alluvial gravels).

Potential exists for discovery of new deposits in the area and additional mineralization down-dip of the

known San Ramon deposit mineralization.

Page 9.1




AMENDED NI 43-101 TECHNICAL REPORT 9.0 EXPLORATION

9.1 Geologic Mapping

Red Eagle Mining has undertaken an extensive program of mapping and sampling old colluvial gold

workings, adits, and quartz veins within the Santa Rosa Gold Project area. Mapping and sampling of

several adits located in the eastern and western portions of the San Ramon deposit shear zone have

identified a significant number of zones of gold mineralization.

Red Eagle Mining has attempted to define the course of mineralized veins in the area of the San Ramon

shear zone by mapping the position of all identifiable adits, shafts, and crown holes (see Figure 7.6). In

the block of concessions (H5790005, B7171005, H5791005) west of the San Ramon deposit Red Eagle

Mining has been conducting surface mapping of geology, and the distribution of adits as well as surveys of

the baticiones in an attempt to better understand the distribution of veins and high grade shoots.

9.2 Geochemistry

Quartz veins, shear zones, and granodiorite diorite schist wall rocks exposed in artisanal adits and mines

have been extensively grab, channel, and panel sampled. A total of 4,637 samples have been taken.

Channel or panel sampling was used to test the quartz veins, while a program of continuous channel

sampling in adits that cut across the broader vein and shear structures in the oxide zone was initiated in

January 2011. A number of key adits were sampled in which the San Ramon deposit shear zone and

some of the hanging-wall structures are exposed.

Further sampling was carried out where sulfides were found to be present. Hilo Yaruma had been

previously sampled underground at a depth of 65 m and returned values of 95.8 g Au/t in a vein

approximately 30 cm wide (the shaft has since collapsed). The Hilo Azul shaft (quartz veins within the

shear zone) was sampled at a depth of 45 m and returned 41.6 g Au/t in the principal vein intersected; a

second vein approximately 20 m into the footwall gave values to 18.8 g Au/t. The shear zone away from

the quartz vein gave values up to 7.4 g Au/t. The reopened El Gato adit, which lies between the San

Francisco adit and the Hilo Azul shaft, also gave access to old stopes in partially oxidized mineralization

both in quartz veins and shear zone. Initial samples returned up to 19.6 g Au/t in the sulfide-rich mylonitic

shear zone material and 10.3 g Au/t in quartz vein samples.

A 200 m by 200 m ground auger saprolite geochemical sampling program was completed over the eastern

part of concession B7560005 with the objective of comparing the results with a mobile metal ion (“MMI”)

sampling over the same area. It was concluded that MMI was superior as an exploration tool, and

significantly easier to implement.

A soil sampling program was conducted over the Santa Rosa Gold Project, and samples were analyzed

for MMI. The grid covering the concessions was on 50 m spacing north-south and 200 m spacing east-

west, with some areas expanded to 400 m east-west spacing to expedite the initial reconnaissance

survey. Anomalies were identified corresponding to the San Ramon deposit shear zone as well as other

areas that have not been fully explored to date (Figure 9.1).

Page 9.2




AMENDED NI 43-101 TECHNICAL REPORT Figure 9.1 MMI Geochemistry of the Santa Rosa Concessions Showing Gold Anomalies

Red Eagle Mining has undertaken exploration of the concessions beyond the San Ramon deposit. This

work included stream-sediment, rock-chip and trench, adit, and MMI and conventional soil geochemical

sampling. This work is being conducted with the objective of identifying potential targets for mineralization

that may become future satellite feed to the San Ramon Gold Project. A total of 264 stream-sediment

samples and 440 soil samples have been taken to date.

MDA has not analyzed the sampling methods, quality, and representivity of surface sampling on the Santa

Rosa Gold Project, and those are not used in the mineral resource estimate.

Page 9.3




AMENDED NI 43-101 TECHNICAL REPORT 9.3 Geophysics

A helicopter-borne, high resolution, magnetic and radiometric survey was completed over the western part

of the original concessions and applications by MPX Geophysics on November 18 and 19, 2010. A total

of 451.9 line-Kilometers of data were acquired over the project area (total area 19.6 km2 at the time). The

survey was flown at a nominal mean terrain clearance of 70 m along north-south-oriented flight lines

spaced at 50 m with tie lines spaced at 500 m.

Red Eagle Mining retained consultants Paterson, Grant & Watson to interpret the aeromagnetic survey

(Ugalde and Misener, 2011). Results were received in January 2011 and are shown in Figure 9.2 and

Figure 9.3.

A similar aerial survey was subsequently flown over the eastern part of the project area, and is also shown

in Figure 9.2.

Figure 9.2 First Vertical Derivative of Total Magnetic Intensity

Note: (i) The area shown in red is the surface expression of the mineralized shear zone but is not the outline of MDA’s mineral resource estimate. The resource outline is shown on Figure 10.1.

Page 9.4




AMENDED NI 43-101 TECHNICAL REPORT Figure 9.3 Total Potassium Radiometric Plot

Note: (i) The area shown in red is the surface expression of the mineralized shear zone but is not the outline of MDA’s

mineral resource estimate. The resource outline is shown on Figure 10.1.

The goals of the aeromagnetic interpretation were:

Definition and mapping of important structures in the area that might have a role in gold

mineralization;

Lithological mapping based on the airborne magnetic data;

Overall refinement of the known geology of the area; and

Definition of targets and areas of interest for an upcoming ground IP survey.

Page 9.5




AMENDED NI 43-101 TECHNICAL REPORT Red Eagle Mining notes that the magnetic data (Figure 9.2) indicated major structures that influenced the

geology, particularly the San Ramon deposit shear zone. The radiometrics, especially potassic alteration

(Figure 9.3), correlated to ”baticiones” and the extensive artisanal mining areas that were being

uncovered.

An induced-polarization (“IP”) survey was conducted by Geofisica TMC SA de CV of Mazatlan, Mexico

over a 50 m (north-south) by 200 m (east-west) grid in the San Ramon deposit area (most of concession

B7560005) in June and July 2011 (Simard, 2011). The survey consisted of 13 north-south lines of 2.5 to

2.8 km in length, totalling 32.6 line-km. The pole-dipole array was used. Red Eagle Mining felt that the

results were inconclusive and decided not to continue IP surveying following unsuccessful drilling into

interpreted IP targets.

9.4 Topography

As described in more detail in Section 12.1.1, initial topography for the project was provided by

Terranalisis, Ltda., based in Santiago, Chile, and MPX Geophysics Colombia SAS (“MPX”) based in

Medellín, Colombia. Subsequently, a ground survey for topography was commissioned through the

project surveyor, MSc. Ricardo Lozano Botache with Estudio T-Rural Consultores. Finally, MPX

performed a high resolution Lidar topographic survey that covers concession B7560005, on which the San

Ramon deposit is located.

Page 10.1




AMENDED NI 43-101 TECHNICAL REPORT 10.0 DRILLING

10.1 Summary

Red Eagle Mining has conducted all of the known drilling at the San Ramon deposit. From September

2011 to May 2012, Red Eagle Mining drilled 42 holes, including re-drills. From May 2012 to July 2012, an

additional 27 holes were drilled, consisting entirely of in-fill drilling in the sulfide mineralization to achieve

100 m centers within the San Ramon deposit shear zone. From July 2012 through October 2012, 74

holes were completed, designed to infill the oxide mineralization to 50 m centers. From October 2012

through May 2013, 95 holes were completed to test continuity of the mineralization in the sulfide zones

and extend the known deposit down dip. Holes drilled at the San Ramon deposit by Red Eagle Mining as

of the effective date of the resource estimate are SR-001 through SR-233 plus SR-028A, SR-032A, SR-

045A, SR-050A and SR-146B. These 238 holes total 45,609 m.

Further drilling for geotechnical and hydrogeological purposes was undertaken between May and July

2014 to investigate the ground conditions that will be encountered during plant construction and decline

development. Two of the 11 holes were drilled into the San Ramon deposit shear zone (SR-234 and 235).

Figure 10.1 shows drill-hole locations.

In general, holes were drilled sub-perpendicular to the strike of the generally east-trending shear structure

at various dips. Drill holes for the initial phases (SR-001 to SR-065) were collared some distance north of

the shear zone and intercepted the zone at depth (up to 300 m vertically). Subsequent holes (SR-066 to

SR-139) were collared closer to the shear and intercepted the shear mineralization closer to the surface.

New drilling for the resource update (SR-140 to SR-233) was intended to infill sulfide zones to 50 m

spacing and to extend known mineralization down-dip; the deepest shear zone intercept is approximately

550 m.

10.2 Drilling by Red Eagle Mining

This subsection describes drilling by Red Eagle Mining that comprises the database for the resource

estimate described in this report, and which formed the basis of the MDA PEA of October, 2013.

All drilling through May 2013 by Red Eagle Mining was core drilling. The drilling through July 2012 was

conducted by Cabo Drilling (Colombia) Corp. (“Cabo”), based in Bucaramanga, Santander Department,

Colombia, using two skid-mounted Boyles drill rigs, a BBS-37 and a BBS-56. Through July 2012, holes

were drilled with HQ core, reducing to NQ core as required by drilling conditions.

Drilling from July 2012 to October 2012 was conducted by Energold Drilling Corp. using a smaller skid-

mounted EGD II (Hydracore 600 Series II) for drilling in oxidized areas. This drilling consisted entirely of

HQ core.

For the drilling from October 2012 through May 2013, drilling of HQ-diameter core was performed by Cabo

using four rigs, two skid-mounted Boyles (BBS-37 and BBS-56) and two man-portable rigs (Duralite 1000

and Hydracore 2000). Skid-mounted rigs were moved between sites with a Caterpillar D6 bulldozer, while

the man-portable rigs were moved by a crew of up to 12 labourers and/or using winches.

Page 10.2




AMENDED NI 43-101 TECHNICAL REPORT The core was laid out, reassembled to a best fit, and cleaned prior to collection of core recovery and rock

quality determinations (RQD) and logging. Geologists logged lithology, alteration, mineralization, and

structural data. The core was photographed using a digital camera mounted on a tripod; MDA observed

that the photographs were of excellent quality.

10.3 Drill-Hole Collar Surveys

Drill holes were staked-out by GPS and oriented by setting up a string-line laid out with a Brunton

compass. Collars were surveyed after completion of drilling by Ricardo Lozano Botache, a licensed

professional surveyor with Estudio T-Rural Consultores based in Bucaramanga. At the end of MDA’s

2012 site visit, the surveyor was setting up to resurvey all drill-hole collars, primarily to check suspected

discrepancies between drill collars and topography. All collar surveys through May 2013, as well as all

project data, are now normalized to the currently-used coordinate system (GCS_Bogotá, Datum:

D_Bogotá).

10.4 Down-Hole Surveys

Holes drilled through July 2012 were initially surveyed down-the-hole by Cabo’s drillers, and subsequently

by Red Eagle Mining, using a Reflex EZ-Trac® tool. Readings were taken every 100 m, starting at 100 m.

The planned azimuths (azimuths of string lines placed on pads to set up the drill rig) and dips were used

for the orientation at the surface; actual azimuths and dips were not measured once the drill rig was set up

and drilling. MDA noted some large differences between the surface azimuth and dip and the first

REFLEX reading at 100 m in previous phases of drilling. The most significant of these differences were

checked by Red Eagle Mining, and modifications to the collar and down-hole survey were made as

appropriate.

No down-hole surveys were conducted for drilling from July to October 2012, primarily because the oxide

holes are short and Red Eagle Mining determined that the deviations at shallow depths in previous drilling

phases were not significant.

For the drilling from October 2012 through May 2013, holes over 100 m deep (66 of 95 holes) were

surveyed at 100 m intervals using a Reflex EZ-Shot®/manual single-shot instrument with an external

reader, operated by Cabo’s drillers. First readings were taken between 6 m and 100 m depths. The

planned azimuths and dips were used for the orientation at the surface.

Page 10.3




AMENDED NI 43-101 TECHNICAL REPORT Figure 10.1 Drill-Hole Location Map in the San Ramon Deposit Area

Page 11.1




AMENDED NI 43-101 TECHNICAL REPORT 11.0 SAMPLE PREPARATION, ANALYSIS, AND SECURITY

This section describes sampling, sample preparation, analysis, and security for drilling by Red Eagle

Mining that comprises the database for the resource estimate described in this report. The information in

this section includes observations by MDA, information provided by Red Eagle Mining, and descriptions of

early work from Jemielita (2011b).

11.1 Sampling Procedures

11.1.1 Surface and Adit Sampling

Red Eagle Mining geologists performed and/ or supervised composite rock-chip, channel, and panel

sampling from surface outcrops and adit walls from all known accessible locations throughout the San

Ramon deposit area. Channel sampling was used on zones of dispersed mineralized quartz veinlets

using a maximum of 2 m sample lengths where there appeared to be little mineralization. Areas of

possible mineralization were sampled across that specific zone. The channels were cut with a pick with a

width of 5 to 10 cm, with best efforts to maintain consistent and unbiased samples; generally,

approximately 3 kg samples were collected for each sample interval.

MMI soil sampling was conducted by digging shallow pits 10 cm to 15 cm deep into the “B” soil horizon on

a 50 m (north-south) by 200 m (east-west) grid. Saprolite was sampled using an auger (to 5.5 m

maximum depth) on a 200 m by 200 m grid and analysed using conventional geochemistry.

11.1.2 Drill Sampling

Geologists determined sample intervals using the observed geology as a guide. Maximum sample length

for drilling through July 2012 was 2 m; from July through October 2012, it was 1 m. For drilling since

October 2012, mineralized material was sampled at 1m intervals, except in narrow zones where the

sample length could be reduced to a minimum of about 0.6 m. If less than 0.4 m of a geological zone

would remain when a 1 m sample was taken, the sample length was adjusted to match the geology and

provide maximum sample lengths. Outside of a mineralized zone, samples up to 2 m in length were

taken.

For the earliest drilling, the entire hole was sampled. As the geologists became familiar with the various

rock and alteration types and the grades associated with them, sampling was only undertaken in core in

the shear zones or zones that appeared to be mineralized, with two or more buffer samples above and

below these zones. This selective sampling covered the primary mineralized zones, as long as they were

recognized, but did not allow for a holistic assessment of a deposit that includes country rock.

For drilling through July 2012, the core was halved with a diamond saw; for infill drilling in oxide material

from July to October 2012, the entire core was sampled. For drilling from SR-140 and higher (October

2012 to May 2013), core was again cut in half with a diamond saw for sampling. The sample for assay

was placed into plastic bags containing the sample ID tag and sealed immediately. The bags were zip-

tied closed. Individual samples were placed into rice bags labelled with the contained sample number

range and zip-tied closed.

Page 11.2




AMENDED NI 43-101 TECHNICAL REPORT Sampling was carried out or supervised by Red Eagle Mining geologists.

11.2 Sample Preparation and Analysis


Rock-chip and channel samples originally were sent to ALS Minerals (“ALS;” formerly called ALS

Chemex) for sample preparation in Bogotá and analysis in Lima. Subsequently Red Eagle Mining

changed laboratories and sent samples for analysis to SGS del Perú, S.A.C. (“SGS”) for precious-metal

and multi-element assays and then later sent samples to Acme Analytical Laboratories S.A. (“Acme”).

Some channel samples from adits were re-sampled and then assayed using metallic screen analyses.

Saprolite auger samples were delivered to SGS Colombia S.A. (“SGS”) in Medellín or Bogotá for

preparation, and then 30 g pulps were forwarded to SGS in Lima for analysis. Analyses were performed

by SGS using gold fire assay plus 52 elements by ICP. Duplicates were analyzed by ALS using gold fire

assay plus 35 elements by ICP. Some samples were screened for coarse gold.

MMI soil samples were delivered to SGS in Medellín and forwarded directly to Lima for sample

preparation followed by analysis in Australia.

11.2.2 Drill Samples

Drill samples were shipped to Acme in Medellín for sample preparation. The samples were dried,

crushed, and pulverized to 200 mesh.

Following sample preparation, pulps were forwarded to Acme in Santiago, Chile for analysis. If there was

a backlog at the Santiago laboratory, the pulps were sent to Acme’s laboratory in Vancouver, Canada, for

analysis. For drilling through October 2012, the analytical lab performed gold fire assays on 30 g samples

with an AA finish plus a 36-element ICP scan. All assays returning values greater than 10 g Au/t (>0.2 g

Au/ton in early drilling) were fire assayed again but with a gravimetric finish. For the drilling from October

2012 through May 2013, 50 g charges were used for both fire assay-AA and fire assay-gravimetric

analyses.

Field and preparation duplicates at the time of MDA’s 2012 site visit were analyzed in sequence with the

original at Acme.

For SR-001 to SR-028A, checks were made on samples suspected of coarse gold and were subjected to

metallic screen analyses; no metallic screen assays have been performed on drilling samples since

completion of those holes.

Page 11.3




AMENDED NI 43-101 TECHNICAL REPORT 11.3 Sample Security


Chain of custody is maintained for all samples. Rock samples (average 2 kg) were placed into a plastic

bag marked with the sample number on both sides and on a piece of flagging tape that was placed inside

the bag; the bag was sealed immediately. The sample and sample location were then photographed.

Bagged samples were put into larger sacks in the field, and when filled, these sacks were sealed in the

field. Sample sacks were then securely transported to the sample storage location. Every Friday or

Saturday, the sample sacks were transported directly to the sample preparation laboratory in Medellín,

together with dispatch documents. Logistics were supervised and monitored by Red Eagle Mining’s

security staff and support. The company then confirmed receipt of the samples with the laboratory.

11.3.2 Drilling Samples

Core was checked and collected by Red Eagle Mining geologists from the drill rig at least once per shift

and transported to the core-processing facility at the Red Eagle Mining camp. The wooden core boxes

were covered and kept closed with inner-tube strips. The core-logging area and storage buildings consist

of metal-framed structures with corrugated plastic roofs, and brick walls and/or plastic-covered wire

fencing on all sides. The building is locked when no one is on site. On a weekly basis, Red Eagle Mining

personnel transported the samples via company truck to the sample preparation laboratory in Medellín,

where custody was transferred to Acme.

The remaining core was stored in locked buildings at the camp, as are coarse rejects and pulps returned

from the laboratory.

11.4 Quality Assurance / Quality Control


Red Eagle Mining purchased certified standards and certified blanks and uses field duplicates for QA/QC

on their analytical work for rock samples. A duplicate, standard, or blank was included with every

10 samples sent for analysis.

For MMI soil geochemistry samples, a duplicate was included with every 30 samples, but no blanks or

standards were used.

11.4.2 Drill Sampling

Approximately 10% of the samples shipped to Acme were either a blank, a standard comprising one of

four standard samples, or a duplicate. Pulp blanks and standards were submitted with the original

samples and were not “blind” to the laboratory.

Page 11.4




AMENDED NI 43-101 TECHNICAL REPORT Standards for gold were inserted into the sample stream every 40 samples. For all drilling, a total of

11 certified standards were obtained and submitted with original drill samples. Of the 11 standards, nine

were used for all but four of the submitted pulps. No sample pulps were re-assayed as a result of

standard failures for all drilling prior to the initial resource estimate (SR-001 to SR-139). For the drilling

from October 2012 to May 2013, when standards failures occurred, Red Eagle Mining assembled and

renumbered all pulps in their possession in the batches associated with the failures, and resubmitted

samples to the laboratory. These were accompanied by two to four standards; two to four pulp blanks,

and one or two pulp duplicates for each batch. In all, 173 original assays were submitted for re-assay

from SR-160 (66), SR-197 (58), and SR-213 (49). At least three different standards were alternately used

at any given time for drilling prior to the initial resource estimate; however, only one at a time was

submitted for drilling associated with the resource update (SR-140 to SR-233).

For all drilling, four certified blank pulps have been inserted into the sample stream at a rate of one for

every 40 samples prior to shipment to the preparation laboratory in Medellín.

A field or preparation duplicate was submitted with samples from the deposit at a rate of one for every

20 samples. All duplicates were assigned a consecutive sample number after the original, and both were

analyzed. Field duplicates consisted of quarter-core splits (also quarter-core original) in the first two

phases of drilling (SR-001 to SR-065), half-core splits (also half-core original) in the next phase of drilling

(SR-066 to SR-139), and quarter-core splits for the remainder of the drilling (SR-140 to SR-233).

Preparation duplicates were splits of coarse rejects.

In addition to the standards, blanks, and duplicates, 169 samples were sent for check assays to SGS at

their laboratory in Lima, Perú; these were discussed in MDA’s first Technical Report (Lindholm and Schlitt,

2013a). Two batches totalling 131 pulp samples and representing a randomly selected 5% of original

mineralized samples were submitted to ALS in Lima for check assaying in April 2013. Samples were

analyzed by 50 g fire assay with an AA finish (ALS code Au-AA26), with analysis by 50 g fire assay with a

gravimetric finish (ALS code Au-GRA22) for samples exceeding 100 ppm Au. QA/QC results are

discussed in Section 12.2.

11.5 Bulk Density and Specific Gravity

Bulk wet-density data were regularly collected from drill holes SR-001 to SR-155, with measurements for

selected samples taken about every 10 m. The samples were weighed in air and then, after wrapping in

cling film to prevent any water absorption, were suspended in water, and the weight of the water displaced

was measured, using a simple overflow system.

At MDA’s request, 65 bulk specific gravity measurements were performed by Red Eagle Mining on oven-

dried samples. A similar process and apparatus for measuring bulk wet densities was used, except that

samples were dried, and the weight of wrapped and suspended sample submerged in water rather than

the weight of the water displaced was used to calculate specific gravities.

Page 11.5




AMENDED NI 43-101 TECHNICAL REPORT For drill samples from holes SR-156 through SR-233, bulk dry densities were obtained using the same

procedures that were used for wet-density determination, except that all samples were oven dried at

125°C. Fresh rock was usually dried between two and three hours, saprock for three to five hours, and

saprolite no less than 12 hours. According to Red Eagle Mining, only saprolite samples were wrapped in

plastic before being suspended in water.

11.6 Summary Statement

The San Ramon deposit drill-core sampling procedures, sample security protocols, and analytical

methods are acceptable. A QA/QC program has been in place for all drilling that is adequate for

evaluation of assay quality. The overall results of the QA/QC program indicate that the analytical data are

of sufficient quality for use in the resource estimate.

Page 12.1




AMENDED NI 43-101 TECHNICAL REPORT 12.0 DATA VERIFICATION

This section describes data verification for drilling by Red Eagle Mining that comprises the database for

the resource estimate described in this report but not holes drilled since May 2013 that post-date the

resource estimate and are discussed in Section 10.1.

12.1 Database Audit

12.1.1 Drill-Collar Audit

Initial topography for the project was provided by Terranalisis, Ltda., based in Santiago, Chile, and MPX

Geophysics Colombia SAS (“MPX”), based in Medellín, Colombia. After discrepancies were noted

between topography and drill-hole collars and after transferring project data to the currently-used

coordinate system (GCS_Bogotá, Datum: D_Bogotá), the project surveyor, MSc. Ricardo Lozano Botache

with Estudio T-Rural Consultores, was contracted to perform a ground survey for topography. MPX then

performed a Lidar topographic survey that covers concession B7560005, on which the San Ramon

deposit is located. A visual comparison of collars (initial and 2013 drilling) and the Lidar topography

revealed a reasonable correlation. This topography, with some of the Terranalisis topography outside the

limits of the Lidar survey attached to include areas where pits and mine facilities might be located, is used

for the current modelling and resource estimation work.

Certified collar coordinates for all except four of the 238 holes drilled at the San Ramon deposit were

provided directly to MDA by MSc. Ricardo Lozano Botache. All coordinates were compared to Red Eagle

Mining’s database, and no errors were found.

12.1.2 Down-Hole Survey Audit

No original REFLEX data were available for audit of down-hole survey data for holes drilled prior to the

initial resource estimate. The data recorded on individual spreadsheets for each drill-hole log were the

only data available for audit; MDA compiled these data into a single spreadsheet and compared them to

Red Eagle Mining’s database. Of a total of 175 records in Red Eagle Mining’s database, 156 records

(89%) were audited. A total of four significant errors, three in depths and one in dip, and 20 insignificant

discrepancies were found. All errors and discrepancies were corrected in the final database used for the

resource estimate. Also, 12 additional records found on spreadsheets for the drill-hole logs were added to

the database.

Page 12.2




AMENDED NI 43-101 TECHNICAL REPORT During the course of the audit, MDA noted that there were radical bends in the upper 100 m of some drill

holes. No REFLEX readings were taken at or just below the collars of the drill holes, nor were azimuth

and dip readings taken with a compass while the rig was set up on the pad. The planned azimuths and

dips were used for the collar orientation in the Red Eagle Mining database. Red Eagle Mining reviewed

the cases where the record at the collar differed from the first REFLEX record at 100m by more than 10°

in azimuth and 5° in dip, and modified as necessary. Many orientations at the collar were replaced by

Reflex records taken at 100 m depth; any errors discovered were corrected; and other records were

unchanged and accepted. The resulting down-hole survey database appears reasonable, with radical

bends in the upper parts of some holes removed, although some insignificant discrepancies of less than

5° could still exist. MDA recommended that direct azimuth and dip readings be recorded while the drilling

rig is set up on site and used as the survey record at 0 m depth.

Red Eagle Mining supplied MDA with original REFLEX down-hole surveys in both .pdf and spreadsheet

format for all new drilling (SR-140 to SR-233). The logs indicate that Red Eagle Mining staff performed

the down-hole surveys using a leased REFLEX EZ-Trac tool. In addition, similar files were provided for

some of the previous drilling (SR-059 to SR-069). MDA compiled all data from the .pdf files into a

spreadsheet for comparison to Red Eagle Mining’s database. In total, the files contained 388 records

from 90 drill holes.

Initial evaluation of the data revealed numerous records that were not used in Red Eagle Mining’s

database, and a significant number of modified azimuth, dip, and depth values. It was apparent that,

beyond correction of azimuths for magnetic declination, the down-hole survey data had been manually

manipulated to a large degree. The data modifications were done based on reasonable criteria; however,

none of the changes were documented and none could be verified. MDA reconstructed the down-hole

survey database for the new drilling, as well as for holes SR-059 to SR-069, from the .pdf files and applied

a correction for magnetic declination as needed. Records were removed only if anomalous magnetic

strength measurements suggested the readings were suspect (35 records) or if there were implausible

differences in azimuth or dip over a short distance (9 records, mostly first or last reading). A total of

46 records were added to Red Eagle Mining’s database, and all work was documented and agreed to by

Red Eagle Mining staff.

The data reconstructed from original REFLEX logs were merged with the down-hole survey database

previously audited for the initial resource estimate. The resulting down-hole survey information in the

database is considered suitable for use in the resource estimate.

12.1.3 Geological Data Audit

The down-hole geological data available to MDA for data verification and modelling include lithology,

weathering (saprolite), redox, alteration, geotechnical, and some structural data provided by Red Eagle

Mining. Much of this information, particularly the lithology and weathering data from earlier drilling

programs, had been normalized since the last resource estimate. During the process of geologic and

mineral domain modelling, core photos were consulted and the geological data were found to be

reasonably accurate. Red Eagle Mining’s geologic data were confirmed, although differences in

interpretations will always occur.

Page 12.3




AMENDED NI 43-101 TECHNICAL REPORT 12.1.4 Assay Database Audit

For the initial database audit (SR-001 to SR-139), MDA downloaded all assay certificates directly from

Acme’s website. These certificates were compiled into one spreadsheet, and quality-control records were

removed. Using the data for sample intervals with multiple assay types, an evaluation of standard fire

assays versus metallic-screen data and then versus fire-assay gravimetric data was undertaken, and it

was determined there is no apparent bias between groups that would preclude use of one assay type over

another. Accepted assays were determined, in agreement with Red Eagle Mining staff, for use in the

resource estimate; metallic screen assays, fire assay gravimetric, and standard fire assays were used in

that order of preference.

For the initial database audit, the assay data compiled by MDA from Acme certificates were compared to

Red Eagle Mining’s database. A total of 13,350 samples from 203 assay certificates were compiled; these

included QA/QC samples. These assays were merged with Red Eagle Mining’s assay database, which

contains 11,665 samples, excluding QA/QC. Only 118 assays in the database did not have matching data

from assay certificates, so 99% of the database was audited. No significant errors or discrepancies were

found. The only systematic discrepancies, though all are considered insignificant, were for values below

detection, which were entered into the Red Eagle Mining database as half the detection limit.

For the current estimate, MDA again downloaded all assay certificates for SR-140 to SR-233 directly from

Acme’s website, compiled them into one spreadsheet, and removed quality-control records. The new data

contained only standard fire assay-AA and fire assay-gravimetric assays, of which the latter were

analyzed only for over-limit values. MDA and Red Eagle Mining staff agreed that standard fire assays

should be replaced by fire assay-gravimetric assays when fire assay-gravimetric assays were available.

The assay data compiled by MDA from Acme certificates were compared to Red Eagle Mining’s database.

A total of 5,651 samples from 96 assay certificates were compiled; these included QA/QC samples.

These assays were merged with Red Eagle Mining’s assay database, which contains 5,320 new samples,

excluding QA/QC. No assays in the database were unmatched by data from assay certificates, so 100%

of the database was audited. No significant errors or discrepancies were found. Below detection limit

assays which had been entered into the Red Eagle Mining database as “0”, were replaced with values of

“0.0025”, which is half the detection limit.

MDA reviewed all available core photos and recorded intervals of no core recovery. These intervals of no

recovery had been recorded as part of one or both adjacent sample intervals, so that a portion of the

assay represented recovered core plus the unrecovered interval. In these cases, MDA added intervals

with values that indicate ‘no assays’ to the database, and shortened the adjacent intervals so they

represent the proper length of assayed core. MDA also merged specific gravity/density intervals to assay

intervals, which required some splitting of assay intervals, so the data could be loaded into the assay

database. Finally, after review of Red Eagle Mining’s QA/QC summary, MDA replaced initial assays

associated with standard failures in SR-160, SR-197, and SR-213 with pulp re-assay values.

All assay database changes were made in communication with Red Eagle Mining staff and were mutually

agreed upon. The final, combined, and fully audited assay database is considered adequate for use in the

resource estimate.

Page 12.4




AMENDED NI 43-101 TECHNICAL REPORT 12.1.5 Specific Gravity Data

The 65 specific gravity measurements used for the initial resource estimate, including all raw data used for

calculations, were provided by Red Eagle Mining. MDA checked all calculations, and no errors were

found. The specific gravity measurements were performed after MDA’s site visit in September 2012, so

MDA was unable to verify the process. However, MDA did inspect the similar process and apparatus for

measuring bulk wet densities during the site visit, and these were determined to be adequate. Also, Red

Eagle Mining reportedly used procedures for determining specific gravities of oven-dried samples that

were provided by MDA. Specific gravity/density measurement procedures are discussed in Section 11.5.

The 862 additional dry-density tests on samples from SR-157 to SR-233 were provided by Red Eagle

Mining with raw data. MDA again checked all calculations; three errors were found and corrected. The

apparatus and procedure used to determine dry density were the same as those used for wet-density

determination, except that an oven-dried sample was used (Section 11.5). The wet-density procedure

was observed during MDA’s initial site visit in September 2012 and is, with the exception of the lack of

sample drying, considered to be adequate for measurement of dry densities. The drying process for dry-

density determination was never verified. The specific gravity/dry-density data are considered suitable for

use in the resource estimate.

12.1.6 Geotechnical Data

All geotechnical data (for SR-001 to SR-233), which include core recovery, RQD, and hardness, were

provided by Red Eagle Mining. The data were collected on core-run intervals, and measured lengths

used to calculate recovery and RQD were included. MDA checked all calculations and found numerous

discrepancies. All were corrected in agreement with Red Eagle Mining. These modifications included

addition of missing data or calculations, correction of calculated interval lengths and recovery / RQD

percentages, removal of extraneous records, and movement of decimal points in lengths or percentages.

The data were then merged with assay intervals, and recovery and RQD values of split intervals were

weight averaged to produce the combined values; these data were used to evaluate sample integrity.

RQD values of “0” were entered for saprolite and saprock, regardless of the condition of the core. This is

somewhat misleading, since an RQD value of “0” indicates that all recovered core pieces were smaller

than 10cm diameter, which suggests the sample integrity is poor. However, rock that is converted mostly

or completely to clay is treated as a soil rather than a rock in terms of engineering properties, so RQD

does not apply. It is more appropriate to assign no value to RQD in heavily weathered rocks. MDA

replaced RQD values of “0” with “-1” for all records that had no measured RQD lengths. However, it was

impractical to investigate all cases where RQD values were “0”, so many of the values of this type in Red

Eagle Mining’s database have been retained.

In general, the geotechnical data are considered adequate for use in the resource estimate. However, the

RQD values of “0” in saprolite and saprock must be taken into account for statistical evaluation of sample

integrity.

Page 12.5




AMENDED NI 43-101 TECHNICAL REPORT 12.2 Quality Control and Quality Assurance

The quality control and quality assurance program implemented by Red Eagle Mining involved the use of

standards, blanks, field and preparation duplicates, and check assays of pulp duplicates at a second lab.

Procedures are discussed in Section 11.4. The standards used have been supplied by CDN Resource

Laboratories Ltd., Calgary, Canada.

12.2.1 Standards Assays

Standards data were supplied to MDA in the assay database by Red Eagle Mining for all assayed drill-

hole samples. Certificates containing standards assays, which were identified using Red Eagle Mining’s

spreadsheet, were also downloaded directly from Acme’s website.

Initial Resource Estimate (SR-001 to SR-139)

Prior to the initial resource estimate (SR-001 to SR-139), the database contained records for 351 analyses

of six certified standards. The range of values of the commercial standards was appropriate to the

samples being processed. MDA evaluated the standards using charts similar to the common Shewhart

charts. An example of one such chart appears below as Figure 12.1.

Figure 12.1 Control Chart for Gold in Standard #3 (SR-001 to SR-139)

Page 12.6




AMENDED NI 43-101 TECHNICAL REPORT In Figure 12.1 the red line indicates the accepted value of the standard, and the blue lines indicate the

mean plus and minus two standard deviations. The purple lines indicate the control limits, set at the mean

plus and minus three standard deviations, using the statistics provided by the manufacturer of the

standard. Analyses falling outside the control limits are deemed to be failures.

The analyses displayed in Figure 12.1 include one failure above three standard deviations, representing a

1.3% failure rate for that standard. The chart also indicates that six analyses fall outside the two standard

deviation limits, all of them high. The analysis shows that the lab has an overall high bias of 2.5% for

analyses of this standard.

Charts such as Figure 12.1 were prepared for all of the standards. Only the one example is provided in

this report. A summary of the results obtained for all of the standards appears in Table 12.1.

Table 12.1 Summary of Gold Results for Standards (SR-001 to SR-139)

Standard ID Count Failures

±3SD Consecutive ±2SD ±2SD

#1 69 9 1 17

#2 37 2 1 6

#3 77 1 0 7

#4 16 2 0 3

#5 131 0 1 14

#6 21 0 0 2

Counts 351 14 3 49

Percent 4.0 0.8 14.0

In addition to analyses falling outside the three standard deviation control limits, consecutive analyses

outside of two standard deviations may also be considered failures. The tabulated results indicate a total

of 17 failures (of both types) of 351 standards analyses, for a failure rate of 4.8%. While the error rate is

not unusual, there is no evidence that any follow-up has been done to investigate the errors and, where

necessary, obtain re-analyses of suspect batches. The lack of follow-up is a deficiency. It is interesting

that more than half of the failures occur on one standard sample (Standard #1), which accounts for 10 of

the 17 failures; this may indicate a problem with the standard rather than the laboratory results.

Page 12.7




AMENDED NI 43-101 TECHNICAL REPORT Although not considered failures, having 14.0% of the analyses fall outside the two standard deviation

limits is somewhat high. The majority of these are high relative to the accepted value of the standard.

The high rate of values above two standard deviations (positive relative differences) is more likely a

reflection of the overall high bias of analyses of the standards relative to the certified accepted values.

Any statistically significant group of analyses of a standard performed by a single lab will have a bias

relative to the accepted or expected value for that standard. There is a positive mean of difference for five

of the six standards that range from 0.8% to 3.5%, and average overall mean is about +1.7%. Average

means of difference may be a general indicator of bias but are often heavily skewed by high assay values,

so are not always an accurate method of quantifying bias.

Resource Update (SR-140 to SR-233)

For the current resource update (SR-140 to SR-233), the database contained standard sample assay

records for 171 analyses of three certified standards. The range of values of the commercial standards

was limited; the accepted value of two of the standards was 4.075g Au/t and 3.85g Au/t, and the other was

7.72g Au/t. Also, only one standard was used over any given period of time. The same standard was

submitted with samples repeatedly until the supply of that reference material was exhausted, and then

another was used. An example of one of the standard charts appears below as Figure 12.2.

Figure 12.2 Control Chart for Gold in Standard #7 (SR-140 to SR-233)

Page 12.8




AMENDED NI 43-101 TECHNICAL REPORT In all, the analyses of standards produced two failures above three standard deviations, which represent a

1.2% failure rate for the program. There were also eight (4.6%) analyses that fell outside the two standard

deviation limits, all of them high, but none of them consecutive. While these do not represent failures, it is

a consequence of a slight high analytical bias. This is corroborated by visual inspection of the charts; the

majority of data on all three charts plot above the accepted value of the respective standards. In Figure

12.2, the standard assays plot below the accepted value in November 2012, and over time drift above the

line and back below by February 2013. The drift stays within acceptable limits and might be attributed to

self-correction by the lab in response to results of internal QA/QC programs.

The error rate of 1.2% is not unusual. More importantly, Red Eagle Mining immediately identified the

failures and followed up by re-numbering and re-submitting all pulps in the batches associated with each

failure. Pulp standards and blanks were submitted with these checks at the same rate as with the original

batches. MDA examined the assay results of the reference materials that were sent with the re-submitted

pulps and confirmed there were no failures. Red Eagle Mining had identified a third standard failure in the

original assays but determined that the discrepancy resulted from a mislabelled standard number. The

pulps in that batch were submitted for re-assay anyway. A total of 173 original assays were replaced by

the re-assays for all three batches from SR-197, SR-213, and SR-160 (the non-failure).

Initially, Red Eagle Mining identified as many as 12 standard failures outside the three-standard deviation

threshold. However, 10 of these were determined to be mislabelled standard numbers. In all, there were

22 cases of mislabelled standards or blanks. All plotted within the three-standard deviation threshold on

the correct charts. MDA examined these cases and determined that the conclusion that they are not

standard failures is reasonable. However, the high rate of mislabelled reference materials is problematic

and suggests that procedural improvements should be made in order to eliminate sample-handling errors.

Summary of Results

The results of the standards analyses for all drilling support the acceptability of Red Eagle Mining’s assays

as part of the basis for the resource estimate described in this report. Initially, there was a lack of follow-

up and explanation for the approximately 5% of standard failures, which detracted from the overall quality

of the support for the data from drill holes SR-001 to SR-139; this was taken into account in MDA’s

classification of the resource estimate. However, for SR-140 to SR-233, Red Eagle Mining immediately

addressed the few standard failures that occurred, which increases the level of confidence in the

associated assays.

12.2.2 Field and Preparation Duplicate Sample Assays


For the initial resource estimate, a total of 316 field duplicate samples and 306 preparation duplicate

samples were submitted with core samples from San Ramon deposit. Field duplicates consisted of

quarter-core splits (duplicating quarter-core original samples) in the first two phases of drilling and half-

core splits (duplicating half-core original) in the last phase of drilling. Preparation duplicates are splits of

coarse rejects. All original and duplicate sample pairs were analyzed by Acme, and assay results were

downloaded directly from Acme’s website by MDA.

Page 12.9




AMENDED NI 43-101 TECHNICAL REPORT Figure 12.3 depicts a relative-difference graph which shows the difference (plotted on the y-axis) of each

Acme field duplicate fire assay relative to its paired original Acme analysis. Calculation of relative

difference in this way gives a worst-case value for the difference, but also gives a better presentation of

the real variability. The x-axis of the graph plots the mean gold grades of the paired data in a sequential,

non-linear fashion ordered by increasing grade. The red line shows the moving average of the relative

differences of the pairs and provides a visual guide to trends in the data. Positive relative-difference

values indicate that the duplicate analysis is greater than the original. Pairs for which both original and

duplicate values are below the detection limit have been excluded to shorten the graph. Comparison of

means and basic statistics are given in Table 12.2.

Figure 12.3 Relative Percent Differences for Field Duplicates (SR-001 to SR-139)

Table 12.2 Comparison of Means for Field Duplicates (SR-001 to SR-139)

Gold (mean of pairs > 0.0025 ppm)

With 5 extreme outliers removed

Valid N Median Mean Wt Avg Std.Dev. CV Minimum Maximum Units

Original 239 0.018 0.194 0.177 0.673 3.47 0.003 5.250 ppm

Duplicate 239 0.022 0.209 0.195 0.750 3.59 0.003 6.900 ppm

Difference

22% 8% 10%

Page 12.10




AMENDED NI 43-101 TECHNICAL REPORT The apparently large relative differences of ±200% on the left side of Figure 12.3for means of pairs at and

below 0.005 g Au/t are due to original or duplicate samples below the detection limit. For mean of pair

calculations and statistics, assays below detection were entered as half the detection limit, or 0.0025 g

Au/t. The small differences in pairs that include these values are large relative to the mean of pairs and

therefore return misleadingly pronounced relative differences on the chart. These are not a reflection of

significant bias or variability in the samples and can be ignored.

Five extreme outliers that had relative differences greater than 1,000% were excluded from the graph and

basic statistics. It should be noted that a large portion of the field duplicate data are below potentially

mineable grades. Only 25 of the 239 duplicate pairs shown in Figure 12.3 have average grades at or

above 0.3 g Au/t. Any observed variability or apparent bias between original and duplicate sample assays

at lower grades are immaterial. At grades above 0.3 g Au/t, variability is still apparent, and there is no

obvious bias that is not driven by extreme outlier grades, however, the data set is too small to draw

meaningful conclusions.

Figure 12.4 plots the relative difference between original and duplicate samples versus the mean of

sample pairs for preparation duplicates. Pairs for which both original and duplicate values are below the

detection limit have been excluded to shorten the graph. Comparison of means and basic statistics are

given in Table 12.3.

Figure 12.4 Relative Percent Differences for Preparation Duplicates (SR-001 to SR-139)

Table 12.3 Comparison of Means for Preparation Duplicates (SR-001 to SR-139)

Page 12.11








Valid N Median Mean Wt Avg Std.Dev. CV Minimum Maximum Units

Original 215 0.027 0.533 0.498 2.626 4.93 0.003 31.500 ppm

Duplicate 215 0.027 0.539 0.505 2.572 4.77 0.003 30.100 ppm

Difference

0% 1% 1%

Two extreme outliers that had relative differences greater than 1,000% were excluded from the graph and

basic statistics. Again, most of the preparation duplicate data are below potentially mineable grades.

Only 37 of the 221 duplicate pairs shown in Figure 12.4 have average grades at or above 0.3 g Au/t. At

grades above 0.3 g Au/t, some variability and little or no bias are apparent.

Examination of the full data sets shows that, unlike the chart for field duplicates, the chart for preparation

duplicates has very little bias towards either the original or duplicate value. This is also indicated by the

higher mean of relative differences and absolute value of relative difference means for field duplicates (8%

and 105%) relative to preparation duplicates (1% and 59%). Both graphs indicate that variability

diminishes in the higher grade ranges.

12.2.3 Resource Update (SR-140 to SR-233)

All original and field duplicate sample pairs were analyzed by Acme, and assay results were downloaded

directly from Acme’s website by MDA.

For the most recent drilling, a total of 114 and 107 field and preparation duplicates, respectively, were

submitted with samples from the San Ramon deposit. Field duplicates consisted of quarter-core splits

(original was half-core). All original and duplicate sample pairs were analyzed by Acme, and assay results

were downloaded directly from Acme’s website by MDA. Figure 12.5 plots the relative difference between

original and field duplicate samples versus the mean of the sample pairs. Comparison of means and

basic statistics are given in Table 12.4.

Page 12.12




AMENDED NI 43-101 TECHNICAL REPORT Figure 12.5 Relative Percent Differences for Field Duplicates (SR-140 to SR-233)

Table 12.4 Comparison of Means for Field Duplicates (SR-140 to SR-233)



Count Median Mean Std. Dev. CV Min. Max. Units

Original 77 0.038 0.294 1.031 3.507 0.0025 6.920 ppm

Duplicate 77 0.031 0.288 0.930 3.232 0.0025 6.269 ppm

Diff.

-18% -2%

0.000 -9%

Fifteen outliers that had relative differences greater than 400% were excluded from the graph and basic

statistics. This is an atypically large number of outliers to remove. The moving average line alternates

between ~100% bias with duplicate greater than original and vice versa. At material grades above 0.3 g

Au/t, there are only 13 pairs, which is not a large enough data set to draw meaningful conclusions. The

absolute value of the relative difference above 0.3 g Au/t is 81%, but the high value is driven by two

relatively large means of pairs.

Page 12.13




AMENDED NI 43-101 TECHNICAL REPORT A total of 107 preparation duplicate samples, which consisted of coarse reject splits, were submitted with

samples from the San Ramon deposit. Figure 12.6 plots the relative difference between original and

duplicate samples versus the mean of sample pairs for preparation duplicates. Comparison of means and

basic statistics is given in Table 12.5.

Figure 12.6 Relative Percent Differences for Preparation Duplicates (SR-140 to SR-233)




Valid N Median Mean Std. Dev. CV Minimum Maximum Units

Original 79 0.067 0.242 0.564 2.327 0.0025 3.298 ppm

Duplicate 79 0.054 0.241 0.563 2.340 0.0025 3.147 ppm

Difference

-19% -1%

Page 12.14




AMENDED NI 43-101 TECHNICAL REPORT Four outliers that had relative differences greater than 200% were excluded from the graph and basic

statistics. Unlike the chart for field duplicates, the chart for preparation duplicates shows only minor bias

towards either the original or duplicate. At material grades above 0.3 g Au/t, there are only 10 pairs, which

is not a large enough data set to draw meaningful conclusions. The absolute value of the relative

difference for all data is 36%, which decreases to 12% for data above 0.3 g Au/t, suggesting decreasing

variability at higher grades.

Summary of Results

The high positive and negative relative differences observed in field, and to a lesser extent, preparation

duplicate pairs is indicative of natural heterogeneity of gold within the San Ramon deposit (and gold

deposits in general). This imparts some risk to the resource estimate due to local grade variability,

although the additional risk is difficult to quantify.

Variability decreases significantly, and overall bias is less apparent from field duplicates to preparation

duplicates. Some of the natural heterogeneity inherent in the samples that was apparent in analyses of

the field duplicates was likely homogenized by the first phase of crushing that produced the coarse rejects.

Variability appears to diminish in the higher grade ranges.

It is important to note that the majority of duplicate pairs are below potentially mineable grades. Although

the amount of variability between original and duplicate sample assays decreases at higher grades, it is

still present, and it may not be appropriate to evaluate heterogeneity or to determine risk based on results

of the small portion of the data set that is material.

Aside from the bias noted in the field duplicate samples with higher duplicate grades relative to original

assays, these data suggest that sampling and subsampling protocols are adequate.

12.2.4 Blanks

The QA/QC data set for the Red Eagle Mining drilling includes 320 analyses of blank pulps for SR-001 to

SR-139 and 173 analyses for the remainder of the drilling that were included in sample shipments. A total

of four certified blanks were purchased by Red Eagle Mining for insertion in both drilling programs.

Results of the analyses are presented in Figure 12.7 and Figure 12.8. Because these samples are pulps,

they are considered “analytical” blanks and are essentially standards with values of “0.” No coarse blanks,

which would also test for contamination during the sample preparation, were included in the QA/QC

program. MDA recommends that Red Eagle Mining submit coarse blanks in the sample stream,

preferably after mineralized samples, in future drilling programs.

Page 12.15




AMENDED NI 43-101 TECHNICAL REPORT Figure 12.7 Gold in Pulp Blanks vs. Preceding Sample (SR-001 to SR-139)

Figure 12.8 Gold in Pulp Blanks vs. Preceding Sample (SR-140 to SR-233)

Page 12.16




AMENDED NI 43-101 TECHNICAL REPORT In Figure 12.7 and Figure 12.8, gold in the blank pulp sample is plotted on the vertical axis versus gold in

the immediately preceding sample, plotted on the horizontal axis. MDA has used a line arbitrarily set at

five times the detection limit to designate “failures” in Figure 12.7 and Figure 12.8. No tendency for in-lab

analytical contamination is suggested by the two plots. However, since no coarse blanks were used,

contamination during sample preparation has not been tested, which decreases the level of confidence in

the assays.

12.2.5 Check Assays


For the initial resource estimate, a total of 169 pulp check samples, originally analyzed by Acme, were

submitted to SGS del Perú, S.A.C. in Lima, in order to test the analytical accuracy of Acme’s assays. Of

these, 20 did not match sample numbers in the drill-hole database and were not included in the check-

assay evaluation. Two of the remaining 149 samples were pulps from field duplicates and are included in

the evaluation. No pulp standards or blanks were submitted with the pulp checks.

Figure 12.9 plots the relative differences between the original Acme assays and the SGS pulp check

assays versus the mean of sample pairs. The red line is a 20-sample-pair moving average. Positive

relative difference values (above the “0” line) indicate that the SGS pulp check assay is greater than the

Acme original assay.

Figure 12.9 Relative Percent Differences for Pulp Check Assays (SR-001 to SR-139)

Page 12.17




AMENDED NI 43-101 TECHNICAL REPORT Fourteen outlier sample pairs, or ~10% of the pulp checks, were removed from the chart in Figure 12.9.

Even without these outliers, there is a significant amount of variability between Acme and SGS assays.

Visual inspection of the data points indicates a significant bias up to a mean grade of the pairs of ~0.4g

Au/t, indicating the original Acme assays are higher relative to the SGS checks in this range. Above ~0.4g

Au/t, there is no definitive bias. The low-grade bias is reflected in the means of the analyses of the two

laboratories (Table 12.6), but that bias is at grades generally below economic cut-off. Table 12.7 shows

the descriptive statistics of the data at means of the pairs greater than 0.4g Au/t.

Table 12.6 Comparison of Means for Pulp Check Assays, Mean of Pairs = 0.01 to 0.4g Au/t

(SR-001 to SR-139)

Mean 0.01 to 0.4 Mean Original Duplicate Diff. Rel. Diff. A.V. Rel. Diff.

Count 29 29 29

29 29

Mean 0.194 0.228 0.160 -30% -69% 82%

Median 0.176 0.205 0.140 -32%

Std. Dev. 0.081 0.098 0.088

CV 0.416 0.429 0.549

Min. 0.105 0.088 0.064 -27% -469% 0%

Max. 0.403 0.415 0.390 -6% 110% 469%

Table 12.7 Comparison of Means for Pulp Check Assays, Mean >0.4g Au/t

(SR-001 to SR-139)

Mean >0.4 Mean Original Duplicate Diff. Rel. Diff. A.V. Rel. Diff.

Count 47 47 47

47 47

Mean 5.948 5.930 5.964 1% 6% 53%

Median 2.870 2.191 2.692 23%

Std. Dev. 7.024 7.140 7.125

CV 1.181 1.204 1.195

Min. 0.403 0.206 0.390 89% -268% 1%

Max. 28.310 27.220 29.400 8% 382% 382%

Since no standards were submitted to SGS with the sample pulps, the accuracy of the check assays

cannot be evaluated, and there is no way to determine which laboratory is returning the more accurate

data. However, the original samples sent to Acme had standards inserted in the sample stream and

those, on the basis of those standards, were determined to be reliable. It should be noted that analyses of

standards assays did also indicate a slight positive bias of Acme assays overall. The high bias of Acme

original assays relative to the pulp checks is apparent up to ~0.4g Au/t. The lack of bias above this grade

is encouraging, although there is a risk that potentially mineable blocks in the resource may be influenced

by the possible bias in the lower-grade samples.

Page 12.18




AMENDED NI 43-101 TECHNICAL REPORT Resource Update (SR-140 to SR-233)

Since the original estimate and QA/QC analyses, 131 pulp check samples, originally analyzed by Acme,

were submitted to ALS Colombia in Medellín, in order to test the analytical accuracy of Acme’s assays.

Pulp standards, duplicates, and blanks were submitted with the pulp checks. MDA evaluated the results

of the QA/QC data and confirmed that there were no standard or blank failures and that the duplicate pair

assays were within reasonable limits. Figure 12.10 plots the relative differences between the original

Acme assays and the ALS pulp check assays versus the mean of sample pairs.

Figure 12.10 Relative Percent Differences for Pulp Check Assays (SR-140 to SR-233)

Four outlier sample pairs were removed from the chart in Figure 12.10. Some variability between Acme

and ALS assays is indicated; the absolute value of the relative differences is 16% for all data (Table 12.8).

By progressively excluding lower-grade pairs, the variability decreases to 10% for assays above 10g Au/t.

There are a numerous individual pair differences that exceed 30% above and below the 0%-line, but the

frequency of these decreases perceptibly above ~7.0 g Au/t. Visual inspection of the data points indicates

little to no definitive bias.

Page 12.19




AMENDED NI 43-101 TECHNICAL REPORT Table 12.8 Comparison of Means for Pulp Check Assays, Mean of Pairs > 0.10g Au/t

(SR-140 to SR-233)

All Pairs (>0.10) Mean Original Duplicate Diff. Rel. Diff. A.V. Rel. Diff.

Count 127 127 127

127 127

Mean 9.510 9.507 9.512 0% -2% 16%

Median 4.814 4.526 4.560 1%

Std. Dev. 18.970 18.603 19.378

CV 1.995 1.957 2.037

Min. 0.112 0.123 0.100 -19% -90% 0%

Max. 163.100 157.700 168.500 7% 108% 108%

The extreme variability and bias shown in the check assay data for the initial resource estimate were not

evident in the second set of check assays. This could be due in part to differences between the check

laboratories. The results of both sets of check assay data are indicative of the higher variability and bias

at grades below 0.4g Au/t, and the lower variability with little to no bias at higher grades. This is important,

because the most significant variability occurs at grades at or below ~0.3 g Au/t, and well below the

potential underground cut-off grade of 1.2 g Au/t.

As part of the sample integrity evaluation, MDA examined the relationship between gold grades and core

recoveries. Figure 12.11 is a histogram that plots core recovery in bins of 5% versus gold grade. All

samples are included, regardless of lithology, oxidation state, or degree of weathering. The approximate

average grade of samples at recoveries of 70% or greater is 1.0 g Au/t. However, below 70% recovery,

the gold grade decreases significantly to about 0.4 to 0.6 g Au/t. This may be the result of loss of gold

during core drilling. In soft rocks, such as saprolite with high clay contents, free gold may be washed out

of the sample by circulated drilling fluids and formation water, causing the observed grade decrease in

poor-recovery samples. There is a precedent for this in that the artisanal miners extracted gold from soft

material near the surface. Additionally, free gold or gold associated with sulfide minerals in soft, clayey

gouge in the shear zone could be washed out, resulting in poor recovery and preferential loss of gold in

the sample. If this is the case, the assays, particularly for samples of saprolite or gouge zones with low

recoveries, may be understating the gold grade.

Page 12.20




AMENDED NI 43-101 TECHNICAL REPORT Figure 12.11 Core Recovery versus Gold Grade

Table 12.9 gives the mean core recoveries for the various lithologic, weathering, and oxide types. The

mean core recovery of ~81% in saprolite/oxidized rock is much lower than that of unweathered and

unoxidized rocks, which ranges from 94% (shear zone) to 98% (outside shear zone). Given the lower

average gold grade for recoveries below 70%, due possibly to loss of gold in heavily weathered, oxidized

and possibly intensely sheared samples, assays associated with the lower recoveries in saprolite and

oxide, and to a lesser extent, in gouge zones, may be understating the gold grade in these rocks.

Table 12.9 Core Recoveries for by Lithologic, Weathering and Oxidation Types

Core Recovery by Lithology, Weathering and Oxidation State

Count Median (%) Mean (%)

All Saprolite 3,207 86.7 80.5

All Saprock 1,472 92.5 87.5

All Oxide 3,869 88.8 81.9

Transition - Shear Zone 210 89.8 86.8

Transition - Country Rock 800 93.3 88.9

Shear Zone - Unweathered, Unoxidized 4,308 96.7 94.0

Schist - Unweathered, Unoxidized 376 99.9 98.1

Mafic Dikes - Unweathered, Unoxidized 229 99.7 96.5

Felsic Dikes - Unweathered, Unoxidized 322 98.1 95.0

Granodiorite - Unweathered, Unoxidized 5,670 99.7 97.3

All Data 17,388 97.6 92.3

0

200

400

600

800

1000

1200

1400

1600

1800

0.00

0.20

0.40

0.60

0.80

1.00

1.20

1.40

1.60

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 80 85 90 95 100

Co

un

t

Au

(p

pm

)

Core Recovery (%)

San Ramon Recovery vs Gold Grade

Page 12.21




AMENDED NI 43-101 TECHNICAL REPORT 12.3 MDA Independent Verification of Drill-Collars

12.3.1 Initial Resource Estimate (SR-001 to SR-139)

During the initial site visit in November of 2012, MDA took GPS readings on 10 drill pads and 15 drill-hole

collars in the field to spot-check coordinates in Red Eagle Mining’s collar table. MDA used a Garmin -

Oregon 550 t non-differential GPS; the device owner’s manual indicates it is accurate to within 10 m.

MDA also recorded GPS measurements taken by the Red Eagle Mining geologist who accompanied MDA

in the field, for comparison.

MDA’s independent collar measurements are intended only as a check of reasonableness against the

collar coordinates in the database. Because Red Eagle Mining was using the older coordinate system at

the time MDA took the measurements, and the collar coordinates in the Red Eagle Mining database are in

the new coordinate system, elevations of MDA’s readings were decreased by 21.67 m. Northings and

eastings were not adjusted, but there is a difference of up to 1.5m between the older and current

coordinate systems. All but one (northing for SR-023) of the northings, eastings, and elevations measured

by MDA compared to the database are within the level of accuracy of the GPS, and all are within 12 m.

This is the case despite the approximated collar locations for SR-019/043 and SR-052 on the drill pads

based on the location of washed material and cuttings on the pads. Also, GPS measurements of

elevations tend to be less accurate in steep terrain where most acquired satellites are more directly

overhead. MDA considers the collar coordinates in Red Eagle Mining’s database to be acceptable based

on the results of the independent GPS checks.

Resource Update (SR-140 to SR-233)

During the second site visit in May 2013, GPS readings were taken on 10 drill sites and 16 new drill-hole

collars in the field to spot check coordinates in Red Eagle Mining’s collar table. MDA used a Garmin

GPSmap® 62s non-differential GPS borrowed from Red Eagle Mining staff; the device’s owner’s manual

indicates it is accurate to within 10 m. MDA recorded the GPS measurements in the field; Red Eagle

Mining’s GIS staff also converted the latitude and longitude easting and northing readings downloaded

from the GPS unit to local coordinate space. Elevation readings did not require conversion to the local

system.

The collar coordinates downloaded from the GPS and converted to local space from latitude and longitude

readings were compared to certified coordinates in Red Eagle Mining’s database provided by the

surveyor. MDA’s independent collar measurements are intended only as a check of reasonableness

against the collar coordinates in Red Eagle Mining’s database. Given the level of accuracy of the GPS

used by MDA, all northings and eastings measured by MDA when compared to Red Eagle Mining’s

database are well within acceptable limits. Only two eastings differed by more than 5m but did not exceed

7.1 m. All measured elevations were consistently high by 9.5 m to 17 m relative to surveyed elevations (in

steep terrain, hand-held GPS elevation readings are generally much less accurate than northing and

easting readings). Also, the eastings recorded in the field directly from the GPS unit were consistently

higher than the converted eastings (which also match the surveyor’s eastings more closely) by 3 m to

4.5 m. The GPS checks confirmed the location of the drill-hole collars in Red Eagle Mining’s database

within reasonable limits.

Page 12.22




AMENDED NI 43-101 TECHNICAL REPORT 12.4 MDA Independent Verification of Mineralization

During the initial site visit, MDA collected 15 samples to confirm the presence of gold at the San Ramon

deposit. These samples are not intended to be duplicate samples for grade verification. Check samples

were selected to provide a range of sample grades from mineralized zones; original sample grades range

from 0.5 to 9 g Au/t.

Eight quarter-core samples of sulfide mineralization were taken from four core holes. These were saw cut

by Red Eagle Mining technicians. Since all mineralized core from the oxide program was being consumed

for assay, the only oxide material available for check sampling was coarse rejects from the current drill

program. Two were selected, and since there was no riffle splitter available, these were split by MDA

using the quartering method on a plastic bag. The split was far from perfect but was expected to return

results that were broadly similar to the original. Four channel samples were taken from the Bernadina (1),

Guaguas (1), and San Francisco (2) adits, and one was taken from a road cut on a switchback at the east

end of the San Ramon shear zone; these samples were from oxide mineralization. Confirmation sampling

results are presented in Table 12.10.

Samples were collected by MDA, placed into polyethylene bags, and zip-tied shut. Samples were kept in

the core-logging building and placed into rice bags, which were also zip-tied shut. MDA travelled with the

two sealed rice bags as they were transported to Red Eagle Mining’s Medellín office but was unable to

follow the samples from that point. Red Eagle Mining took possession of the samples and shipped them

to MDA’s office in Reno via DHL Express. Upon arrival, the rice bags had been opened for inspection by

DHL, but the samples remained zip-tied and appeared intact and uncompromised.

The results of the independent check sampling confirm the presence of mineralization in the samples.

The grades are roughly similar to original assays in Red Eagle Mining’s database, but, as expected, there

is some variability. The differences in grades between original and check assays could be an indicator of

the heterogeneity of the gold at the San Ramon deposit.

Page 12.23




AMENDED NI 43-101 TECHNICAL REPORT Table 12.10 MDA Confirmation Samples from Core, Outcrop, and Adits

12.5 Summary Statement on Data Verification

The QA/QC programs implemented by Red Eagle Mining for all drilling are generally thorough but do have

deficiencies. Despite the short-comings, which are deemed minor and are described elsewhere, the

overall results of the QA/QC programs indicate that the assays are of sufficient quality for use in the

resource estimate. Furthermore, all of MDA’s verification work showed that Red Eagle Mining’s sample

data and database are suitable for use in a Measured/Indicated/Inferred-classified resource estimate.

Location Sample Type Redox

Zone

Red Eagle (Acme) MDA (Chemex)

Sample Number Au (ppm) Ag (ppm) Sample Number Au (ppm) Ag (ppm)

SR-015 1/4 core Sulfide SR-06784 4.607 7.8 MDA-SR-001 13.85 21.9








SR-073 1/2 split of coarse rejects Oxide SR-04579 0.591

MDA-SR-009 0.789 2.2

SR-080 1/2 split of coarse rejects Oxide SR-04901 3.147

MDA-SR-010 2.95 3.9

Bernadina Adit Channel Oxide 2100-RS-AC- 1.082

MDA-RC-001 0.983 7.1

East Switchback

Roadcut Channel Oxide 2100-RS-TN-057 1.455 1.3 MDA-RC-002 1.235 0.6

Guaguas Adit Channel Oxide 2100-RS-AC-003 0.638

MDA-RC-003 0.334 0.25

San Francisco Adit Channel Oxide 2100-RS-VP-208 1.470

MDA-RC-004 1.465 1.1


MDA-RC-004 1.465 1.1


MDA-RC-005 31.6 13.2


MDA-RC-005 31.6 13.2

Page 13.1




AMENDED NI 43-101 TECHNICAL REPORT 13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

Most of the metallurgical test work to-date has been completed at two laboratories: Kappes, Cassiday &

Associates (KCA), and McClelland Laboratories, Inc. (MLI). A limited preliminary (sighter) testing program

was undertaken by Acme Metallurgical Limited. As identified throughout the following text, several other

laboratories have performed specialized tests, mainly as subcontractors to MLI or KCA.

These various programs are discussed in chronological order in the following subsections. There are a

few exceptions to this organization. One is a separate subsection that covers the various comminution

studies as a single unit. This follows the subsections on MLI’s work. Standalone sections on the ore and

concentrate mineralogy and the metallurgical QA/QC programs follow the discussion of the comminution

parameters. The section concludes with a summary of the metallurgical results, including risks and

opportunities.

Note: The term “ore” is used in this section with regard to metallurgy and is not intended to imply that a

particular sample could be treated economically.

13.1 Acme Metallurgical Limited Test Work

Initial metallurgical test work on the San Ramon deposit mineralization was conducted by Acme

Metallurgical Limited (“AcmeMet”) in Vancouver, British Columbia between April and September, 2012.

Testing was conducted on four bulk samples that included both oxide and sulfide mineralization from the

San Ramon deposit shear zone (Kwok and Choi, 2012). At this stage of early metallurgical and mineral

resource considerations, oxide and sulfide metallurgical processes were investigated (“sighter’ tests),

including heap leach and CIP/CIL leaching of the potential oxide and sulphide mineral resources. The

following summarizes that testing and is taken from AcmeMet’s report.

Table 13.1 summarizes the head grades for the four samples. The sample identified as “Primary” came

from sulfide and shear-zone mineralization from the underground Hilo Azul workings. The sample

identified as “Sulfide” came from a quartz vein in sulfide and shear-zone mineralization, also from the Hilo

Azul workings. The oxide samples came from existing adits and a roadside shear zone outcrop.

Table 13.1 Head Grades for Metallurgical Samples from San Ramon Shear Zone

(Kwok and Choi, 2012)

Sample ID Au

g/t

Ag

g/t

S

%

As

ppm

Cu

ppm

Pb

ppm

Fe

%

Primary 0.90 2 0.48 74 55 262 4.48

Oxide #1 7.48 4 0.10 5134 48 184 6.21

Oxide #2 2.43 <2 0.03 1299 74 223 4.30

Sulfide 24.3 38 6.96 316 229 4679 6.68

Page 13.2




AMENDED NI 43-101 TECHNICAL REPORT Preliminary cyanide leaching tests, as simulated by 144-hour bottle-roll tests, showed that the two oxide

samples are amenable to cyanide leaching at 0.50 inch size, and the sample identified as “Sulfide” is

amenable to cyanide leaching at a 75-micron grind size. However, at the coarse 0.50 inch size, the

Sulfide sample was not amenable to leaching. Test results are summarized in Table 13.2. In most cases

gold extraction was continuing when the bottle roll tests were terminated. Thus, the reported figures may

not represent ultimate extractions.

Table 13.2 Summary of Preliminary Cyanide Leaching Tests from San Ramon Samples

(Modified from Kwok and Choi, 2012)

Sample ID Top Size

inch/um

Au Extraction @

144 Hours (%)

Ca(OH)2 Addition

kg/t

NaCN Consumption

kg/t

Primary 0.50 inch 43.4 2.4 0.6

Oxide #1 0.50 inch 85.3 11.7 0.6

Oxide #2 0.50 inch 72.8 10.4 0.4

Sulfide 75* micron 94.6** 3.3 0.4

* 80% passing 75 microns ** Based on 48 hours

As can be seen, the two types of mineralization exhibit the same low cyanide consumptions, but the

oxides have much higher lime requirements than the sulfides.

Throughout the AcmeMet program, reagent consumptions were reasonable. Consumption of sodium

cyanide in the ore leach tests was generally between 0.4 and 0.6 kg/t ore, with a few values slightly lower

or higher. Lime consumption was more variable; it ranged from 1.8 to 3.3 kg/t for the sulfides and about

9 to 12 kg/t for the oxides. Cyanide and lime consumption in the concentrate leach tests were both in the

range of 4 to 6 kg/t.

13.2 Kappes, Cassiday & Associates Test Work

Kappes, Cassiday & Associates (“KCA”) of Reno, Nevada, conducted metallurgical test work on a range

of mineralization types from the San Ramon deposit. This section summarizes the results of that testing,

which commenced in January 2013 and was completed as of August 1, 2013. KCA did not issue a formal

report on their test work, but described test parameters and provided results in the form of log sheets and

spreadsheet compilations. Information presented in this report is drawn from this raw data.

Two main parallel processing routes were tested, whole-ore leaching, and flotation and concentrate

leaching

whole-ore leaching involved crushing and grinding all ore-grade material, followed by leaching

the entire sample with cyanide in an agitated cyanide circuit; and

Page 13.3




AMENDED NI 43-101 TECHNICAL REPORT flotation and concentrate leaching involved primary grinding, followed by a rougher flotation step

on the ore. Then the concentrate was reground and leached to recover the gold.

The main difference between the two alternatives is the amount of material that is in contact with cyanide

and must be detoxified. In the whole-ore leach option, 100% of the ore will have to be detoxified and

stored. With the latter approach, most of the material will be uncontaminated flotation tailings, which can

be mixed with the mine waste and dry stacked in a single location. A separate process evaluation was

performed to determine which route is the more attractive.

In addition to the two leach routes that were tested, selected samples were subjected to gravity

concentration on the mill discharge.

13.2.1 Sample Description

KCA’s program was a major expansion of the four-sample preliminary program completed in 2012 by

AcmeMet and described in Section 13.1. The bulk of the KCA testing focused on composites of the oxide,

transition and sulfide mineralization types. The composites are believed to be representative of each

specific type of mineralization as all material was selected by the Red Eagle Mining project geologist on

the basis of head grade, mineralization, and host rock. However, a few individual samples that met these

criteria actually lie in stringers outside the main mineralized trend.

The oxide composite was made up of assay reject material from 118 intervals (nominal 1 m in

length) drawn from 21 drill holes along the mineralized trend.

The sulfide composite was prepared by cutting drill core into quarters, and pulling approximately

500 grams of material from each selected interval in the core trays. In all, the sulfide composite

contained material from 202 intervals in 24 drill holes. As the sulfide mineralization made up

about 90 % of the known resource at the time of the study, sulfide variability samples were also

tested. There were six composites included in this part of the program, two from the western

end of the mineralized shear zone and four from the eastern end. Each composite was made up

of core intervals taken from two to six closely spaced drill holes. In all, the six samples included

material from 204 intervals in 25 drill holes.

The transition mineralization is a minor type (< 5% of the known resource). This composite was prepared

by taking material from 12 channel cuts in the San Francisco adit.

In addition to the nine composites described above, two individual samples were included in the test work

program:

a high-grade sulfide sample from the Hilo Azul adit; and

a low-grade saprolite ore sample taken from a surface channel.

Page 13.4




AMENDED NI 43-101 TECHNICAL REPORT The various composites used in the KCA test work are generally representative of the major lithological

groups (oxide/saprolite, saprock/transition, sulfide, quartz vein), although spatial representivity varies.

Sulfide mineralization was well represented; spatially, a majority of the samples for the sulfide and variable

sulfide composites were from the eastern half of the deposit, which comprises the bulk (+80%) of the

resource, but a representative number were included from the western part. Samples were taken at a

range of depths to ~350 m below the surface. Samples for the oxide composite were evenly distributed

across the eastern half of the deposit. Access to and availability of transition samples was limited to San

Francisco adit at the west end of the deposit, although, as noted, the material makes up a very small

portion of the deposit. Some of the samples that were collected for the oxide, sulfide, and E1, E2, E4, and

W1 variable sulfide composites (Table 13.3) were taken from mineralized core intercepts that occur in the

hanging-wall and footwall of the primary shear zone. These samples were taken from secondary fault

structures, but generally have the same geologic characteristics as the primary shear zone.

With the exception of the Hilo Azul composite (47.19 g Au/t), the head grades for all composites ranged

from 0.66 to 2.30 g Au/t. These represent potential open-pit grades. The composite grades cover a range

above and below a potential underground mining cut-off grade 2 g Au/t), and further testing on higher-

grade samples is required to represent average underground mineable grades.

13.2.2 Whole-ore Leaching

A total of 37 leach tests was completed, 30 on whole mineralised rock and seven on rougher

concentrates, comprising 16 tests on sulfide material, 10 on oxide samples, eight on transition material,

two on saprolite, and one on Hilo Azul shaft high-grade material.

In the initial series of tests, grind size was a major variable. Tests were conducted on samples where the

P80 varied from 200 microns down to 40 microns. To obtain maximum extraction, the sulfide samples

required a somewhat finer grind than the oxide or transition samples. For direct leaching of the combined

mineralization types, a P80 grind size of 75 microns (200 mesh) appears to be the optimum, as it provides

a good balance between grinding costs, gold recovery, and leach cycle time.

All types of mineralization, including the sulfides, responded well when leached directly. Except for the

transition samples and one variability sample, gold extraction exceeded 90 % in all tests. These results

are summarized in Table 13.3. In the tests listed, the samples were rod milled to a nominal grind size

(P80) of 75 microns. Bottle roll leaching was performed at a cyanide concentration of 1 g/L using sodium

cyanide (NaCN) as the lixiviant. These are all the tests run using the parameters selected for the PEA. A

96-hour leach cycle was used in all tests, although the samples typically reached their maximum

extraction levels in about 24 hours. The rougher concentrate leach tests are not included in this Table and

are discussed as part of the flotation program.

Page 13.5




AMENDED NI 43-101 TECHNICAL REPORT Table 13.3 Results of Direct Leaching on All Ore Types and Variability Samples

Sample Type

Calculated Head

Grade,

g Au/t

96-h Gold

Extraction, %

Reagent Usage

NaCN, kg/t Ca(OH)2, kg/t

Oxide Composite 0.840 95 0.30 3.0

Transition

Composite 1.561 87 1.96 3.0

Sulfide Composite 1.489 92 1.20 1.5

Hilo Azul Sulfide 47.19 96 1.93 1.0

Saprolite 0.554 94 0.22 3.5

E1 Variable Sulfide 1.747 96 4.95 1.0




W1 Variable Sulfide 0.858 86 1.99 1.75

W2 Variable Sulfide 2.295 93 1.56 1.00

Reagent consumption was variable and was determined at the end of the 96-hour leach tests. Most

samples required less than 2 kg of NaCN per tonne (kg/t), with a range from 0.2 to 5 kg/t. The sulfide

variability samples showed something of a geographic trend, with cyanide consumption declining from

5 kg/t at the far east end to 1.5 kg/t at the far west end. Lime consumption was generally low with most

sulfide samples only requiring 1 kg/t. The non-sulfide samples required 3.0 to 3.5 kg/t. This may be due

to the higher calcite levels in the unweathered sulfides.

Due to the significant variability in the NaCN consumption, seven samples were sent to McClelland

Laboratories, Sparks, Nevada for confirmatory QA/QC tests replicating the KCA test conditions (see

Section 13.2.7 below).

13.2.3 Flotation with Concentrate Regrinding and Leaching

Flotation

KCA performed 40 rougher flotation tests, including 17 on various sulfide samples, 11 on the oxide

composite, 10 on the transition composite and one each on the Hilo Azul and saprolite samples.

Initial testing focused on two areas. One was identification of a reagent suite that would provide effective

recovery of gold while minimizing the mass pull required. The other was a grind size study to determine

grinding requirements needed to achieve good gold recovery.

The reagent selection study showed that all ore types could be floated effectively at their natural pH levels.

A combination of copper sulfate at 250 g/t, PAX at 125 g/t and 208 at 5 g/t, together with about 25 g/t of

the frother F-579, was found to give good gold recovery at mass pulls of 5 to 10 % of the ore feed.

Page 13.6




AMENDED NI 43-101 TECHNICAL REPORT The grind size study showed that grinding the sulfide ores to a P80 finer than 150 microns provided no

observable improvement in gold recovery. However, both the oxide and transition composites showed

modest increases in gold recovery when ground to a P80 of 75 microns. Since the combined oxide and

transition ore types represent less than seven percent of the known resource, the coarser grind appears to

be the optimum for the San Ramon deposit.

A summary of optimum rougher flotation results at nominal grinds of both 150 and 75 microns is provided

in Table 13.4. As might be expected, the sulfides responded to flotation better than the other types of

mineralization. Silver recovery is also shown, although this would not be a major contributor to project

revenue. As can be seen, silver recoveries were generally lower than the corresponding gold recoveries.

Table 13.4 Summary of Optimum Rougher Flotation Results

Sample Type Actual P80 Grind Size,

µm

Mass Pull, % of

Feed

Percent Recovery

Gold Silver

Oxide Composite 163 10.0 60 64

Oxide Composite 82 6.0 76 64

Transition Composite 129 7.9 82 73

Transition Composite 92 9.4 90 78

Hilo Azul 148 32.8 98 96

Saprolite 79 6.9 58 39

Sulfide Composite 153 6.8 98 87

Sulfide Composite 78 8.9 97 88

E1 Sulfide Composite 138 4.5 95 87




W1 Sulfide Composite 200 5.7 88* 86

W2 Sulfide Composite 119 5.2 98 85

* May have been adversely impacted by coarse grind size (200 µm rather than 150 µm).

Concentrate Regrinding and Leaching

Rougher flotation concentrates from all three mineralization types have been subjected to cyanidation for

final gold recovery.

Leach tests were limited to just one transition sample and two oxide samples. These three samples were

floated at a P80 grind size of about 75 µm using the optimum reagent suite. Leaching of these rougher

concentrates was done at a cyanide concentration of 1 g/L NaCN without prior regrinding. Results are

summarized in Table 13.5.

Page 13.7




AMENDED NI 43-101 TECHNICAL REPORT Six leach tests were run on sulfide rougher concentrates. Two were leached without regrinding. Two

were leached at 1 g/L NaCN after regrinding, and two were leached at 5 g/L NaCN after regrinding.

These results are also included in the table.

It is important to note that the reagent consumptions shown in the table are stated on a per-tonne of

concentrate basis, not a per-tonne of ore basis. At a 10 % mass pull in flotation, the reagent

consumptions shown in the table should be divided by 10 to give the per-tonne of ore values. Reagent

consumptions are those measured after 96 hours at the end of the tests and were lower at shorter leach

times.

Table 13.5 Summary of Concentrate Leach Tests

Sample Type Calculated Head Grade,

g Au/t

Cyanide

Concentration, g/L NaCN

Gold Leach

Extraction %

Reagent Consumption kg/t

NaCN Ca(OH)2

Oxide Composite 15.918 1.0 95 6.15 3.97

Oxide Composite 16.023 1.0 81 4.85 6.69

Transition Composite 18.227 1.0 92 3.32 2.33

Sulfide Composite1 20.902 1.0 59 4.26 1.18






1 No regrind

2 30 minute regrind

3 60 minute regrind

4 120 minute regrind

As can be seen from Table 13.5 the concentrate leach results were erratic. The two oxide trials were

essentially replicates run on rougher concentrates produced from splits taken from the same composite.

Thus, there is no obvious reason for the disparity in the gold extraction and reagent consumption. Unlike

the oxide and transition concentrates, which do not appear to require regrinding, the sulfide concentrates

leached poorly, unless reground. However, the sample with a short regrind and low cyanide concentration

leached as well or better than concentrates ground longer or leached at a higher cyanide level. The

sample with the two-hour regrind essentially represents ultra-fine grinding. While this test yielded the

highest extraction, the cyanide consumption was also far higher than it was in any other test. In

conclusion, when flotation-concentrate leaching is considered in the future, then further concentrate leach

tests will be required to clarify leach performance and establish the optimum leach parameters.

Page 13.8




AMENDED NI 43-101 TECHNICAL REPORT 13.2.4 Gravity Concentration

A limited program of gravity concentration was conducted on three sulfide composites. One was the

overall sulfide composite used in many of the test work trials. One was a master composite prepared from

the four eastern end of the deposit variability samples, and one was a master composite formed by

combining the two western end variability composites. All three composites were tested using the same

protocol. A nominal 1.0 kg sample was fed to the gravity separation unit. The initial gravity concentrate

was then hand panned to simulate final concentration of the gold by tabling. The tailings from the two

gravity concentration steps were combined and then floated to recover a portion of the remaining gold.

Flotation was done using the optimum conditions previously determined. The resulting rougher

concentrates were then reground for 30 minutes. These were leached at 5 g/L NaCN using a 96-hour

leach cycle. There were no tests that involved direct leaching of the gravity tails.

Gold recoveries for each unit operation, along with the overall recoveries, are summarized in Table 13.6.

The final gravity concentrate represented about 0.5 % of the feed with an average grade of about 100 g

Au/t and 230 g Ag/t. However, the silver recoveries were lower than the gold recoveries when expressed

as a percentage. Direct leaching achieved higher recoveries than the gravity-flotation-leach route.

Table 13.6 Summary of Gold Recovery from Sulfides with Gravity Concentration

Sample

Total Gold

Recovery in

Gravity, %

Total Gold

Recovery in Float-

Leach, %

Overall Gold

Recovery with

Gravity, %

Gold Recovery in

Direct Leach, %

East Composite 49 41 91 95.1

West Composite 27 45 72 90.1

Sulfide Composite 27 61 89 91.2

1 Average of all direct leach tests

2 Average of comparable tests at 5 g/L NaCN

13.2.5 McClelland Laboratories, Inc. Confirmatory Test Work

Because of some inconsistent results on reagent consumption at KCA, an initial short, confirmatory, test

work program was undertaken at McClelland Laboratories, Inc. (“MLI”) of Sparks, Nevada. The work was

limited to direct leaching of whole ore using splits from five sulfide composites previously tested at KCA.

The samples were delivered to MLI in August, 2013, and the work was complete in October, 2013. The

following information is taken from various spreadsheets and other sources of raw data, as MLI has never

issued a final report on this portion of the work.

Page 13.9




AMENDED NI 43-101 TECHNICAL REPORT The samples tested included the variability composites E2, E3, E4, and W2, plus the main sulfide

composite. Duplicate tests were performed on the latter to gauge experimental reproducibility. Single

tests were performed on the others due to the limited amount of material available. The MLI bottle-roll

tests replicated the KCA protocol as closely as possible. This included rod-mill grinding of the feed to a

P80 of 75 µm and a 96-hour leach at a nominal cyanide concentration of 1 g/L NaCN and a pH range of

10.5 to 11.0. The only discernible difference in the two sets of tests was in the dissolved oxygen (“DO”)

levels. In the MLI tests, the DO level averaged just over 7 ppm. In the KCA runs, the DO level was

generally below 5 ppm and dropped below 4 ppm on several occasions, necessitating oxygen sparging.

Extraction was very fast in all tests, with leaching essentially complete in just 24 h. An example of the

leach curves for the main sulfide composite is provided in Figure 13.1. This shows the rapid completion of

the leach extraction of gold. The very close agreement for the duplicate runs also demonstrates that the

MLI testing protocol is highly reproducible.

Figure 13.1 Gold Leach Rate Profiles, Bottle Roll Test

The comparable gold extractions obtained by MLI and KCA are shown in Table 13.7 while the cyanide

consumptions are compared in Table 13.8. As can be seen, the gold extractions are nearly identical and

average 93.6 %. However, the cyanide consumptions are markedly different. The average value for both

sets of data is 1.47 kg of NaCN. This value was used until additional confirmatory tests resolved the

discrepancy.

0

10

20

30

40

50

60

70

80

90

100

0 24 48 72 96 120

Cu

mu

lati

ve R

eco

very

, % o

f T

ota

l

Leach Time, hours

San Ramon Master Sulfide 66809 Drill Core Composite,

80%-75µm Feed Size

Au (Initial) Au (Duplicate)

Page 13.10




AMENDED NI 43-101 TECHNICAL REPORT Table 13.7 Comparable Gold Extractions in the MLI and KCA Tests

Composite / Test Gold Recovery in Direct Leach, %

MLI KCA Average

Master Sulfide Composite 66809 CY1 91.5 92 91.7

Master Sulfide Composite 66809 CY2 92.1 No Test - - -

E2 Sulfide 66850 CY3 94.2 93 93.6

E3 Sulfide 66851 CY4 95.7 96 95.8

E4 Sulfide 66852 CY5 96.1 94 95.0

W2 Sulfide 66854 CY6 93.4 93 93.2

Average All Tests 93.8 93.6 93.6

Table 13.8 Comparable Cyanide Consumptions in the MLI and KCA Tests

Composite / Test Cyanide Consumption, kg/t NaCN

MLI KCA Average

Master Sulfide Composite 66809 CY1 0.32 1.20 0.76

Master Sulfide Composite 66809 CY2 0.46 No Test - - -

E2 Sulfide 66850 CY3 0.66 5.01 2.84

E3 Sulfide 66851 CY4 0.36 3.92 2.14

E4 Sulfide 66852 CY5 0.45 1.95 1.2

W2 Sulfide 66854 CY6 0.48 1.56 1.02

Average All Tests 0.46 2.47 1.47

In addition to determining the overall gold extraction in each test, MLI also conducted tail-screen analyses

on each final leach residue, with gold assays by size fraction. Results are shown in Table 13.9.

Unfortunately, there was insufficient material to run comparable head-screen analyses to determine

extraction by screen fraction. Nonetheless, the results show that the residue assays decline progressively

as the particle size gets finer. Thus, it is likely that extraction increased with decreasing particle size.

Therefore, further testing was undertaken to confirm the optimum grind size for direct leaching of the

sulfides.

Table 13.9 MLI Gold Assays by Size Fraction in the Leach Residues

Size

Fraction, µm

Tailing Assays by Screen Fraction, g/t Au

66809 CY-1 66809 CY-2 E2 CY-3 E3 CY-4 E4 CY-5 W2 CY-6

+106 µm 0.230 0.216 0.174 0.213 0.062 0.325

-106+90 µm 0.210 0.185 0.138 0.189 0.060 0.290

-90+75 µm 0.214 0.205 0.211 0.152 0.062 0.213

-75+63 µm 0.181 0.186 0.145 0.204 0.035 0.227

Page 13.11




AMENDED NI 43-101 TECHNICAL REPORT -63+53 µm 0.167 0.161 0.131 0.162 0.048 0.238

-53+45 µm 0.162 0.152 0.114 0.160 0.048 0.199

-45+37 µm 0.145 0.148 0.123 0.118 0.048 0.212

-37 µm 0.047 0.046 0.041 0.047 0.013 0.097

13.2.6 McClelland Laboratories, Inc. Test Work on High Grade Ore

Much of the test work done at KCA and the confirmation tests at MLI was conducted on samples with low

gold grades typical of the open pit operation that was originally anticipated for this project. Subsequent

development indicated than an underground mining operation would be more appropriate and six higher

grade composites (Comp 1 through 6) were prepared to represent the range of expected ore grades.

Work on these new samples started in October 2013 and is approaching final conclusion to a very

successful program.

MLI has completed several programs to date:

Full composite characterization;

A limited diagnostic test to determine the effects of dissolved oxygen (DO) levels on gold

extraction and reagent consumption;

Whole ore leach tests as a function of grind size;

Two types of gravity concentration, extended gravity recoverable gold (EGRG) and

gravity/cyanidation tests; and

A combined approach, in which flotation concentrate was reground, then recombined with the

flotation tails for leaching of the combined product.

Work still in progress includes optimizing the regrind size, cyanide neutralization (detox) tests, liquid-solid

separation tests and carbon capacity tests. As the work is ongoing, MLI has yet to issue a final report on

their overall program. Various portions of the program are discussed in the following subsections. The

discussion is based on interim results issued in spreadsheets issued as the work has progressed. These

programs shall continue to completion and will be used during the next detailed engineering and specific

metallurgical testwork stages.

Page 13.12




AMENDED NI 43-101 TECHNICAL REPORT High Grade Composite Preparation

MLI’s most recent program has focused on testing composites that are representative of the higher grades

in the underground-mineable resource. Material from 88 intervals in 52 core holes was combined to

produce six composite samples used for testing. Coarse assay laboratory rejects made up the majority of

the composited materials, although some intact core was also used. All composite head grades exceeded

the potential underground cut-off grade of 2.0 g Au/t. Figure 13.2 shows the location of all drill intervals

with grades exceeding 2.0 g Au/t that were used in the metallurgical composites. These demonstrate the

wide spatial distribution of the samples, particularly in the eastern and central portions of the mineralized

trend that account for the bulk of the minable resource. The figure also includes the 2012 and 2013

samples that are still at or above the current cut-off grade.

Page 13.13





Figure 13.2 The Location of All Drill Intervals with Grades >2 g Au/t Used in the Metallurgical Composites

Page 13.14




AMENDED NI 43-101 TECHNICAL REPORT All of the 2014 metallurgical samples are also unoxidized and unweathered, and are hosted almost

entirely in sheared granodiorite. All but 5 of the 88 intervals are from within the main shear zone.

Therefore, with only a few exceptions, the samples that make up the six composites represent the primary

geologic characteristics of the bulk of the mineralization to be mined. Each composite consists of intervals

that are broadly scattered throughout all but the western quarter of the deposit. Therefore, in total, the

central and eastern portions of mineralized material are spatially well represented in each composite.

Since each of the six composites is composed of material from throughout the eastern and central

portions of the deposit, the 2014 samples will not demonstrate any local variability in metallurgical

characteristics. The only variability amongst these six composites is in the head grade, which ranges from

2.7 to 9.6 g Au/t which were selected specifically to cover the anticipated overall ROM grade range that

could be delivered from underground mining. Since the whole ore leach tests on the KCA variability

samples did exhibit a decreasing trend in both recovery and reagent consumption going from east to west,

it may be useful to conduct variability testing on the western area in advance of mining.

Sample Characterization

Chemical characterization of the six composites included head assays (in triplicate) for gold and silver,

multi-element ICP scans, and classical whole-rock analyses. Key assays are summarized in Table 13.10.

These assays represent the lower and upper grade ranges the were expected to be encountered, typically

during mining operations,

Table 13.10 Partial Chemical Composition of the Six Test work Underground Composites

Composite

Chemical Analyses

Au, g/t Ag, g/t As, g/t Cu, g/t Fe, % Pb, g/t Zn, g/t Sulfide S,

%

Comp. 1 2.72 5 92.4 72.3 3.92 415 1,020 1.13

Comp. 2 3.26 7 552 88.7 3.98 456 1,890 1.84

Comp. 3 4.34 10 244 130.0 4.87 729 2,280 2.89

Comp. 4 5.36 10 903 109.5 5.05 664 2,700 3.24

Comp. 5 9.60 17 659 108.0 4.94 1,185 3,640 3.02

Comp. 6 5.29 7 160 99.9 5.0 397 1.880 2.01

During the exploration program, silver assays were not always included, as the main focus was on the

gold. As shown in Figure 13.3 , there is a linear relationship between the gold and silver head assays for

the six composites. This relationship can be used to project the silver grade for any ore sample where the

gold grade is known. For example, at the life-of-mine (LOM) gold grade of 4.57 g Au/t the estimated LOM

silver grade would be 8.48g Ag/t making the ratio of silver to gold 1.85:1.

Page 13.15




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.3 The Gold-Silver Head Grade Relationship

The whole rock analyses show that the makeup of all composites is remarkably similar. Silicon (as SiO2)

makes up 60 to 62 % of the material, probably mostly as quartz, along with feldspar and mica. This is

followed by aluminum (as Al2O3) at 10 to 13 %, mostly in the mica and feldspar. Iron (as Fe2O3) makes

up 6 to 8 % of the composites, mostly as pyrite and pyrrhotite; calcium (as CaO) accounts for about 6 %,

likely there as calcite. Loss on ignition, probably the carbonate in the calcite, also makes up 6 to7 % of

the material.

Dissolved Oxygen (DO) Tests

In both the KCA program and the MLI confirmatory tests, there was some indication that the DO levels

during the direct leach tests were influencing the recovery and reagent consumption. Therefore a short

diagnostic study was undertaken before embarking on the direct leach program with the grade

composites. The sample used in the diagnostic tests was the 66809 (all-in) master composite obtained

from KCA. Four tests were conducted: a closed bottle roll run with no sparging and three mechanically

agitated tests in an open vessel, one with no sparge, one with an air sparge and one with an oxygen

sparge.

y = 1.6119x + 1.1155 R² = 0.8723

0

2

4

6

8

10

12

14

16

18

20

0 2 4 6 8 10 12

Silv

er

He

ad G

rad

e, g

Ag/

t

Gold head Grade, g Au/t

Page 13.16




AMENDED NI 43-101 TECHNICAL REPORT All four tests followed the same protocol. This included a nominal 1-kg ore charge with a P80 of 75 µm.

This was slurried in a nominal 1.5 L of solution to give 40 % solids. After determining the natural pH of the

slurry, 1.5 g of NaCN was added to obtain a concentration of 1.0 g/L. Sufficient hydrated lime was added

to raise the pH to 11.0. Tests were run for a total of 72 hours, with intermediate sampling times of 2, 4, 8,

16, 24, 48, and 72 hours. At each sampling interval a nominal 100 mL sample of solution was withdrawn

and analyzed for NaCN, pH, DO, and gold. If necessary, NaCN or lime was added to bring the

concentration or pH back to the set point. At the end of the test, the slurry was filtered, the filtrate being

the pregnant leach solution (PLS). This was assayed for gold and thiocyanate and also analyzed using an

ICP multi-element scan.

Key results of the four tests are shown in Table 13.11. As can be seen, increasing the DO level had no

discernible effect on gold recovery but did appear to reduce lime consumption. However, cyanide

consumption increased as the available oxygen increased. Based on these tests it appeared that no

sparging was necessary and that agitation alone pulled sufficient air into the slurry to maintain an

adequate DO level.

Table 13.11 Summary of Results of the Dissolved Oxygen Tests

Test Description Dissolved O2 Level

ppm

72-Hr. Gold

Recovery %

NaCN

Consumption kg/t

Lime

Consumption kg/t

Closed Bottle Roll 4.2 93.4 0.27 1.9

Mech. Agitation (no sparge) 3.3 91.4 0.59 2.1

Mech. Agitation (air sparge) 5.2 93.2 1.18 1.6

Mech. Agitation (O2 sparge) 28.2 92.3 1.62 1.3

Whole Ore Leach Tests

The initial program with the underground ore was a continuation of the whole ore leach processing route

undertaken in the confirmatory program. The difference was that the samples used were the underground

composites listed in Table 13.10. The principal process variable in these tests was the crush size, as

determined by the P80 value. Five values (150 µm, 106 µm, 75 µm, 53 µm and 37 µm) were tested for

each composite, except Comp 6, which was only tested at 75 µm. Grind size studies were conducted on

each composite in order to determine the grind times needed to achieve the various P80 values.

Page 13.17




AMENDED NI 43-101 TECHNICAL REPORT In all, 30 leach tests were run on whole ore samples. The total included three repeat tests and a single

diagnostic test intended to test the benefits of an alkaline aeration pre-treatment. All tests were run using

the same protocol. A nominal 0.5 kg sample (exact weight recorded) was combined with 750 mL of water

(exact volume recorded) to give a slurry with 40 % solids. The natural pH of the slurry was determined.

This ranged from 7.3 to 8.2 for Comps 1 through 5, with most values either 7.8 or 7.9. The natural pH of

Comp 6 was slightly higher at 9.0. At the start of the test sufficient NaCN was added to give a cyanide

concentration of 1.0 g/L NaCN and sufficient lime (as hydrated lime) was added to raise the pH to about

11. Then the slurry was agitated using a bottle roll technique. Agitation was interrupted after 2, 4, 8, 16,

24, and 48 hours to collect intermittent solution samples. The cyanide concentration and pH were also

checked and reagents were added if necessary to re-establish the desired set points. The test was

terminated after 72 hours and the slurry was filtered to obtain the final PLS. The filter cake was double

rinsed. The rinse volumes were measured and any gold and silver in the rinse solutions were included in

the overall mass balance. Once each filter cake was dried, it was weighed and then subjected to a

standard screen size analysis with size fractions ranging from -37 µm to +150 µm. Each of the seven

screen fractions was assayed for both gold and silver.

As shown in Figure 13.4, the tests demonstrated that there is a strong inverse relationship between grind

size and gold extraction. In the figure, each point is the average of the tests run on all composites at the

given particle size. At a P80 of 3 µm, recoveries were in the range of 95 % and were generally

independent of head grade. The corresponding silver recovery averaged about 73 % and was also

independent of head grade.

In spite of the simplicity of the whole ore leach flowsheet and the high recoveries, this approach to ore

treatment has some disadvantages. One is that the entire mine output must be ground to a very fine size,

requiring a multi-stage crushing and grinding circuit and a high power consumption. With a P80 of 37µm,

about 80 % of the material will be sub-sieve, leading to possible difficulties during thickening and filtration.

In addition, the leach kinetics are slow, even at the finest grind. In every case extraction was continuing

when the tests were terminated after 72 hours. With such slow kinetics, agitated leach circuit would

require large tanks and agitators. Finally, the long leach cycle would lead to high cyanide consumption, as

this increased continuously throughout the leach due to ongoing cyanate formation, which will occur after

the gold has been dissolved, but the cyanide ion (HCN-) continues to oxidise to form various cyanates.

Page 13.18




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.4 The Relationship between Grind Size and Gold Recovery in a 72-Hr. Direct Leach

In addition to the foregoing problems, parallel mineralogical examination of the San Ramon samples

showed that the ore contained pyrrhotite and other sulfides that are potential oxygen consumers.

Therefore, one test was conducted with an alkaline aeration pre-treatment, which should oxidize these

sulfides, reducing cyanide consumption and possibly increasing recovery. The results were disappointing.

The cyanide consumption did drop by about 15 %, to about 0.82 kg/t ore. However, gold extraction

dropped slightly and lime consumption almost tripled. Therefore, no further consideration was given to

use of an alkaline aeration pre-treatment.

Gravity Concentration

As discussed in Section 13.2.8, the various mineralogical studies clearly demonstrated that the San

Ramon deposit contains some coarse gold with individual particles having at least one dimension in the

range of 50 to up to over 200 µm in size. In addition, tail screen analyses on the direct leach residues

generally exhibited very high gold grades in the coarsest fractions. A typical example is direct leach test

CY-23, which involved Composite 4 ground to a P80 of 53 µm. Key results are shown in Table 13.12.

y = -0.0704x + 97.186 R² = 0.9977

86

87

88

89

90

91

92

93

94

95

96

0 20 40 60 80 100 120 140 160

Go

ld R

eco

very

, %

P80 Grind Size, µm

Page 13.19




AMENDED NI 43-101 TECHNICAL REPORT Table 13.12 Tail Screen Analysis for Direct Leach Test CY-23 (Composite 4 Ground to a

Nominal P80 of 53 µm)

Size Faction, µm

Weight Assays, g/t Metal

Distribution, %

Wt., g Cum. Wt.,

% Gold Silver Gold Silver

+150 1.9 0.4 12.50 8.9 9.5 1.2

-150 +106 15.9 3.6 4.49 5.0 27.4 5.2

-106 +75 38.4 11.3 0.87 2.7 12.8 6.8

-75 +53 47.2 20.7 0.84 3.7 15.1 11.4

-53 +45 26.8 26.0 0.57 3.3 5.8 5.7

-45 +37 30.3 32.0 0.53 3.7 6.1 7.3

-37 341 100.0 0.18 2.8 23.3 62.4

Total/Average 501.5 0.52 3.1 100.0 100.0

The results in Table 13.12 show that the actual grind is very close to the target P80, with 79.3 % of the

material passing the 53 µm set point. However, much of the material is quite fine, with 68 % being sub-

sieve. That said, there is strong evidence of coarse gold, as the 20 % of the residue that is + 53 µm

contains 65 % of the gold. The coarsest tail fraction has a grade of 12.5 g Au/t, more than twice the

calculated feed grade of 5.50 g Au/t. From there the grades drop progressively as the particle size

decreases. Silver behaves somewhat differently. Although the two coarsest fractions contain the highest

silver contents, the grades in all finer fractions are about the same. The 68 % of the material that is sub-

sieve contains about 62 % of the silver.

Due to the apparent presence of coarse gold, two different types of gravity test work were conducted.

One was the extended gravity recoverable gold (E-GRG) test. The other was gravity concentration with

cyanidation of the gravity tails.

E-GRG Program

The E-GRG test involves multiple grinds with tail screen analyses after each grinding stage. The initial

sample size was 8 kg crushed to the point where 100 % passed 20 mesh (M) (840 µm) [P80 of 25M

(700 µm]. In Stage 1 the feed was run through the laboratory Knelson gravity concentrator once and the

concentrate photographed, dried and assayed for gold. The tail fraction was dried and 250 g were split

out for a screen analysis. In Stage 2 the remaining tail fraction was ground to a P80 of 60M (250 µm) and

again run through the Knelson concentrator one time. Both fractions were then handled as in Stage 1. For

Stage 3 the remaining tail fraction was ground to a P80 of 200M (74 µm) and processed as in the first two

stages.

Page 13.20




AMENDED NI 43-101 TECHNICAL REPORT Composites 1 through 5 were subjected to the E-GRG tests. When analysing these tests, the gravity

recoveries in each stage are added together and the sum represents the maximum amount of gravity

recoverable gold in the sample. Actual commercial plant performance is usually slightly lower. The

detailed test results can also be used by the manufacturers of centrifugal gravity concentrators to provide

an optimum gravity circuit design. Typical test results for an individual sample (Composite 3) are plotted

in Figure 13.5 and Figure 13.6. Results for all five tests are summarized in Table 13.13 and Figure 13.7.

Figure 13.5 Gravity Recoverable Gold and Silver vs. Grind Size (Composite 3)

0

10

20

30

40

50

60

70

80

90

100

10 100 1000

Cu

m. G

old

Re

cove

ry, %

of

tota

l

Nominal Grind Size, µm

Cum. Au Recovery Cum. Ag Recovery

Page 13.21




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.6 Gravity Recoverable Gold by Size Fraction (Composite 3)

Table 13.13 Summary results of the E-GRG Tests for Gold

Composite

Gold Recovery vs. P80 Grind Size,

% of Gold in Feed Head Grade, g/t

700 µm 250 µm 75 µm Total Calc. Assay

Composite 1 15.0 27.7 15.1 57.8 2.49 2.76

Composite 2 10.3 23.6 15.0 48.9 3.41 3.46

Composite 3 19.2 18.5 16.3 54.0 4.49 4.88

Composite 4 16.6 14.5 17.6 48.7 5.18 5.58

Composite 5 10.4 16.8 16.3 43.5 9.78 9.74

0%

10%

20%

30%

40%

50%

60%

70%

80%

90%

100%

10 100 1000

Cu

m. G

old

Re

cove

ry, %

of

tota

l

Particle Size, µm

Stage 1 Stage 2 Stage 3

Page 13.22




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.7 Total Gravity Recoverable Gold vs. Head Grade

Typically the proportion of gravity recoverable gold increases as the head grade increases. However, the

San Ramon deposit is anomalous in this regard, trending downward as the grades increase. Although not

shown in Table 13.13, silver exhibits the same trend. The gravity recoverable percentage drops from

42.7 % in Composite 1 to just 24.9 % in Composite 5. Based on the trend line in Figure 13.7 gravity

recoverable gold would average less than 50 % life-of-mine.

Gravity / Cyanidation Program

The protocol for the gravity / cyanidation program started by splitting out a separate 5 kg sample of each

of the five grade composites. Each split was stage ground to a P80 of 65M (225 µm). Then the ground

product was passed through the Knelson concentrator one time. The resulting rougher gravity

concentrate was hand panned to produce a cleaner gravity concentrate, typically weighing less than 10 g.

This was photographed, dried, weighed and assayed in its entirety for gold and silver.

Both the cleaner and rougher gravity tails were separately dried and weighed. The two separate 1 kg

samples were reconstituted using the proper proportion of rougher and cleaner tails. The remaining

rougher tails were blended and splits cut for triplicate gold/silver assays. After regrinding to a P80 of 200M

(75 µm), one of the reconstituted tailings samples was given a direct leach following the leach protocol

given in above, but with an additional sampling time at 36 hours. The second sample was also leached

directly, but there was one additional slight change to the protocol. In the second test, the initial cyanide

concentration was set at 1 g NaCN/L, but no further additions were made at the intermediate sampling

times so that the cyanide concentration was allowed to coast down. The results of both sets of tests are

compared in Table 13.14. The original direct leach results at the same grind size are included in the table

for comparison.

y = 66.361x-0.183 R² = 0.7418

0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

0.00 2.00 4.00 6.00 8.00 10.00 12.00

Co

mb

ine

d G

ravi

ty R

eco

very

, %

Gold Head Grade, g/t

Page 13.23




AMENDED NI 43-101 TECHNICAL REPORT Table 13.14 Summary of the Gravity / Cyanidation Tests

Composite Test Type Gold Recovery, % of Total Reagent Consumption

Gravity Leach Combined NaCN, kg/t Lime, kg/t

Comp. 1 Standard 17.0 75.3 92.2 0.38 2.0

Comp. 1 Coast Down 16.9 77.8 94.7 0.68 1.6

Comp. 1 Direct Ore 91.8 0.74 4.2

Comp. 2 Standard 6.3 86.9 93.2 0.82 1.6

Comp. 2 Coast Down 6.3 86.1 92.4 0.35 1.6


Comp. 3 Standard 12.3 82.5 94.8 0.57 2.2

Comp. 3 Coast Down 12.4 81.8 94.2 0.60 1.9


Comp. 4 Standard 7.4 85.7 93.1 0.78 1.6

Comp. 4 Coast Down 7.3 85.0 92.3 0.45 1.6


Comp. 5 Standard 7.5 84.8 92.4 0.42 1.7

Comp. 5 Coast Down 8.7 83.4 92.1 0.44 1.8


For each composite there is some variability in the recovery in standard 1 g NaCN/L and the coast down

tests. However, this appears to be due simply to experimental variability as the average combined

recovery for all composites is 93.1 % for both types of tests. In addition, there is no apparent relationship

between head grade and combined recovery. On the other hand, there is much more variability in both

the gravity recovery (6.3 to 17.0 %) and the leach recovery (75.3 to 86.9 %). The conclusion is that the

leach will do a good job extracting whatever gold remains in the gravity tails. Although silver recovery is

not shown in Table 13.14, it generally follows the same trends as gold. The average recovery in the

standard test was 70.9 % vs. 71.0 % using the coast down.

While the average combined recoveries are the same for both types of tests, there are some important

differences with respect to reagent consumption. Not surprisingly, the final cyanide concentration was

lower in the coast down tests than in the standard tests. The respective average concentrations were

0.66 g NaCN/L vs. 0.94 g NaCN/L. This favours the coast down approach, as reagent consumption will

be lower in the subsequent detox operation. Furthermore, the actual reagent consumptions were also

lower with the coast down approach. Cyanide consumption was 0.50 kg/t ( and may go lower) with coast

down vs. 0.59 kg/t in the standard test. Lime consumption was 1.7 kg/t when using coast down vs

1.82 kg/t in the standard approach. Thus, the coast down approach would be preferred for plant

operation. In addition, the results suggest that further testing to optimize reagent concentration and

addition practices would be beneficial, as the 1 g NaCN/L appears to be higher than necessary to get

good leach extraction.

Page 13.24




AMENDED NI 43-101 TECHNICAL REPORT At the same primary grind (P80 of 75 µm) the gravity / cyanidation route appears to be more effective than

direct leaching. Recovery with the latter was incrementally lower, averaging 92.1 %. Silver recovery was

also lower, averaging 69.8 %. However, as noted above, higher recoveries can be achieved with a finer

grind. Furthermore, reagent consumptions in the direct leach were higher than in either

gravity / cyanidation option. Cyanide consumption in the direct leach was 0.80 kg/t and lime consumption

doubled to 3.5 kg/t.

Flotation-Combined Leach Program

As discussed above, a simple direct leach of finely ground ore was capable of extracting 94 to 95 % of the

gold and 73% of the silver. However, a number of problems were also associated this flowsheet.

Obviously the leach cycle time could be reduced and/or a coarser grind could be used to reduce costs.

However, these changes would likely reduce gold recovery to 91 or 92 %. These are lower than the

recoveries that were obtained by gravity concentration and tailing cyanidation. Therefore, an effort was

made to identify a less costly alternative process that would still give a recovery of at least 95 %. The

alternative selected for testing involves a fairly coarse primary grind; a flotation step focusing on recovery,

rather than grade; and an ultra-fine grind of the concentrate. The reground concentrate is then blended

back into the flotation tails and the combined product is agitation leached.

Due to the limited amount of material remaining in Composites 1 through 5, optimization of the flotation-

combined leach process was done using Composite 6. However, grind studies were conducted on all six

to establish the individual grind times needed to achieve a P80 of 125 µm for each composite. A P80 of 125

µm was selected upon analysis of KCA flotation work detailed in Table 13.4. When KCA conducted their

study of flotation recovery vs. grind size, they tested sizes that spanned the 125 µm size by using P80

levels of 150 and 75 µm. Recovery exceeded 95 % at the coarser grind and showed no improvement at

the finer grind. These results suggest that use of a coarser grind than the 125 µm one could reduce

grinding costs without jeopardizing recovery. In addition, it should be noted that recovery to concentrate is

not critical. Any gold that remains in the flotation tails may still be recovered in the combined leach.

A number of trials were run as part of the optimization study. In four of these the concentrate was

recombined with the flotation tails for a whole ore leach. In one of these, the concentrate was not

reground. In the other three, the concentrate was reground for either 4 or 11 minutes, or was lightly

reground to a P80 of 25 µm. An additional one-off test was run with a 4 minute regrind and mercury

deportment was tracked from the initial feed through to the final leach residue and PLS. A second one-off

test was run with a 4 minute regrind and the flotation tails and concentrate were leached separately. A

third one-off test was conducted merely to produce an unground concentrate sample for mineralogical

assessment (See Section 13.2.8).

In all these tests the primary grind was targeted to a P80 of 125 µm. Following the primary grind, 1 kg

charges were floated in four stages using 50 g/t PAX as the collector with a 2 minute conditioning time and

a 30 minute float time. The frother was Flomin F-579 and was added as needed. The tests were run at

the natural pH and 37 % solids. After any regrind, the concentrate was recombined with the flotation tails

and leached using essentially the same protocol as described for the gravity / cyanidation tests. The only

difference was that the tests were terminated after 48 hours, rather than 72. Results of the optimization

tests on composite 6 are summarized in Table 13.15.

Page 13.25




AMENDED NI 43-101 TECHNICAL REPORT Table 13.15 Results of the Optimization Tests Using Composite 6

Test Description Mass Pull,

% of Feed

Conc.

Regrind

Leach Recovery, % Reagent Usage

Gold Silver NaCN,

kg/t

Lime,

kg/t

Combined Leach 11.2 80% -25µm 91.1 70.1 0.22 1.8

Mercury Deportment 10.1 4 min. 94.8 65.7 0.61 1.6

Separate Product Leach 10.5 4 min. 95.11 70.02 0.393 2.04

Combined Leach 9.8 None 86.3 59.4 0.51 1.6

Combined Leach 10.7 4 min. 92.9 66.7 0.42 1.8

Combined Leach 9.7 11 min. 96.0 70.0 0.50 1.6

Note 1. 95.9% recovery from concentrate and 76.9% from float tails for weighted recovery of 95.1%.

Note 2. 74.2% recovery from concentrate and 37.5% from float tails for weighted recovery of 70.0%

Note 3. 2.11 kg/t NaCN for concentrate and 0.19 kg/t for float tails for weighted total of 0.39 kg/t.

Note 4. 4.8 kg/t lime for concentrate and 1.7 kg/t for float tails for weighted total of 2.0 kg/t.

Composite 6 responded very well to flotation. Mass pull to concentrate averaged just over 10 % of the

feed. Assuming an ore processing rate of 1,000 tonnes per day (tpd), only about 100 tpd will be advanced

to the regrind mill. Not all concentrates were assayed, but those that were contained an average of 95.8

% of the gold in the feed at an average assay of 42.7 g Au/t. Silver recovery to concentrate was nearly as

good, averaging 90.7 % of that in the feed. The silver content of the concentrate averaged 57.4 g Ag/t.

The gold recovery during leaching is clearly influenced by the degree of concentrate regrinding. As can

be seen in Table 13.15, gold recovery is only 86 % with no regrind. With just a minimal regrind to 25 µm,

recovery jumps to 91 %. Recovery increases to an average of 94 % with a 4 minute regrind and to 96 %

with the 11 minute regrind. These results are generally in line with the earlier KCA tests, which utilized a

coarser primary grind (P80 of 150 µm). In these tests recoveries were below 60 % with no regrind,

increasing to 94 or 95 % with a 30 minute regrind.

Silver behaves somewhat differently. With no regrind silver recovery is only 59 %. However, recovery

fluctuates in a narrow range between 65 and 70 % with any sort of regrind.

It is not clear what role the regrind plays in enhancing gold recovery. As discussed below (13.2.8) only

about 5 % of the gold is actually encapsulated in pyrite or quartz and would require breakage to expose

the gold to the lixiviant. The remainder of the gold is either fully liberated (30 %) or is partially rimmed or

encapsulated (65 %) and is accessible to the cyanide solution. Thus the main function of the regrinding

may be mainly surface scrubbing or abrasion to remove reagents occluded on the surface or cleaning off

any rims of pyrite of other constituents. This is supported by the particle size distributions shown in Figure

13.8. This shows that there is almost no difference in the particle sizing of the leached tailings with or

without regrinding.

Page 13.26




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.8 Concentrate Particle Size Distributions as a Function of Regrind Conditions

While the regrind appears to have little effect on the particle size distribution, it has a significant impact on

the residue assays as a function of particle size. This is shown in Figure 13.9. As can be seen, the assay

profiles with and without grinding are completely different. Without regrinding, the residue assays drop

progressively as the particle size drops. With both the 4 and 11 minute regrinds, just the opposite occurs.

The assays are lowest for the coarsest material and trend upwards as the particle size decreases. The

higher assays in the coarse fractions explain the lower recovery observed without a regrind.

The recovery with the 11 minute regrind is skewed by the one datum point at 2.5 g Au/t. Unfortunately this

was a very small sample (see Figure 13.9) and there was insufficient material left for another assay. If this

outlier is deleted, recovery at the 11 minute regrind would be even higher and the tail assays would be

lower than with the 4 minute regrind. This supports the decision to use the 11 minute regrind as the

design basis.

-10

0

10

20

30

40

50

60

0 20 40 60 80 100 120 140

We

igh

t o

f Sc

ree

n F

ract

ion

, %

Mid-Point of Screen Fraction, µm

Blue - no regrind Brown - 4 min. regrind Green - 11 min. regrind

Page 13.27




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.9 Residue Assays by Size Fraction with and without Regrinding

The reagent consumptions listed in Table 13.15 show that the cyanide consumption is generally low,

averaging 0.44 kg/t. However, the consumptions were somewhat erratic and ranged from 0.22 to 0.61 kg

NaCN/t. Lime consumption varied from 1.6 to 1.8 kg/t, except for the tests where the concentrate and

float tails were leached separately. It should be noted that these values are those measured when the

tests were terminated.

Two other results from the tests on Composite 6 deserve some discussion. With regard to the mercury

deportment, the approach was to track the mercury levels throughout the process. This involved running

mercury analyses on the feed, the flotation concentrate and tails, the final leach residue and the final PLS.

The object was to determine if mercury levels in the gold would be high enough to require inclusion of a

mercury retort in the gold doré smelting area. The head grade was 0.72 g Hg/t (ppm). The result was

very positive, as the mercury level in the PLS was just 0.003 ppm, showing that 99+ % of the mercury

reported to the final leach residue. Since the mercury does not follow the gold through the process, a

retort is not warranted.

In one of the tests on Composite 6, the reground rougher concentrate and the flotation tails were leached

separately, rather being combined and then leached. Gold recovery in the separate leach was 95.1 %.

This is within experimental error of the combined leach recoveries and suggests there is no benefit in

using separate leaches.

-0.5

0

0.5

1

1.5

2

2.5

3

0 20 40 60 80 100 120 140

Leac

h R

esi

du

e A

ssay

s, g

Au

/t

Mid-Point of Screen Fraction, µm

Test CY-46 w/ no regrind

Test CY-48 w/ 11 min. regrind

Test CY-47 w/ 4 min. regrind

Page 13.28




AMENDED NI 43-101 TECHNICAL REPORT The results from the separate leach test also provide some perspective on the leaching both the

concentrate and tailings. With the 4 minute regrind, gold extraction from the concentrate was 96 % vs.

just 77 % from the tails. There was also an order of magnitude difference in cyanide consumption, which

was 2.1 kg NaCN/t for the concentrate and less than 0.2 kg NaCN for the tails.

Based on the optimization studies on Composite 6, it was decided to use the 125 µm primary grind and

the 11 minute concentrate regrind when testing the five grade composites. In addition, no changes were

made to the flotation or combined leach protocols. Two extra tests were included in this part of the

program, one using Composite 3 and the other using Composite 6. In these tests the initial cyanide

concentration was adjusted to the standard set point of 1 g NaCN/L. However, no further cyanide

additions were made and the cyanide concentration was allowed to coast down until the tests were

terminated. Results of all tests with an 11 minute regrind are summarized in Table 13.16.

Table 13.16 Results of the Float-Combined Leach Tests on the Grade Composites

(with 11 min. concentrate regrind)

Sample Mass Pull,

% of Feed

Leach Recovery, % Gold Head Grade Reagent Usage

Gold Silver Calc.,

g Au/t

Assay,

g Au/t

NaCN,

kg/t

Lime,

kg/t

Composite 1 13.0 94.8 70.0 2.48 2.70 0.34 1.3

Composite 2 15.9 95.2 65.7 3.57 3.15 0.35 1.5

Composite 3 19.1 95.7 68.8 4.66 4.44 0.36 1.5

Composite 3

(coast down) 18.8 92.1 64.6 4.69 4.44 0.22 1.7

Composite 4 21.1 97.7 69.4 5.14 5.35 0.34 2.2

Composite 5 20.6 98.1 85.3 9.87 9.50 0.45 1.6

Composite 6 19.7 96.0 70.0 5.22 5.32 0.50 1.6

Composite 6

(coast down) 19.7 96.5 63.2 4.92 5.32 0.36 2.0

The key parameter in these tests is the gold recovery. As hypothesized in the PEA, there is a positive

correlation between head grade and recovery. This is shown in Figure 13.10A. As can be seen, there is a

single point well off the trend line, which raises that line by about 0.2 percentage points. Statistically, this

point proved to be an outlier. Figure 13.B shows the head grade-recovery relationship plotted without the

outlier. The resulting trend liner is almost identical to the original one. The main difference is the much

higher degree of correlation (R2 value) without the outlier (removal of which makes the data more

conservative). The standard deviation for the latter is 0.2% and the 95% confidence interval is +/- 0.6%.

It should be noted that this expression just relates to the leach step. It does not take into account the

small post-leach losses due to gold in the tail solution, etc.

Page 13.29




AMENDED NI 43-101 TECHNICAL REPORT Figure 13.10A A Relationship between Gold Head Grade and Recovery

Figure 13.10B Adjusted Relationship between Gold Head Grade and Recovery

Leach kinetics are not reflected in Table 13.16 or in Figure 13.10A and B. All results shown are final

extraction levels when the tests were terminated after 48 hours. In some cases the extraction reached

completion in 36 hours. However, in about half the tests, extraction was continuing, albeit slowly, when

the tests were terminated.

y = 0.451x + 93.924 R² = 0.7191

94.5

95

95.5

96

96.5

97

97.5

98

98.5

99

0 2 4 6 8 10 12

Go

ld R

eco

very

, %

Calculated Gold Head Grade, g Au/t

y = 0.4519x + 93.628 R² = 0.9991

94.5

95

95.5

96

96.5

97

97.5

98

98.5

0 2 4 6 8 10 12

Go

ld R

eco

very

, %

Calculated Gold Head Grade, g Au/t

Page 13.30




AMENDED NI 43-101 TECHNICAL REPORT Silver recovery does not show the same grade-recovery relationship as gold. The average silver recovery

for the tests is 71.5 %, exclusive of the coast down tests. However, there is again a single very high value,

which is associated with the highest grade sample. Without this, the average drops to 68.8 % and is

clearly independent of head grade. Note that the highest and lowest grade samples have the highest

silver recoveries. The silver recovery dropped to about 64 % in the two coast down tests.

One unexpected aspect of the results presented in Table 13.16 is the generally high mass pull in flotation.

The KCA program suggested a mass pull level of about 10 %; as did the earlier MLI tests (see Table

13.15). However, the KCA value was obtained with low grade samples that are below the current cut-off

grade. With the five higher grade underground samples, the lowest grade sample has a mass pull of

almost 13 % and the mass pulls are above 20 % for the highest grade composites. Such high mass pulls

may represent a risk in the sizing of the tower mill selected for the plant.

Another possible risk arises because the laboratory used a pebble mill with ceramic media to do the

concentrate regrinding. However, a tower mill will be used in the plant. Thus, the laboratory concentrate

grind curves and particle morphology may not replicate the product produced in the plant. As a result, a

new grind curve (grind time vs. particle size) will likely be required for the plant. Reagent consumptions

were generally in line with the optimization tests on Composite 6, with an average cyanide consumption of

0.39 kg/t, exclusive of the coast down tests. Lime consumptions were also similar. Consumptions of both

reagents were lower than in the direct leach tests. As with extractions, the reagent consumptions given

are those at test termination. However, almost all reagent consumptions appeared to stabilize after the

first 16 hours, with virtually no increase after that. Interestingly, the cyanide consumption in the coast

down tests was lower than in the tests at a constant 1 g NaCN/L, averaging just 0.28 g NaCN/L.

Cyanide Destruction Tests

Cyanide destruction (detox) tests using sulfur dioxide (SO2) as the reagent are underway on the final

leach residues. No results are available as of the effective date of this report; however, very conservative

estimates for consumables have been adopted for the operating cost estimate

Liquid / Solid Separation

A series of thickening and filtration tests are being conducted on the detoxified leach residue by Pocock

Industries in Salt Lake City, Utah. No results are available as of the effective date of this report; however,

the results are not expected to affect the design criteria adopted in the Feasibility Study

Carbon Loading Tests

Carbon loading and stripping tests are underway at MLI. No results are available as of the effective date

of this report; however, results are not expected to affect the design criteria adopted in the Feasibility

Study.

Page 13.31




AMENDED NI 43-101 TECHNICAL REPORT 13.2.7 Comminution

The initial comminution test work was performed by Phillips Enterprises LLC of Golden, Colorado on a

subcontract basis. When this facility closed due to retirement of the owner, further comminution work was

performed at ALS Metallurgy, Kamloops, British Columbia on a direct contract basis.

Phillips Comminution Tests

The initial comminution tests were run on the six KCA variability samples, the Hilo Azul and saprolite

samples, and a barren granodiorite sample that was not tested metallurgically (Phillips 2013, 2-part letter

reports). Samples were received in two batches, one on January 25, 2013 and the other on April 30, 2013.

Work was completed on May 22, 2013. The resulting rod mill and ball mill Bond work indices and

abrasion indices are summarized in Table 13.17. The work indices are given in kilowatt hours per tonne

(kW-h/t). Further details are available in the referenced reports.

An important caveat regarding these tests is that they were conducted on material that represented the

open pit / underground mining scenario. Except for the high grade Hilo Azul material, all samples have

head grades at or below the underground cut-off grade (2.0 g Au/t). Samples E2 and W3 are at or just

above the cut-off level. The other five samples would be classified as waste in the underground operation.

Thus, these samples may not be closely representative of typical mine output and could be either harder

or softer than the higher grade material.

Table 13.17 Comminution Test Results

Sample Rod Mill Work Index, kW-

h/t

Ball Mill Work Index, kW-

h/t1 Abrasion Index

Hilo Azul Sulfide 12.03 12.87 0.3345

Granodiorite Waste 15.49 18.32 0.2068

Saprolite 4.01 4.52 0.0010

E1 Variability N/A 15.15 0.1215




W1 Variability N/A 13.56 0.0952

W2 Variability N/A 14.93 0.1635

Note 1. Bond ball mill tests were conducted using a closing screen size of 106 µm.

Exclusive of the saprolite, the ore samples have fairly consistent work indices, with an average ball mill

work index of 14.85 kW-h/t. This value indicates that the ore is of average hardness. The granodiorite

waste is somewhat harder, while the high grade Hilo Azul is softer, probably due to its high sulfide

component. There is a wider range for the abrasion indices. Values below 0.1 indicate rather non-

abrasive material, while values of 0.2 or more are fairly abrasive. As can be seen, the saprolite is very soft

and non-abrasive.

Page 13.32




AMENDED NI 43-101 TECHNICAL REPORT ALS Comminution Tests

Two separate comminution programs have been conducted by ALS. One involved the determination of

standard comminution parameters – low impact crushing and ball mill work indices, plus abrasion indices

(Mehrfert, 2014). The other was a SAG mill comminution (SMC) test program (Mehrfert, 2014; Weier,

2014).

The first program involved two different sets of samples. The four samples used for the bond low impact

crushing work index (CWi) determinations are shown in Table 13.24. In these tests, individual pieces of

rock from each sample are tested. The abrasion index (Ai) and ball mill work index (BMWi) were

determined for each of the six underground grade composites previously described (Table 13.18).

Table 13.18 Description of Crusher Work Index Samples

Sample ID Kg No. of Pieces Tested Description

Sample 1 4.5 6 Sericite altered granodiorite wallrock

Sample 2 6.4 10 Low sulfide ore, 1-4% sulfides + quartz, calcite & sericite

Sample 3 6.7 8 Medium sulfide, 4-10% sulfide, mainly quartz, minor

calcite matrix

Sample 4 6.2 7 >10% sulfides, gangue mainly quartz or brecciated

sulfides

The CWi results are summarized in Table 13.19. The abrasion and ball mill work indices are summarized

in Table 13.20. Additional details on both programs are provided in the referenced reports.

Table 13.19 Bond Low Impact Crusher Work Index Results

Sample ID Crusher Work Index (kW-h/t) Specific

Gravity Average Minimum Maximum Std. Deviation

Sample 1 10.03 1.81 16.35 6.40 2.72

Sample 2 7.77 1.80 24.12 6.55 2.76

Sample 3 8.08 4.56 14.25 3.43 2.84

Sample 4 7.75 4.17 12.63 3.09 2.91

Page 13.33




AMENDED NI 43-101 TECHNICAL REPORT Table 13.20 Bond Ball Mill and Abrasion Index Results

Sample ID Ball Mill Results

1 Abrasion

Indices F802, µm P803, µm Work Index, kW-h/t

Composite 1 2243 83 16.6 0.089

Composite 2 1815 81 15.6 0.047

Composite 3 2117 81 14.6 0.122

Composite 4 2036 81 16.7 0.085

Composite 5 2067 82 14.1 0.041

Note 1. Bond ball mill tests were conducted using a closing screen size of 106 µm.

Note 2. F80 is the screen size that passes 80% of the feed.

Note 3. P80 is the screen size that passes 80% of the ground product.

Standard protocol for the CWi determinations is to break 20 individual pieces. Averaging the results then

gives a reliable result, even though there is typically considerable variation from piece to another. Since

an average of only eight (8) pieces was broken per ore sample, there is some uncertainty in the individual

results. However, in all, 25 pieces of ore were broken and the results were quite consistent for all three.

The average was 7.87kW-h/t, which indicates that the ore is very soft with respect to crushing energy

requirements. The single waste sample appears to have a somewhat higher crushing energy

requirement, as the average CWi is just over 10kW-h/t. This is consistent with the Phillips work, which

showed that the barren material also had a higher milling work index than the ore.

In addition to determining the CWi value for each sample, the specific gravities were also determined. Not

unexpectedly, the specific gravities increased progressively as the sulfide level increased. Because the

comminution tests only affect the physical nature of the samples, and not their chemical make, ALS also

ran head assays on each of the four samples as requested by Red Eagle Mining. Key analyses are

shown in Table 13.21.

Table 13.21 Head Assays on CWi Samples

Sample ID Au, g/t Ag, g/t Fe, % S, % Cu, % Pb, % Zn, % As, %

Sample 1 <0.01 0.7 3.89 0.20 0.017 0.001 0.01 0.002

Sample 2 11.0 25 3.81 3.65 0.009 0.07 0.90 0.028

Sample 3 22.7 37 6.14 8.58 0.042 0.38 3.10 0.255

Sample 4 62.6 91 13.95 >10 0.042 0.83 3.35 4.52

The close agreement of the average CWi values for the three ore samples in spite of the disparity in the

head assays suggests that a CWi value of 8 kW-h/t is a reasonable value for initial crusher design.

However, it should be noted that the head grades of gold-bearing samples are all well above the life-of-

mine average.

Page 13.34




AMENDED NI 43-101 TECHNICAL REPORT The BMWi values for the five ore-grade composites average 15.5 kW-h/t, but do exhibit a slight negative

correlation with increasing head grade. See Figure 13.11. This trend is consistent with the BMWi of

12.87 kW-h/t for the high-grade Hilo Azul sample. Using the equation for the trend line and the life-of-

mine head grade gives a BMWi value of 15.4 kW-h/t, the same as the average value. This value is also

consistent with the earlier Phillips results.

Five of the six grade composites have Ai values below 0.100, with an average value of 0.0768. Values

below 0.100 are associated with soft, nonabrasive material. The earlier values determined by Phillips on

the six variability samples were slightly harder and more abrasive, with an average Ai value of 0.1216.

The inverse relationship between head grade and the Ai values is consistent with the increased sulfide

levels at higher head grades, as the sulfides should be softer than the host rock.

Figure 13.11 Relationship between Head Grade and BMWi Values

Six samples were submitted to ALS for the SAG mill comminution (SMC) tests. These were simply

identified as 1 through 6 and were selected to represent stoped material that would be mined

progressively over the life of the operation. Each sample contained approximately 14 kg of material and

consisted of either quarter core or individual fragments that would provide at least 100 pieces suitable for

breakage in the test. No consideration was given to grade, but all material came from intervals with a

potentially workable grade that is likely to go through the mill.

y = -0.2697x + 16.885 R² = 0.4014

0

2

4

6

8

10

12

14

16

18

0 2 4 6 8 10 12

Ball M

ill W

ork

In

dex,

kW

-h/t

Gold Head Grade, g/t

Page 13.35




AMENDED NI 43-101 TECHNICAL REPORT ALS utilizes the JK Tech proprietary SAG mill design system. With this system ALS stage crushed each

sample and then selected 100 fragments in the 19 x 22.4 mm fraction for breakage. Five sets of 20 were

broken; each set run at a specific energy level using a JK Drop-Weight tester and the broken product was

screened at a specified sieve size. The results were forwarded to JK Tech for full evaluation. JK Tech

determines the Drop-Weight index (DWi), which is a measure of the rock strength under impact

conditions. The breakage parameters A and b were also estimated. The derived product A*b is then

compared to the extensive JK Tech data base to determine the hardness with respect to a SAG mill

operation. The higher the product, the softer the ore. Selected SMC test results are shown in Table 13.22.

The average value of A*b for Samples 1 through 5 is 44.5. This is right in the middle of the JK tech data

base and indicates ore of medium hardness with respect to grinding. Sample 6 is significantly harder,

having hardness greater than about 75% of the samples in the JK Tech data base. The SMC test results

are being used to size the SAG mill.

Table 13.22 Selected SMC Test Results

Sample ID DWi,

kWh/m3

A b A*b

Sample 1 5.57 68.0 0.70 47.6

Sample 2 6.46 72.3 0.58 41.9

Sample 3 5.63 59.7 0.81 48.4

Sample 4 6.68 69.1 0.59 40.8

Sample 5 6.36 66.0 0.66 43.6

Sample 6 7.85 73.7 0.47 34.6

13.2.8 Ore and Concentrate Mineralogy

Economic Geology Consulting Studies

A total of five mineralogical studies have been completed as part of the Feasibility Study program.

Economic Geology Consulting (EGC) of Reno, Nevada conducted one study of concentrate produced in

the flotation tests at KCA. ECG also undertook two related studies on slices of drill core. Two additional

studies were performed on samples of concentrate produced in the MLI program on the flotation and

leaching of the underground ore samples. The latter were run by Process Mineralogical Consulting Ltd.

(PMC) of Maple Ridge, British Columbia.

Chronologically, the first three reports were prepared by EGC (EGC January, 2014; February, 2014; and

again February, 2014). These all utilized optical petrographic techniques. The earliest addresses the

mineralogy of four cleaner concentrate samples produced by KCA. Three of the samples (66861 A,

66861 G and 66862 A) were produced from the low-grade master composite 66809, while 66864 A was

produced from the high-grade Hilo Azul composite. Further details regarding the sources of these samples

are given in Section 13.2.1.

Page 13.36




AMENDED NI 43-101 TECHNICAL REPORT All four samples have essentially the same mineralogy, which consists of gold and several sulfides in

varying proportions. Paragenetically the early pyrite is rimmed, veined and replaced with subsequent

mineralization. Each sample is dominated by pyrite, followed by sphalerite and pyrrhotite. There are

lesser amounts of galena and chalcopyrite. Due to the presence of pyrrhotite, all samples are magnetic.

There is a mixture of coarse-grained and fine-grained gold, with particle sizes ranging from >100 µm to

<1 µm. The free gold tends to be finer, but may have been impacted by concentrate regrinding. Gold

encapsulation in pyrite is common, with additional encapsulation in quartz or other sulfides. See

Figure 13.12.

Figure 13.12 Pyrite Grain Cut by Gold (arrow) Veinlet (Credit EGC)

The second EGC report describes a petrographic study of 13 core samples. All are vein samples.

Page 13.37




AMENDED NI 43-101 TECHNICAL REPORT With one exception (SR-50054), all samples have very similar mineralogy and textures. The early pyrite

has been sheared and crushed in multiple events, then cemented by quartz and various sulfides. Late

stage calcite then pushes through the quartz and various sulfides, including pyrite. A typical polished

section is shown in Figure 13.13. The samples contain quartz and pyrite, followed by sphalerite. There

are lesser amounts of galena, chalcopyrite, calcite and sericite. Most samples contain visible gold, often

associated with galena or sphalerite microveinlets cutting through brecciated pyrite. Additional gold

occurs as fine inclusions in pyrite or quartz. SR-50054 is dominantly quartz with fragments or vug filling of

pyrite and sphalerite and does not exhibit any evidence of shearing. Gold occurs in the vugs as

replacement for the pyrite.

Figure 13.13 A Polished Section of Sample SR-50051 Showing Crushed Pyrite Cemented by

Sphalerite, Galena-Gold, Quartz and Late-Stage Carbonate Veinlets (Credit EGC)

Page 13.38




AMENDED NI 43-101 TECHNICAL REPORT The third EGC report is an extension of the second report. In this part of the study three samples (SR-

2958, SR-15615B and SR-50056) were re-examined in reflected light in an effort to differentiate between

pyrite and arsenopyrite. Arsenopyrite was identified in the latter two samples. It appears to predate the

pyrite and has also been crushed and sheared and cemented together as a breccia. Although the

arsenopyrite does contain some gold, the association is less common than for pyrite or the other sulfides.

(See Figure 13.14).

Figure 13.14 Crushed Arsenopyrite (left)-Pyrite (right) Contact with Gold (arrow) Along the

Contact and Extending Along the Fractures (Credit EGC)

Page 13.39




AMENDED NI 43-101 TECHNICAL REPORT Process Mineralogy Consulting Studies

Both the PMC studies (Geoff Lane June, 2014 and August, 2014) were made on concentrates produced

from Composite 6, which has an assay head grade of 5.29 g Au/t. See Section 13.2.6 Sample

Characterization for further details on this sample. In a sense, the first study is preliminary as it was

performed on a reject sample. The later report covers an assessment of the entire concentrate output from

a single test. Thus, it provides a more comprehensive assessment of the concentrate sample. This

started with a particle size determination using a cyclosizer at Blue Coast Research Ltd. in Parksville,

British Columbia. Results are shown in Table 13.23.

Page 13.40




AMENDED NI 43-101 TECHNICAL REPORT Table 13.23 Cyclosizer Particle Sizing on Unreground Composite 6 Rougher Concentrate

Size Designation

Effective Particle

Separation Size,

µm

Sample Weight1,

G Passing Weight, %

Gold Assay,

g Au/t

CS-1 46.3 0.83 99.26 100.6

CS-2 34.7 3.88 95.78 55.5

CS-3 24.2 14.04 83.21 39.8

CS-4 15.8 9.71 74.52 25.9

CS-5 11.6 14.07 61.92 16.1

CS-5 - 3.3

Note 1. Feed sample weight was 111.7 g.

As can be seen, the concentrate is quite fine, even without a regrind. Over 95 % of the material is sub-

sieve (<400 M) and about 60 % is minus 10 µm. The P80 is just over 20 µm, generally in line with the

concentrate grind study reported above.

For purposes of mineralogy, PMC combined selected cyclosizer products and only evaluated four

samples: CS-1 + CS-2, CS-3, CS-4 + CS-5, and minus CS-5. Polished sections were prepared from each

of the four and systematically scanned using the Tescan Integrated Mineral Analyser (TIMA) Scanning

Electron Microscope equipped with additional semi-quantitative elemental analysis software to measure

grain size and elemental composition. Then a subsequent scan was used to determine gold deportment

and degree of liberation. The main conclusions of the principal PMC study are as follows:

Overall, the concentrate consists of mica, pyrite, quartz, feldspar and pyroxene, with

lesser calcite, pyrrhotite and chlorite. Pyrite dominates the coarsest size at 68 %,

decreasing progressively to just 5 % in the finest fraction. Mica shows the opposite

trend increasing from 3 % in the coarsest fraction to 34 % in the fines;

About 95% of the gold is actually present as electrum with an average composition of 57

% gold and 43 % silver. Overall, gold ranges from 83 % down to 29 %. While the

electrum appears to represent all of the gold, the same is not true for the silver. About

half the silver is unaccounted for. As no silver-bearing phases other than electrum were

detected in the TIMA scan, the remaining silver is likely distributed at low concentrations

in various minerals, such as tetrahedrite;

Overall free grains of gold represent 30 to 35 % of the total. Only about 5 % are actually

fully encapsulated in pyrite or other minerals. The balance is either exposed on or

attached to pyrite or other mineral assemblages; and

In terms of size distribution, about 83 % of the gold grains are finer than 8 µm.

However, just the opposite is true for mass. About 82 % of the gold weight is in grains

coarser than 8 µm.

Page 13.41




AMENDED NI 43-101 TECHNICAL REPORT 13.2.9 Metallurgical QA/QC Program

A good QA/QC program for the metallurgical test work is just as important as it is in the drilling program. If

metallurgical samples are somehow switched, test results are not reproducible or are biased, or assays

are unreliable, the proper decision may not be made regarding project viability. There were three main

components to both the KCA and MLI metallurgical QA/QC programs:

The sample tracking system;

Metallurgical testing protocols; and

Assay accountability.

Sample Tracking System

All samples delivered to each laboratory are entered into a similar sample-tracking system. The delivery

date, shipping company, and shipment tracking number are recorded. Then a project name is designated

with a unique project number. Each sample is given a unique sample number based on the project

number, then weighed and in most cases photographed. If samples are composited in some way, the

composite is given a new unique number and cross referenced back to the samples that went into the

composite. When a sample is used in some sort of metallurgical test, that test is also given a unique

designation that indicates the type of test, e.g. rougher flotation, whole ore leach, etc. Each test is cross

referenced back to the sample tracking log. The log remains open until the samples are disposed of or

returned to the client.

Metallurgical Testing Protocols

Where there is sufficient sample, an effort is made to conduct at least 10 % of the metallurgical tests in

duplicate. The purpose is to determine the inherent variability in the testing procedure. This must be

small. Otherwise it will be difficult to determine if a change in metallurgical response is due to a change in

a process parameter, such as grind size, or merely experimental variability.

Written instructions are prepared for each test to avoid any confusion by the technicians actually

performing the work. This includes checking all equipment to be sure it is clean and free of material from

previous tests. Probes for pH, Eh, and dissolved oxygen are checked daily and calibrated before each

use. Balances and scales are checked weekly using certified weights and are checked annually by an

outside service provider. As much of the data relate to bench-scale testing, calculated and assay head

grades are compared and tailings are routinely checked to determine if any re-assays are required or if a

test should be rerun.

Page 13.42




AMENDED NI 43-101 TECHNICAL REPORT Comminution Protocols

The ALS laboratory is ISO 9001:2008 certified and has been since 2001. Annual drop weight tests are

performed as required by JK Tech, to confirm the accuracy of the equipment for the full drop weight and

SMC tests. ALS has prepared internal standard samples to check the calibration of the Bond Ball and

Bond Abrasion mills. These are checked annually. All comminution test equipment is properly maintained

and any variances are addressed immediately. All media loads and balances are checked weekly. All

scales are inspected annually by an external auditor. ALS also participates in regular round robin tests for

the ISA Mill Signature Plot test apparatus, and is certified to conduct these regrinding tests.

No details could be obtained regarding the QA/QC procedures used at Phillips Enterprises. This

laboratory was closed unexpectedly midway through the MLI program and the equipment has been sold to

another laboratory.

Assay Accountability

All the laboratories generally follow similar analytical protocols. The laboratories typically participate in

round robin proficiency test programs to help maintain both analytical accuracy and precision.

Assuming sufficient sample, all assays are run in duplicate or triplicate to ensure that the highest quality

analyses are reported. For the analysis of solutions, duplicate analyses are completed utilizing separate

techniques where possible. In the case of solid samples, MLI periodically submits duplicate pulps and/or

carbon samples to two separate commercial labs for analysis. Typically, around 10 % of the solid

samples are re-submitted for assay to a third party commercial laboratory.

For solids being assayed for gold and silver, the average assay, the standard deviation, and the relative

standard deviation of each assay set are examined, and additional check assays are completed if

required. For duplicate assays, relative standard deviation values of less than 10 % for gold and less than

20 % for silver are expected. Standard deviation results greater than those identified would trigger a re-

assay (if material is available).

Overall reproducibility for a test program is also important, and the average assay, the standard deviation,

and the relative standard deviation for calculated heads from test sets are also examined; tests may be re-

run if a single test deviates significantly from the set average.

Analyses completed in-house are performed utilizing methods and procedures which are standard in the

industry. Assay methods for gold include both a fire assay method with a wet finish as well as a fire assay

method with a gravimetric finish. Silver assays are conducted in a similar manner.

Material for a multi-element analysis is digested using a four-acid digestion. This digestion provides for a

total digestion of the material. Whole-rock analyses are conducted using a lithium metaborate fusion

followed by ICAP-OES analysis. The ICAP-OES units use certified reference solutions for calibration.

ICAP-OES units are maintained by the manufacturing company with preventative maintenance visits

throughout the year.

Page 13.43




AMENDED NI 43-101 TECHNICAL REPORT Checks of individual samples (between 10 and 20 %) are routinely completed on each sample group.

Blanks and certified reference material are utilized throughout the analytical procedures. The blanks,

standards and duplicates are included in each group of samples on a random basis.

Solution assays are made by flame atomic absorption spectrophotometer (“FAAS”) methods using

certified gold and silver standards. Solutions are routinely checked by fire assay methods to confirm the

calibration of the FAAS unit. FAAS units are maintained by the manufacturing company with preventative

maintenance visits throughout the year.

Only reagent-grade chemicals are used in the analytical work.

13.3 Summary of Metallurgical Test Work

Both key metallurgical results and metallurgical risks and opportunities are covered in this section. Each

is additionally covered in a separate section.

13.3.1 Metallurgical Results

The AcmeMet results are of limited value given the small number of grab samples tested and a focus on

different processing options. As with the AcmeMet program, much of the work done at KCA is now of

limited value due to the shift in expected ore grades. Most of the KCA testing was done on samples with

head grades that are now at or below the expected gold cut-off grade of 2.00 g Au/t. Some of the findings

in the KCA program do still relate to the currently planned operation and these are listed below.

Gold recovery averaged 93 % in whole ore leach tests where the grind had a P80 of 75

µm, consistent with results at AcmeMet and the confirmatory tests conducted at MLI;

In tests to determine flotation recovery as a function of primary grind size, recovery was

95 to 96 % at a P80 of 150 µm, with some increased loss at 212 µm. Grinding down to a

P80 of 75 µm provided no observable improvement. The implication is that if the ore is

harder than expected and is coarser than the projected grind size, recovery should not

suffer significantly; and

When leaching the flotation concentrate, KCA only achieved 53 % recovery on unground

material. With a 30 minute regrind, recovery jumped to 94 %. (Note: KCA did not

determine the particle size distribution on the reground material).

MLI is concluding a broad test work program, which is based on the use of several samples that span the

range of expected underground ore grades. Key results from the program are listed below. These are

followed by an identification of risks and opportunities:

There is a strong linear relationship between the gold and silver head grades, which can

be expressed as Ag head grade = 1.6119 x (Au head grad) + 1.1155. Using this

relationship the average silver head grade is 1.85 times the gold grade;

Page 13.44




AMENDED NI 43-101 TECHNICAL REPORT The whole rock analyses show that the makeup of all composites is remarkably similar.

Silicon (as SiO2) makes up 60 to 62 % of the material, probably mostly as quartz, plus

feldspar and mica. This is followed by aluminum (as Al2O3) at 10 to 13 %, probably in

mica and feldspar. Iron (as Fe2O3) makes up 6 to 8 % of the composites, mostly as

pyrite and pyrrhotite;

Diagnostic tests were run on a low grade sample to identify the possible need for air or

oxygen sparging to improve extraction. The results showed that such steps were not

necessary for whole ore leaching. However, these tests were not extended to the higher

grade combined leaches. There is some risk that the combined leaches might have

benefitted from an air sparge, as dissolved oxygen (DO) levels dropped below 2 ppm at

various times during these tests;

Whole ore leach tests showed there was a strong inverse linear correlation between the

P80 grind size and gold extraction, reaching 95% at 37 um. This was independent of

head grade. Silver recovery also increased with decreasing particle size, reaching 73%

at 37 um with no observable dependence on head grade. Cyanide consumption was

erratic, but averaged 1.11 kg NaCN/t;

Due to the presence of reactive sulfides like pyrrhotite, an alkaline aeration pre-

treatment, was tested to oxidize these sulfides and reduce cyanide consumption.

Results were disappointing. While there was a modest decline in cyanide consumption,

gold recovery also dropped marginally and lime consumption increased significantly;

Tail residue assays and mineralogical studies indicated the presence of some coarse

gold. As a result two separate gravity programs were conducted;

E-GRG test results indicate that the maximum gravity recoverable gold ranges from 58%

for the lowest composite grade downward to 43.5% for the highest grade composite.

Such behaviour is anomalous, as the normal trend is for recovery to increase as the

head grade increases; and

A final gravity concentrate was recovered and the combined gravity tails were leached

after being ground to a P80 of 75 µm. Overall recovery averaged 93%, with an average

cyanide consumption of 0.59 kg NaCN/t.

Optimization tests on a single high grade composite showed that a flotation step at a primary grind at a P80

of 125, followed by a combined leach on the concentrate and float tails gave the best results. This

confirmed earlier results at KCA, which showed that extraction was poor without a regrind.

In the flotation-combined leach optimization tests an 11 minute regrind achieved 96 % gold recovery. This

regrind time was selected, as it produced lower tail grades than shorter regrinds.

Page 13.45




AMENDED NI 43-101 TECHNICAL REPORT When the optimized float-combined leach tests were run on the grade composites,

recovery was found to be an inverse linear function of the head grade. After deletion of

a statistical outlier, the relationship is Gold Recovery (%) = 0.4519 x Gold Head Grade +

93.628. The standard deviation is 0.2% recovery and the 95% confidence level is 0.6%.

Life of mine recovery would be 96%, exclusive of any post leach losses;

In the optimized float-combined leach tests silver recovery was independent of head

grade and averaged 68.8 %, exclusive of one statistical outlier and any post-leach

losses;

A float-combined leach test was conducted to track mercury deportment during

processing. Results showed that virtually all mercury reported to the final leached tails.

The mercury content in the final pregnant leach solution going to gold recovery was just

0.003 ppm;

The crushing work index for the ore grade samples averaged 8 kW-h/t, indicating soft

ores in terms of crushing. The barren waste was higher at 10 kW-h/t;

Comminution testing showed that the ball mill grinding index was fairly constant at

approximately 15 kW-h/t over the normal range of ore grades. Barren waste, as might

come from dilution, was over 18 kW-h/t. This material was also moderately abrasive.

Most of the ore grade samples were soft and non-abrasive;

SMC tests indicated that San Ramon deposit ore is of medium hardness with respect to

grinding. One sample was significantly harder than the other five, indicating some

variability in the ore;

Petrographic studies on slices of core showed that the early pyrite and arsenopyrite had

been crushed and sheared then cemented back together with later quartz and other

sulfides, all cut through with late stage calcite. The final mineralized material was

therefore highly brecciated;

Other mineralogical studies showed that most of the gold is present as electrum with an

average content of 57% Au and 43% Ag. This accounts for virtually all of the gold, but

only about half the silver. The remaining silver likely occurs at low levels in tetrahedrite

and other minerals;

About 30% of the gold is free or liberated and only 5% is fully encapsulated in pyrite or

quartz. The remainder is partially rimmed with sulfide or partial inclusions in other

minerals;

In terms of particle size, about 80% of the gold particles are < 8 µm. On a mass basis,

about 80% of the gold is present in particles > 8 µm; and

Page 13.46




AMENDED NI 43-101 TECHNICAL REPORT In general, QA/QC lab practices are adequate and consistent with normal industry

practices.

13.3.2 Risks and Opportunities

In addition to the main sulfide composite, KCA ran direct leach tests on six variability samples. Each was

prepared from intervals in closely space drill holes. Geographically these were spread across the shear

zone from the eastern edge of the concession to the western portion of the deposit. Head grades for the

variability samples were low, ranging from 0.8 g Au/t to just above the current cut-off grade. While the

sample grades are not relevant, the fact that the test results showed a clear trend along the mineralized

zone is important. Going from east to west, there was some falloff in gold recovery and a significant drop

in cyanide consumption. The latter dropped from about 5 kg NaCN/t in the east, to just 1.5 kg NaCN/t in

the west. To mitigate the risk of such variations going forward, a new set of east-to-west variability

samples should be prepared from ore-grade material and tested using the flotation-combined leach

approach, as the mining schedule dictates

There is an additional risk associated with the six MLI grade composites. On the one hand all six are

made up of material drawn from throughout the central and eastern areas of the deposit. As such, all six

are highly representative of the life-of-mine material. As a result, any localized variability in metallurgical

characteristics will be masked by the broad makeup of the composites. Furthermore, the western portion

of the deposit is not well represented in these composites. In effect, no variability tests have been run on

the higher grade underground ore. Mitigation would require preparation and testing of higher grade

variability samples.

As part of the gravity / cyanidation program two parallel tests were run on each composite. In one the

cyanide concentration was held at 1 g NaCN/L by periodically adding cyanide as required. In the second,

the initial cyanide content was adjusted to the 1 g NaCN/L, but the cyanide level was allowed to coast

down during the remainder of the tests. Results were encouraging and warrant further consideration.

There was no decrease in recovery, a 15 % drop in cyanide consumption and a drop in the cyanide

concentration going to cyanide destruction. Additional tests should be run to optimize the cyanide dosage

and addition schedule to reduce reagent consumption.

The float-combined leach tests on the six high-grade composites yielded high mass pulls that increased

from 13 to 21 % as the head grades increased. This was a higher range than the 11 % design basis. As

a result there may be an issue with the capacity of the regrind mill in the design. Mineralogy on these

concentrates showed that they contained high levels of mica, quartz, feldspar and pyroxene, which

together undoubtedly accounts for the higher than expected mass pull. Further tests to optimize flotation,

including reagent selection and dosage, may reduce the mass pull by eliminating the high gangue

carryover.

Thickening of tailings slurry, with recycle of the barren solution prior to cyanide destruction, will recover

cyanide and minimize reagent consumption in the detox operation. An evaluation of this option may show

that it is a viable alternative to detoxification of the entire leach residue.

Page 13.47




AMENDED NI 43-101 TECHNICAL REPORT In conclusion, the project is moving ahead positively with a combined process that involves the following

steps:

A primary grind to a P80 of 125 microns.

A rougher flotation step with 95 to 96% recovery of gold to the concentrate.

Regrinding the concentrate to a P80 of 20 microns.

Leaching the combined concentrate and flotation tails in a CIL circuit for 48 hours at a

cyanide concentration starting at 1 g NaCN/L.

Life-of-mine, the resulting gold recovery is expected to average 96 %, with 69 % silver recovery. Reagent

consumptions are 0.39 kg/t for cyanide and 1.6 kg/t for lime.

Page 14.1




AMENDED NI 43-101 TECHNICAL REPORT 14.0 MINERAL RESOURCE ESTIMATES

The effective date of the mineral resource estimate is August 5, 2013. Open pit and underground

mineable resources were reported separately in a Technical Report prepared by MDA on September 10,

2013 (Lindholm and Schlitt, 2013b). The current resource estimate on which this Feasibility Study is

based is the same as the one reported in September 2013 but is entirely reported at an underground

mining cut-off grade of 1.2 g Au/t. The current resource estimate also includes a small addition on land

not formerly controlled by Red Eagle Mining.

14.1 Introduction

MDA classifies resources in order of increasing geological and quantitative confidence into Inferred,

Indicated, and Measured categories to be in compliance with the “CIM Definition Standards - For Mineral

Resources and Mineral Reserves” (2014) and therefore Canadian National Instrument 43-101. CIM

mineral resource definitions are given below, with CIM’s explanatory material shown in italics:

14.1.1 Mineral Resource

Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated

and Measured categories. An Inferred Mineral Resource has a lower level of confidence than that applied

to an Indicated Mineral Resource. An Indicated Mineral Resource has a higher level of confidence than

an Inferred Mineral Resource but has a lower level of confidence than a Measured Mineral Resource.

A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the

Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual

economic extraction.

The location, quantity, grade or quality, continuity and other geological characteristics of a Mineral

Resource are known, estimated or interpreted from specific geological evidence and knowledge, including

sampling.

Material of economic interest refers to diamonds, natural solid inorganic material, or natural solid fossilized

organic material including base and precious metals, coal, and industrial minerals.

The term Mineral Resource covers mineralization and natural material of intrinsic economic interest which

has been identified and estimated through exploration and sampling and within which Mineral Reserves

may subsequently be defined by the consideration and application of Modifying Factors. The phrase

‘reasonable prospects for eventual economic extraction’ implies a judgment by the Qualified Person in

respect of the technical and economic factors likely to influence the prospect of economic extraction. The

Qualified Person should consider and clearly state the basis for determining that the material has

reasonable prospects for eventual economic extraction. Assumptions should include estimates of cut-off

grade and geological continuity at the selected cut-off, metallurgical recovery, smelter payments,

commodity price or product value, mining and processing method and mining, processing and general and

administrative costs. The Qualified Person should state if the assessment is based on any direct evidence

and testing.

Page 14.2




AMENDED NI 43-101 TECHNICAL REPORT Interpretation of the word ‘eventual’ in this context may vary depending on the commodity or mineral

involved. For example, for some coal, iron, potash deposits and other bulk minerals or commodities, it

may be reasonable to envisage ‘eventual economic extraction’ as covering time periods in excess of 50

years. However, for many gold deposits, application of the concept would normally be restricted to

perhaps 10 to 15 years, and frequently to much shorter periods of time.

14.1.2 Inferred Mineral Resource

An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are

estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to

imply but not verify geological and grade or quality continuity.

An Inferred Mineral Resource has a lower level of confidence than that applying to an Indicated Mineral

Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of

Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.

An Inferred Mineral Resource is based on limited information and sampling gathered through appropriate

sampling techniques from locations such as outcrops, trenches, pits, workings and drill holes. Inferred

Mineral Resources must not be included in the economic analysis, production schedules, or estimated

mine life in publicly disclosed Pre-feasibility or Feasibility Studies, or in the Life of Mine plans and cash

flow models of developed mines. Inferred Mineral Resources can only be used in economic studies as

provided under NI43-101.

There may be circumstances, where appropriate sampling, testing, and other measurements are sufficient

to demonstrate data integrity, geological and grade/quality continuity of a Measured or Indicated Mineral

Resource, however, quality assurance and quality control, or other information may not meet all industry

norms for the disclosure of an Indicated or Measured Mineral Resource. Under these circumstances, it

may be reasonable for the Qualified Person to report an Inferred Mineral Resource if the Qualified Person

has taken steps to verify the information meets the requirements of an Inferred Mineral Resource.

14.1.3 Indicated Mineral Resource

An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality,

densities, shape and physical characteristics are estimated with sufficient confidence to allow the

application of Modifying Factors in sufficient detail to support mine planning and evaluation of the

economic viability of the deposit.

Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing

and is sufficient to assume geological and grade or quality continuity between points of observation.

An Indicated Mineral Resource has a lower level of confidence than that applying to a Measured Mineral

Resource and may only be converted to a Probable Mineral Reserve.

Page 14.3




AMENDED NI 43-101 TECHNICAL REPORT Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person when the

nature, quality, quantity and distribution of data are such as to allow confident interpretation of the

geological framework and to reasonably assume the continuity of mineralization. The Qualified Person

must recognize the importance of the Indicated Mineral Resource category to the advancement of the

feasibility of the project. An Indicated Mineral Resource estimate is of sufficient quality to support a Pre-

feasibility Study which can serve as the basis for major development decisions.

14.1.4 Measured Mineral Resource

A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality,

densities, shape, and physical characteristics are estimated with confidence sufficient to allow the

application of Modifying Factors to support detailed mine planning and final evaluation of the economic

viability of the deposit.

Geological evidence is derived from detailed and reliable exploration, sampling and testing and is

sufficient to confirm geological and grade or quality continuity between points of observation.

A Measured Mineral Resource has a higher level of confidence than that applying to either an Indicated

Mineral Resource or an Inferred Mineral Resource. It may be converted to a Proven Mineral Reserve or to

a Probable Mineral Reserve.

Mineralization or other natural material of economic interest may be classified as a Measured Mineral

Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such that

the tonnage and grade or quality of the mineralization can be estimated to within close limits and that

variation from the estimate would not significantly affect potential economic viability of the deposit. This

category requires a high level of confidence in, and understanding of, the geology and controls of the

mineral deposit.

14.1.5 Modifying Factors

Modifying Factors are considerations used to convert Mineral Resources to Mineral Reserves. These

include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing,

legal, environmental, social and governmental factors.

MDA reports resources at cut-offs that are reasonable for deposits of this nature given anticipated mining

methods and plant processing costs, while also considering economic conditions, because of the

regulatory requirements that a resource exists “in such form and quantity and of such a grade or quality

that it has reasonable prospects for economic extraction.”

Page 14.4




AMENDED NI 43-101 TECHNICAL REPORT 14.2 Database

The initial resource estimate for the San Ramon deposit (Lindholm and Schlitt, 2013a) utilized a database

with an effective date of November 21, 2012. Red Eagle Mining had completed 143 drill holes (SR-001 to

139, SR-028A, SR-032A, SR-045A and SR-050A) for a total of 23,015.25m of core in the San Ramon

deposit area. The San Ramon deposit database contained 11,844 records, of which 11,665 had gold

assays. Other elements, including silver, base metals, and trace elements, were assayed in SR-001 to

065. The database also contained logged lithology and recovery data, summarized redox data, and bulk

wet-density and specific gravity measurements. All of these drilling data were used in the initial estimate.

From November of 2012 through May 2013, Red Eagle Mining completed an additional 95 drill holes

(SR-140 to SR-233, SR-146B) for a total of 22,593.75 m of core. Of these, 76 were infill drill holes within

the previous resource area, and three were exploration holes drilled north of the San Ramon deposit

shear zone. Twelve holes extended mineralization down-dip within the shear zone, and four extended

mineralization to the west. Red Eagle Mining supplied the data for the new drilling to MDA on June 5,

2013, which is the effective date of the database for the updated resource estimate. The new collar,

survey, and assay data were audited and appended to the previously audited database. As noted in

Section 12.1.2, additional down-hole survey certificates were provided by Red Eagle Mining and were

used by MDA to audit some of the older survey data, which were modified as needed; none of the

previous collar and assay data were changed. Table 14.1 summarizes the data in the updated San

Ramon deposit database, which contains 19,938 records, of which 17,286 have gold assays. Only gold

was analyzed for the new drilling program. The other 2,652 records include blank records that are

automatically added by database software to indicate drill-hole gaps and intervals added for specific

gravity data. Red Eagle Mining also provided geological and geotechnical data for all drilling, including

major and minor lithologies, alteration, weathering, oxidation, structure, density, recovery, RQD and

hardness. Much of this information came from re-logging from core photos in an effort to standardize

geologic interpretations.

Page 14.5




AMENDED NI 43-101 TECHNICAL REPORT Table 14.1 Exploration and Resource Database Descriptive Statistics

Valid N Median Mean Std.Dev. CV Minimum Maximum Units

Length 19,938 1.00 2.29 0.01 205.35 m

Au 17,286 0.01 0.31 2.59 8.37 0.0025 157.70 g/t

Specific Gravity 932 2.61 2.32 0.53 0.23 0.97 3.45

14.3 Underground Workings

For the initial resource estimate, Red Eagle Mining provided MDA with digital outlines of 14 underground

workings in the San Ramon deposit. With average floor-to-back heights and floor elevations, MDA

constructed solids of the adits and shafts. These were used to code the block model by assigning

percentages of mined material to individual blocks, from which mined tonnes was calculated. The mined

tonnes were subtracted from the resource on a block-by-block basis. Red Eagle Mining is continually

discovering new underground workings, which are rehabilitated to the extent practical and sampled if they

intersect the shear zone. Although the total volume of mined material is likely very small, there is certainly

some unknown quantity of the resource that has already been mined, which is not taken into account in

the reported estimate.

14.4 Mineral Domains

Low- and higher-grade gold domains were modelled on north-south, 50m-spaced cross sections for the

resource at the San Ramon deposit. The lower boundary of the low-grade domain, which represents

mineralization in weak to moderately sheared rock that contains scattered small quartz veins, is at ~0.05 g

Au/t. The grade break between the low- and higher-grade domains is gradational between 0.35 and 0.6 g

Au/t. The higher-grade domain (>~0.6 g Au/t, but may contain grades down to 0.35 g Au/t) represents

zones of strong shearing that variably contain relatively thick quartz veins (less than 2 m) with coarse-

grained sulfides (sphalerite, pyrite, and galena), coarse-grained sulfide veins, finer-grained sulfide

minerals and quartz vein fragments in gouge zones, and sericite. The domains, and the primary shear

zone, strike roughly east-west and dip to the north at around 70° to near-vertical. At depths below an

elevation of approximately 2,100 m, the dip shallows to 50° to 60°. Figure 14.1 and Figure 14.2 show

these domains.

A quantile plot of coded higher-grade domain samples indicates the higher-grade domain is trimodal, with

one relatively sharp grade break at ~1.0 and another gradational break between 5.0 and 9.0 g Au/t. The

grade change at ~1.00 g Au/t is not abrupt, and no geological characteristics that would distinguish rock

above and below ~1.0 g Au/t were observed. Therefore the modelling did not separate these samples into

distinct domains. This is not considered a risk.

Page 14.6




AMENDED NI 43-101 TECHNICAL REPORT Above ~5.0 g Au/t, the samples commonly consist of quartz veins with coarse-grained sulfides. Gold

grade appears to increase with increasing content of sulfide minerals, and galena, abundant coarse-

grained sphalerite, and massive pyrite are generally associated with extremely high grades. Samples with

grades above ~5 g Au/t constitute only ~15 % of the higher-grade domain, but they are geologically

distinct and are treated during estimation as a sub-domain within the higher-grade domain. It was

observed that only about one third of samples with grades above ~5.0 g Au/t had similarly high-grade

samples from different drill holes at distances greater than 50 m. Consequently, at the existing 50 m drill-

hole spacing, the above ~5.0 g Au/t sub-domain within the higher-grade domain cannot, for the most part,

be modelled separately.

Like all deposits with distinct highest-grade populations whose continuity is generally less than existing

drill spacing, there is an added risk. This risk at the San Ramon deposit was lessened, if not removed, by

using estimation search restrictions above 9.0 g Au/t. To illustrate, about 8% of the metal at the 2.0 g Au/t

cut-off was lost between the 2012 estimate (Lindholm and Schlitt, 2013a) with drill spacing of ~100 m and

the 2013 estimate (Lindholm and Schlitt, 2013b) with drill spacing of ~50m. MDA expects that since the

drill spacing is approaching the average continuity of the highest-grade vein sub-domain, that the risk of

further losses is greatly diminished, and the magnitude of the losses, if they occur at all, would be

substantially less than 8%. Because the highest-grade mineralization occurs in zones less than 50 m, any

model from the current drill spacing will likely be a poor predictor of the location of the highest-grades,

although the overall estimated metal content is reasonable.

To further characterize the highest-grade mineralization, MDA identified a total of 98 sample intervals

within the modelled deposit. Core photos for 45 of these were examined, and various geological criteria

were summarized. Veins, sulfide minerals, and breccia make up ~50% of the total length of the samples.

Angles to core axes of mineralized zones are most commonly 40° to 60°, although structures were

intersected as shallow as 20° and as steep as 80°. Also, and most importantly, the mineralized structures

are sub-parallel to the surrounding shear-zone fabric in nearly all cases. This indicates that the

mineralization is strongly associated with the shear-zone and only infrequently occurs as cross-cutting

features.

Two extremely high-grade intercepts (SR-042, 184 m to 189 m and SR-053, 201.8 m to 211 m) deserve

special attention. Core photos indicate that the quartz and sulfide veins were intersected by these holes

at very shallow angles; in other words, the veins are sub-parallel to the core axis. The true widths of the

high-grade zones are, therefore, much narrower than the 9.2 m and 5 m intercepts, which was accounted

for in the modelling. Also, since the holes were drilled at a relatively high angle to the overall shear zone,

the near-parallel orientation of the veins in core might suggest a secondary orientation of the extreme

high-grade mineralization. However, core photos also show that the holes are sub-parallel to the

surrounding shear zone fabric, which may indicate that there is a local bend in the structure that hosts the

veins. These bends and associated very high-grade mineralization are likely localized and could easily be

missed at 50 m, or even tighter, drill-spacing. There is an unquantifiable potential to encounter similar

high-grade zones during mining that were not located by drilling.

Page 14.7




AMENDED NI 43-101 TECHNICAL REPORT The continuity of modelled individual zones within the higher-grade domain and the geologic and mineral

domain models, in general, were tested with the recent 50 m-spacing infill drilling program. The location

and presence of the shear zone and low-grade domain were confirmed within reasonable limits. In a

majority of cases, the location and presence of individual zones within the higher-grade domain were

validated as well, in the sense that strong shear zones with samples that assayed above ~0.6 to ~1 g Au/t

were intersected in the approximate locations indicated by the previous model. In fact, the continuity at

the underground resource cut-off and even potential underground mining cut-off grades is 100s of meters,

well beyond the existing drill density. In short, the model is a good predictor of the modelled domains.

Although there is some silver associated with gold in the San Ramon deposit, it was only analysed in the

first 65 holes drilled on the property. Therefore, it was not modelled, and is not included in the resource.

Page 14.8




AMENDED NI 43-101 TECHNICAL REPORT Figure 14.1 Gold Domains and Geology – Section 856500E

Reno

SCALE

Nevada

DATE

MINE DEVELOPMENT

ASSOCIATES

Lithology:Shear Zone

Gold Domains:low gradehigh grade

drill hole assays(Aucap)

< 0.001>= 0.001>= 0.040>= 0.050>= 0.100>= 0.250>= 0.350>= 0.600>= 2.000>= 3.500>= 5.000>= 8.000>= 30.000

shear zone

saprolite boundary saprock boundary

top of sulfide

bottom of oxide

GOLD V

EIN

FAU

LT

SAN

FRA

NC

ISCO

AD

IT FAU

LT

Gold Domains Section 856500 E

San Ramon Project

Red Eagle Mining

as shown

6 Aug 2013

Page 14.9




AMENDED NI 43-101 TECHNICAL REPORT Figure 14.2 Gold Domains and Geology – Section 857700E

Reno

SCALE

Nevada

DATE

MINE DEVELOPMENT

ASSOCIATES

Lithology:Shear Zone


Gold Domains Section 857700 E

San Ramon Project

Red Eagle Mining

as shown

6 Aug 2013

saprock boundary

top of sulfide

bottom of oxide

saprolite boundary

drill hole assays(Aucap)

< 0.001>= 0.001>= 0.040>= 0.050>= 0.100>= 0.250>= 0.350>= 0.600>= 2.000>= 3.500>= 5.000>= 8.000>= 30.000

< 0.001>= 0.001>= 0.040>= 0.050>= 0.100>= 0.250>= 0.350>= 0.600>= 2.000>= 3.500>= 5.000>= 8.000>= 30.000

Page 14.10




AMENDED NI 43-101 TECHNICAL REPORT 14.5 Density

In the early drilling programs (SR-001 to SR-139), Red Eagle Mining regularly measured core samples for

bulk wet densities at in-situ moisture conditions. At the request of MDA, Red Eagle Mining also measured

64 samples for bulk specific gravities for use in the initial estimate; these were taken from core holes

SR-001 to SR-010 and SR-059. The immersion method was used to obtain both sets of values. Wet-

density samples were wrapped in cellophane, and the weight of the displaced water determined for

calculations. Oven-dried bulk specific gravity samples were also wrapped in cellophane, but the weight of

the dry sample suspended in water was used for calculations per standard testing procedures for specific

gravity determination.

Frequent wet-density measurements were obtained for new drill holes SR-140 to SR-155. Starting with

drill hole SR-157 and continuing to the end of the 2013 program, 862 dry densities were measured, which

provided a total of 926 bulk specific gravity/dry-density values for use in the current estimate. The dry-

density procedure was identical to the immersion method used for wet-density determination, except an

oven-dried sample was tested. Proper drying times for various sample types, including saprolite, were

determined by weighing test samples every hour until a constant weight was achieved.

The combined specific gravity/dry-density data were evaluated in terms of lithology (shear zone, granite

country rock, dikes), redox (oxide, mixed, or sulfide) and saprolite. Table 14.2 summarizes density

measurements and the values applied to the model. Each data set represented in the table (i.e. All

Saprolite) was evaluated individually, and up to seven relatively extreme outliers (nearly all low) were

removed from each.

Table 14.2 Density Measurements and Values Applied to the Block Model

Rock Unit No. of

Samples Range Mean

%

Adjustment

Applied

SG

All Saprolite 229* 1.22-2.39 1.55 0% 1.55

All Saprock 92* 1.21-2.84 2.14 0% 2.14

All Oxide 259 1.21-2.70 1.60 0% 1.60

Transitional Redox – Shear Zone 3 1.83-2.77 2.41 1% 2.38

Transitional Redox – Country Rock 75 2.52-2.75 2.26 0% 2.26

Unoxidized – Shear Zone 152 2.23-3.45 2.68 1% 2.66

Unoxidized – Schist 17 2.52-2.98 2.75 0% 2.75

Unoxidized – All Other Lithologies 414 2.16-2.91 2.67 0% 2.67

All Data 932** 1.21-3.45 2.32 0% 2.32

* Saprolite and saprock values overlap with redox and lithology types

** No outliers removed

Page 14.11




AMENDED NI 43-101 TECHNICAL REPORT The specific gravity / dry-density data are sufficiently representative of the various lithology, redox, and

weathering types in the deposit, both in terms of quantity and spatial distribution. The only exception is

the transitional redox within the shear zone; however, this material makes up a very small portion of the

overall volume in the block model. The shear zone mean density values were adjusted down by one

percent, due to: 1) slightly lower core recoveries in the shear zone (generally by 2-4%), and 2) the

predominance of fractured and broken core, which is difficult to measure for specific gravity, thereby

instilling a sample selection bias against broken (lower density) samples.

14.6 Sample and Composites Descriptive Statistics

Once the domains were defined, the north-south sectional interpretations were used to code drill-hole

samples. Also, all samples outside modelled low- and higher-grade domains were assigned to a unique

domain for statistical and estimation purposes. Quantile plots were made; outlier grades were reviewed

on screen; and descriptive statistics were completed (Table 14.3). Samples from within each of the two

domains, as well as for assays outside modelled mineral domains, were capped. It is noteworthy that the

coefficients of variation (“CV”), in spite of apparently low capping grades, remain high (1.77 and 2.10 for

the low- and higher-grade domains, respectively). As noted earlier, the higher-grade domain appears to

be trimodal, but continuity that would allow for modelling of the separate highest-grade sub-domain was

not evident as its continuity is less than 50 m on average. There is no geological evidence that would

support modelling separate domains within the low-grade domain. As a consequence, and to compensate

for the shorter continuity, the projection distance of highest-grade assays was limited during the estimation

within each mineral domain.

Page 14.12




AMENDED NI 43-101 TECHNICAL REPORT Table 14.3 Descriptive Statistics of Coded Samples

Gold Domain 51 Low-grade; weak shear zone with quartz veinlets


Length 4,183 1 1.07

0.03 4.5 m

AuPPM 4,164 0.092 0.167 0.406 2.43 0.0025 15.1 g/t

AuCapped 4,164 0.092 0.161 0.285 1.77 0.0025 3.5 g/t

SG 87 2.68 2.65 0.11 0.04 2.25 2.86

Gold Domain 52 Higher-grade; strong shear zone, QV w/coarse-grained sulfides

Valid N Median Mean Std.Dev. CV Minimum Maximum Units

Length 1,225 1 1.01 0 0 0.1 3.96 m

Au 1,220 1.037 3.509 9.253 2.64 0.0025 157.7 g/t

AuCapped 1,220 1.037 3.257 6.848 2.10 0.0025 50 g/t

SG 50 2.72 2.70 0.33 0.12 1.39 3.45

Gold Domain 99 Outside domains


Length 14,530 1 2.75

0.01 205.35 m

Au 11,902 0.007 0.061 0.754 12.27 0.0025 57.3 g/t

AuCapped 11,902 0.007 0.044 0.261 5.89 0.0025 4 g/t

SG 795 2.58 2.27 0.55 0.24 0.97 3.21

QV = quartz veins

Capping for each domain was determined, first, by assessing the grade above which the outliers occur.

Then those outlier grades were reviewed on screen to determine materiality, grade and proximity of the

closest samples, and general location. Caps of 3.5 g Au/t, 50.0 g Au/t and 4.0 g Au/t were applied to the

low-grade, higher-grade, and outside domains, respectively. The cap applied to low-grade domain assays

is lower than capping of assays outside mineral domains; however, the influence of assays outside

domains is greatly restricted during the estimation process. Once the capping was completed, the drill

holes were down-hole composited to 2 m intervals, honouring the domain boundaries. The descriptive

statistics of the composite database are given in Table 14.4.

Page 14.13




AMENDED NI 43-101 TECHNICAL REPORT Table 14.4 Descriptive Statistics of Coded Composites

Gold Domain 51 Low-grade; weak shear zone with quartz veinlets


Length 2,566 2 1.73

0 2 m

AUPPM 2,556 0.106 0.167 0.273 1.64 0.0025 7.6 g/t

AUCAP 2,556 0.106 0.161 0.197 1.22 0.0025 3.5 g/t

Gold Domain 52 Higher-grade; strong shear zone, QV w/coarse-grained sulfides


Length 838 1.8 1.47

0 2 m

AUPPM 837 1.200 3.504 7.828 2.23 0.0025 157.7 g/t

AUCAP 837 1.200 3.253 5.706 1.75 0.0025 50 g/t

Gold Domain 99 Outside domains


Length 20,259 0 0.66

0 2 m

AUPPM 7,457 0.009 0.062 0.477 7.75 0.0025 23.5 g/t

AUCAP 7,457 0.009 0.044 0.183 4.12 0.0025 4 g/t

QV = quartz veins

14.7 Estimation

Four estimates were completed: polygonal, nearest neighbour, inverse distance, and kriged. These

estimates were run multiple times in order to evaluate the results and determine sensitivity to estimation

parameters. The inverse distance estimate is the reported estimate. Estimation parameters are given in

Table 14.5.

Two successive estimation passes were run for the low- and higher-grade domains. A first long pass

projecting 300 m along the primary axis in the low-grade domain, and 200 m in the higher-grade domain,

was run to fill in all blocks (all the blocks affected by this long pass were Inferred). This was followed by a

short pass of 100 m in both low- and higher-grade domains. Note that none of the highest-grade sub-

domain material would have been estimated beyond the 100 m pass and therefore does not exist at all in

the Inferred material. The search ellipse is oriented such that the major axis is along an east-west strike,

with a variable semi-major axis in the dip direction. Dips of 85°, 70°, and 55° were used above an

elevation of 2,400 m, between 2,115 m and 2,400 m, and below 2,115 m, respectively, to reflect the

changing orientation of the shear zone and mineralization at various depths.

Page 14.14




AMENDED NI 43-101 TECHNICAL REPORT Numerous test cases of the estimate at various parameters resulted in the selection of the estimation

parameters given in Table 14.5. Modifications made to the initial resource estimation parameters include

addition of multiple search dips for various elevation ranges and application of restriction grades of

0.7 and 9.0 g Au/t to low-grade and higher-grade domains, respectively (changed from 0.4 and 7.0 g Au/t).

Also, the inverse-distance power was increased from third to fourth; anisotropy was narrowed from 3:1 to

4:1; and the maximum number of composites from one drill hole was lowered from three to two for

estimation of higher-grade domains only (these original parameters for the low-grade domain were not

changed).

The highest grades (>5 to 9 g Au/t) at the San Ramon deposit have shown continuity of less than 50 m on

average, so MDA chose to restrict their impact on the estimate by limiting the projection of those high

values. For example, within the higher-grade domain, any composite with a grade over 9.0 g Au/t was

projected only 65 m of the full distance of long and short passes. The 5 to 9 g Au/t limit represents the

approximate lower end of the upper domain of the trimodal population (highest-grade sub-domain, not

modelled) and was chosen based on the quantile plot of the higher-grade domain samples. From the

quantile plot of the low-grade domain, it appears that the upper five percent of values are not part of the

same geological population, so a restriction of 0.7 g Au/t within 65 m for each pass was deemed

necessary. A restriction of 1.0 g Au/t within 10m for each pass was applied to blocks outside modelled

mineral domains.

The block model is not rotated, and the blocks are 2 m north-south by 2.5 m vertical by 2.5 m east-west.

The dimensions were chosen to best reflect possible block sizes for underground mining.

Page 14.15




AMENDED NI 43-101 TECHNICAL REPORT Table 14.5 Estimation Parameters

Description Parameter

Low-grade Domain

Samples: minimum/maximum/maximum per hole 1 / 12 / 3

Rotation/Dip/Tilt (variogram and searches):

Above elevation 2,400 m

Elevation 2,115 m to 2,400 m

Below elevation 2,115 m

90° / 0° /85°

90° / 0° /70°

90° / 0° /55°

Search (m): major/semi-major/minor (vertical)

Long Pass

Short Pass

300 / 300 / 100

100 / 100/ 33

Inverse distance power 3

High-grade restrictions (grade in g/t and distance in m)

Long Pass

Short Pass

0.7/ 65

0.7 / 65

Anisotropic weighting yes

Higher-grade Domain

Samples: minimum/maximum/maximum per hole 1 / 12 / 2

Rotation/Dip/Tilt (variogram and searches):

Above elevation 2,400 m

Elevation 2,115 m to 2,400 m

Below elevation 2,115 m

90° / 0° /85°

90° / 0° /70°

90° / 0° /55°

Search (m): major/semi-major/minor (vertical)

Long Pass

Short Pass

200 / 200 / 50

100 / 100/ 25


High-grade restrictions (grade in g/t and distance in m)

Long Pass

Short Pass

9.0 / 65

9.0 / 65


Outside Mineral Domains

Samples: minimum / maximum/maximum per hole 1 / 12 / 3

Rotation/Dip/Tilt (variogram and searches): 90° / 0° / 70°

Search (m): major/semi-major/minor (vertical) 100 / 100 / 33


High-grade restrictions (grade in g/t and distance in m) 0.1 / 10


Page 14.16




AMENDED NI 43-101 TECHNICAL REPORT 14.8 Mineral Resources

MDA classified the San Ramon deposit resources, giving consideration to a combination of distance to the

nearest sample, number of samples and holes, confidence in the underlying database, sample integrity,

analytical precision / reliability, and geologic interpretations. The criteria for resource classification are

given in Table 14.6. The classification of the resource was upgraded primarily on the basis of the

substantial amount of infill drilling, which brought the overall drill-spacing to ~50 m to a depth of about

200 m in the west half of the deposit and to a depth of about 250 m in the east half. Red Eagle Mining

also made significant improvements to their QA/QC program (i.e. real-time analyses and corrections),

collected much more dry-density and geotechnical data that sufficiently characterize all lithologic types,

improved understanding of the geology of the San Ramon deposit, and performed more comprehensive

metallurgical testwork. As a result, a Measured category was added, and the overall quantity of Measured

and Indicated material increased to about 75%. Measured material makes up approximately 20% of total

Measured and Indicated. The discontinuous nature of the highest-grade material and the demonstrated

natural heterogeneity of the gold in the deposit add some risk and decrease confidence in the resource

estimate somewhat. A delineation drilling program is included and discussed in the mining sections of this

Feasibility Study.

Distance parameters for material in intensely weathered rock were made stricter because core recoveries

are on average 17% less in saprolite and 10% in saprock relative to unweathered rock. Also, there is a

demonstrated decrease in grade at recoveries below ~70%, which may suggest a loss of gold in samples

in soft rock. These factors impart a sample bias and loss of confidence in the data for saprolite and

saprock.

Page 14.17




AMENDED NI 43-101 TECHNICAL REPORT Table 14.6 Classification Criteria

Measured

Inside Domains, in Saprolite,

And

No. of holes / closest distance / avg. distance of all

samples used >=4 / 10m from closest sample / <=40m

Inside Domains, Below Saprolite,

And

No. of holes / closest distance / avg. distance of all

samples used >=4 / 15m from closest sample / <=60m

Indicated

Inside Mineralized Domains, in Saprolite,

And

No. of holes / closest distance >=3 / 25m from closest sample

Or

No. of samples / closest distance >=2 / 15m from closest sample

Inside Mineralized Domains, Below Saprolite,

And

No. of holes / closest distance >=3 / 40m from closest sample

Or

No. of samples / closest distance >=2 / 25m from closest sample

Inferred

Inside any mineral domain that is not Indicated

Or

Outside the mineralized domains within 20m of a sample

Table 14.7 and Table 14.8 present the total Measured, Indicated, and Inferred block-diluted resources for

the San Ramon deposit. The initial resource estimate in 2012 (Lindholm and Schlitt, 2013a) was reported

at a single cut-off grade, because nearly all material had the potential to be mined by open-pit methods.

However, the mining scenario has now been determined to be by underground methods. The resource is

therefore reported at a grade of 1.2 g Au/t. The underground resource does not include material outside

modelled mineral domains, because it is unlikely to be mined from underground.

The reporting cut-off of 1.2 g Au/t is based on preliminary metallurgical testwork and operations cost

estimates for an underground operation. Preliminary metallurgical testwork indicated that whole-ore CIL

processing technology will potentially be suitable for all materials at the San Ramon deposit, with an

average gold recovery in excess of 95% weighted for all ore types. Metallurgical testwork has

demonstrated that all material, regardless of oxidation or weathering state, will be processed in a similar

manner.

Page 14.18




AMENDED NI 43-101 TECHNICAL REPORT Table 14.7 San Ramon Block-Diluted Gold Resources – Measured and Indicated

Total Measured & Indicated

Cut-off


1.00 4,861,000 3.23 504,500

1.20 4,153,000 3.59 479,000

1.40 3,612,000 3.93 456,800

1.60 3,179,000 4.26 435,800

1.80 2,825,000 4.58 416,100

2.00 2,523,000 4.91 398,400

3.00 1,549,000 6.47 322,000

4.00 1,035,000 7.97 265,100

Total Measured

Cut-off


1.000 775,000 3.873 96,500

1.200 678,000 4.270 93,000

1.400 600,000 4.657 89,800

1.600 537,000 5.027 86,800

1.800 488,000 5.362 84,100

2.000 444,000 5.706 81,400

3.000 298,000 7.306 70,000

4.000 218,000 8.725 61,100

Total Indicated

Cut-off


1.00 4,086,000 3.10 408,000

1.20 3,475,000 3.46 386,000

1.40 3,012,000 3.79 367,000

1.60 2,642,000 4.11 349,000

1.80 2,337,000 4.42 332,000

2.00 2,079,000 4.74 317,000

3.00 1,251,000 6.26 252,000

4.00 817,000 7.76 204,000

Page 14.19




AMENDED NI 43-101 TECHNICAL REPORT Table 14.8 San Ramon Block-Diluted Gold Resources – Inferred

Total Inferred

Cut-off


1.00 1,832,000 2.44 144,000

1.20 1,524,000 2.72 133,000

1.40 1,284,000 2.98 123,000

1.60 1,087,000 3.25 114,000

1.80 912,000 3.55 104,000

2.00 767,000 3.86 95,000

3.00 343,000 5.65 62,000

4.00 188,000 7.49 45,000

Page 14.20




AMENDED NI 43-101 TECHNICAL REPORT Figure 14.3 Gold Block Model Section 856500E

Reno

SCALE

Nevada

DATE

MINE DEVELOPMENT

ASSOCIATES

drill hole assays& model blocks(Aucap)


San Ramon Project

Red Eagle Mining

as shown

6 Aug 2013

Block Model Section 856500 E

< 0.040>= 0.040>= 0.050>= 0.100>= 0.250>= 0.350>= 0.600>= 2.000>= 3.500>= 5.000>= 8.000>= 30.000

Page 14.21




AMENDED NI 43-101 TECHNICAL REPORT Figure 14.4 Gold Block Model Section 857700E

Reno

SCALE

Nevada

DATE

MINE DEVELOPMENT

ASSOCIATES

< 0.040>= 0.040>= 0.050>= 0.100>= 0.250>= 0.350>= 0.600>= 2.000>= 3.500>= 5.000>= 8.000>= 30.000

San Ramon Project

Red Eagle Mining

as shown

6 Aug 2013

drill hole assays& model blocks(Aucap)

Block Model Section 857700 E


Page 14.22




AMENDED NI 43-101 TECHNICAL REPORT 14.9 Discussion of Resources

The San Ramon deposit is a steeply north-dipping, shear-zone-hosted gold deposit. New drilling since the

original resource estimate (Lindholm and Schlitt, 2013a) has identified extensions of mineralization that

project 100 m to 200 m deeper and 100 m farther to the west, so that the currently defined limits of the

deposit are roughly 2,000 m in an east-west direction and 350 to 550 m below the surface. A significant

outcome of Red Eagle Mining’s latest work has been improvement of an already good geologic model,

based on 95 new core holes (total 238) providing a solid base for the updated resource estimate.

About 75% of the total resource is classified as Measured and Indicated, with the remainder as Inferred.

Upgrade of the initial resource classification is a reflection of the substantial amount of infill drilling,

improvements in geologic understanding and QA/QC practices, and the addition of more comprehensive

density and geotechnical data and metallurgical testwork. The work done has shown a strong and

predictable mineralized shear zone and indications of varying styles of mineralization with distinctive

geological characteristics that can be modelled within the shear. The infill drilling to approximately 50 m

spacing was generally successful at verifying the modelled location of the shear zone and mineral

domains, although it was somewhat less able to verify the location and grade of mineralization greater

than ~5.0 g Au/t. Globally MDA has demonstrated that the metal content exists as stated. Additional

drilling would help to mitigate the risk associated with the local definition of this highest-grade style of

mineralization, which is the principal reason for the relatively small amount of Measured resources.

There is potential to encounter additional highest-grade quartz/sulfide zones during mining that were not

located by drilling, although this potential cannot be quantified. For example, even at drill spacing tighter

than 50 m, the localized bends and associated highest-grade mineralization similar to that encountered in

SR-042 and SR-053 could easily be missed.

Multiple test cases of the estimate at various capping grades, restrictions on interpolation distances at

varying grades, inverse-distance powers, anisotropies, and maximum number of composites per drill hole

were performed to determine the model’s sensitivity to these parameters. Also, multiple point-validation

(also called jackknifing) tests of inverse-distance to the first through fifth powers, as well as ordinary kriged

estimates, were performed for various parameters. Overall, these test cases confirmed that the average

continuity of the highest-grade mineralization is generally less than 50m, and estimation parameters were

chosen, based in part on the results, to restrict the projection of the highest grades. In other words, MDA

has accounted for the shorter continuity of the highest-grade mineralization in the estimate.

Increased drilling density will upgrade Inferred material into Indicated, and Indicated to Measured. Most of

the resources deeper than 200 m to 250 m below the surface are Inferred because of the wide drill

spacing. Drilling subsequent to the initial 2012 resource estimate has already added new Inferred material

to the resource, and the deposit is still open-ended down-plunge to the east onto newly acquired property.

Page 14.23




AMENDED NI 43-101 TECHNICAL REPORT The previous resource estimate update (September 10, 2013; Lindholm and Schlitt, 2013b) was reported

as a combination of open-pit-mineable material at a cut-off grade of 0.3 g Au/t and underground-mineable

material at a cut-off of 1.2 g Au/t. At that time, the total resource consisted of 10,348,000 tonnes of

Measured and Indicated material at 1.81 g Au/t (601,000 ounces of gold), and 2,966,000 tonnes of

Inferred material at 1.69 g Au/t (161,000 ounces of gold). Because the mining scenario has now been

determined to be by only underground methods, the current estimate’s tabulation applies underground

mining criteria, and the entire resource is reported at a cut-off grade of 1.2 g Au/t. Consequently, the

resource in this report cannot be directly compared to the September 10, 2013 resource estimate.

Changes to the initial estimated resource (January 22, 2013; Lindholm and Schlitt, 2013a) included

increases in the total resource at depth as a result of expansion drilling in 2013, but these increases were

offset by decreases due to lower applied densities in the saprolite, decreased influence of mineralization

~>5.0 g Au/t, and the change in reporting cut-off grades (i.e. in the 2013 estimate (Lindholm and Schlitt,

2013b,. underground resources were reported at 1.2 g Au/t below an elevation of 2,325 m at the west end

of the deposit and 2,110 m at the east end).

Page 15.1




AMENDED NI 43-101 TECHNICAL REPORT 15.0 MINERAL RESERVE ESTIMATES

15.1 Mineral Reserves

Mine Development Associated (“MDA”) classifies reserves in order of increasing confidence into Probable

and Proven categories to be in compliance with the “CIM Definition Standards - For Mineral Resources

and Mineral Reserves” (2014) and therefore Canadian National Instrument 43-101. CIM mineral reserve

definitions are given below, with CIM’s explanatory material shown in italics:

Mineral Reserve

Mineral Reserves are sub-divided in order of increasing confidence into Probable Mineral

Reserves and Proven Mineral Reserves. A Probable Mineral Reserve has a lower level of

confidence than a Proven Mineral Reserve.

A Mineral Reserve is the economically mineable part of a Measured and/or Indicated Mineral

Resource. It includes diluting materials and allowances for losses, which may occur when the

material is mined or extracted and is defined by studies at Pre-Feasibility or Feasibility level as

appropriate that include application of Modifying Factors. Such studies demonstrate that, at the

time of reporting, extraction could reasonably be justified.

The reference point at which Mineral Reserves are defined, usually the point where the ore is

delivered to the processing plant, must be stated. It is important that, in all situations where the

reference point is different, such as for a saleable product, a clarifying statement is included to

ensure that the reader is fully informed as to what is being reported.

The public disclosure of a Mineral Reserve must be demonstrated by a Pre-Feasibility Study or

Feasibility Study.

Mineral Reserves are those parts of Mineral Resources which, after the application of all

mining factors, result in an estimated tonnage and grade which, in the opinion of the

Qualified Person(s) making the estimates, is the basis of an economically viable project

after taking account of all relevant Modifying Factors. Mineral Reserves are inclusive of

diluting material that will be mined in conjunction with the Mineral Reserves and

delivered to the treatment plant or equivalent facility. The term ‘Mineral Reserve’ need

not necessarily signify that extraction facilities are in place or operative or that all

governmental approvals have been received. It does signify that there are reasonable

expectations of such approvals.

Page 15.2




AMENDED NI 43-101 TECHNICAL REPORT ‘Reference point’ refers to the mining or process point at which the Qualified Person

prepares a Mineral Reserve. For example, most metal deposits disclose mineral reserves

with a “ mill feed” reference point. In these cases, reserves are reported as mined ore

delivered to the plant and do not include reductions attributed to anticipated plant losses.

In contrast, coal reserves have traditionally been reported as tonnes of “ clean coal” . In

this coal example, reserves are reported as a “ saleable product” reference point and

include reductions for plant yield (recovery). The Qualified Person must clearly state the

‘reference point’ used in the Mineral Reserve estimate.

Probable Mineral Reserve

A Probable Mineral Reserve is the economically mineable part of an Indicated, and in some

circumstances, a Measured Mineral Resource. The confidence in the Modifying Factors

applying to a Probable Mineral Reserve is lower than that applying to a Proven Mineral Reserve.

A The Qualified Person(s) may elect, to convert Measured Mineral Resources to Probable

Mineral Reserves if the confidence in the Modifying Factors is lower than that applied to a

Proven Mineral Reserve. Probable Mineral Reserve estimates must be demonstrated to

be economic, at the time of reporting, by at least a Pre-Feasibility Study.

Proven Mineral Reserve

A Proven Mineral Reserve is the economically mineable part of a Measured Mineral Resource.

A Proven Mineral Reserve implies a high degree of confidence in the Modifying Factors.

Application of the Proven Mineral Reserve category implies that the Qualified Person has

the highest degree of confidence in the estimate with the consequent expectation in the

minds of the readers of the report. The term should be restricted to that part of the

deposit where production planning is taking place and for which any variation in the

estimate would not significantly affect the potential economic viability of the deposit.

Proven Mineral Reserve estimates must be demonstrated to be economic, at the time of

reporting, by at least a Pre-Feasibility Study. Within the CIM Definition standards the

term Proved Mineral Reserve is an equivalent term to a Proven Mineral Reserve.

Modifying Factors

Modifying Factors are considerations used to convert Mineral Resources to Mineral Reserves.

These include, but are not restricted to, mining, processing, metallurgical, infrastructure,

economic, marketing, legal, environmental, social and governmental factors.

Mineable reserves were developed using the resource modelled high-grade domains along with

undiluted grade estimates. The high-grade domains were used as a basis for stope designs. To

build up the mineable reserves, the following steps were performed:

Page 15.3




AMENDED NI 43-101 TECHNICAL REPORT Expand the high-grade resource estimation domain polygons to a minimum mining width

of 2.5 m resulting in mid-block, level-plan polygons that represent potential stopes;

o Initial designs were completed using 2.0 m minimum mining widths, and then

expanded to 2.5 m minimum mining width in order to include external dilution or

selvedge.

Estimate diluted stope grades using only Measured and Indicated undiluted grade

estimates from both high and low-grade estimation domains;

o Note there is additional dilution from included waste and Inferred resources

added at zero grade.

Apply economic parameters to calculate net value for each block;

Revise stope polygons to remove negative value blocks where possible;

o Blocks with negative values were retained where required to maintain

continuous blocks within a stope.

Resources inside of each stope polygon were summarized and the economic value for

each stope was calculated;

Stope polygons with negative values were either modified or eliminated;

Centerline development was refined to access each stoping area;

Stope economics were reassessed to include an allocation of development costs

required to mine each stope;

Final refinement or elimination of stope polygons was completed; and

Solids of the final stope polygons were created and resources inside of each solid were

summarized.

The stope solids reflect the material that is to be mined and sent to be processed. As such, all

Measured and Indicated resources inside of the stope solids, including resources below the

mining cut-off grades (oxide 1.96 g Au/t, transition 2.14 g Au/t, and sulfide 2.00 g Au/t) (internal

dilution), are considered to be Proven and Probable reserves respectively. Table 15.1 shows

the Proven and Probable reserves for the San Ramon deposit by material type.

Table 15.1 Total Proven and Probable Reserves by Material Type

Page 15.4





Reserves are reported based on Measured and Indicated resources inside of mining

shapes;

Measured and Indicated resources below the mining cut-off grade, above the resource

grade, and inside of mining solids are included as reserves; and

Rounding may result in apparent summation differences.

Table 15.2 shows the total Proven and Probable reserves along with additional dilution. This

additional dilution is waste and Inferred material inside of the stopes that cannot be segregated

out of the material sent to the mill. The waste and Inferred material is included in the mill feed at

zero grade, reducing the overall mill feed grade. The feasibility uses a fully diluted grade of 4.57

g Au/t for production scheduling and economic analysis.

Table 15.2 Total Proven and Probable Reserves and Dilution

Oxide Mixed Sulfide Total Proven and Probable

K Tonnes g Au/t K Ozs Au K Tonnes g Au/t K Ozs Au K Tonnes g Au/t K Ozs Au K Tonnes g Au/t K Ozs Au

Proven 11 4.60 2 3 3.12 0 415 5.99 80 429 5.93 82

Probable 63 3.91 8 17 4.37 2 1,915 5.08 313 1,995 5.04 323

Proven & Probable 74 4.01 10 20 4.20 3 2,331 5.24 393 2,425 5.20 405

K Tonnes g Au/t K Ozs Au

Proven 429 5.93 82

Probable 1,995 5.04 323

Proven & Probable 2,424 5.20 405

Dilution 334 - -

Page 15.5





Figure 15.1 Mineral Reserves

Page 15.6




AMENDED NI 43-101 TECHNICAL REPORT 15.2 Economic Parameters

After design, the economic viability of stope blocks was checked by estimating the value of each stope

block and then modifying or eliminating stope solids with zero or negative gross value. Table 15.3 shows

the economic parameters that were used to estimate the value of each stope block. Cut-off grades and

reserves have been stated using a $1,300 per ounce gold price. Note that these parameters were used

for defining stopes and differ slightly from the final economic analysis. During the feasibility study, the

LMM royalty increased to 3%.

Table 15.3 Economic Parameters

15.3 Dilution and Ore Loss

Initial designs were undertaken using a 2.0 m minimum mining width. Measured and Indicated material

inside of these designs that is not above the mining cut-off grade is added at the respective grade, while

Inferred and un-estimated material is added at zero grade with both being considered internal dilution.

The stope designs were later expanded to 2.5 m minimum widths to account for external wall dilution

(selvedge). This external dilution models what additional material may be captured in the stopes due to

over break from blasting and cleaning of the ribs and back.

Mining Costs

UG Mining Cost 37.36$ $/t Mined

Processing Costs

Milling 25.39$ $/t Processed

Refining 10.00$ $/oz Au Produced

Recoveries Oxide Mixed Sulfide

Milling - Au 95% 87% 93%

Payable 99.5% 99.5% 99.5%

Royalties

Government 3.20% nsr

Liberty Mutual Insurance 2% nsr

G&A Costs

G&A Per Year 3,535,200$

Throughput (t/day) 1,000

Days per Year 360

Throughput (t/year) 360,000

G&A $/t Processed 9.82$

Page 15.7




AMENDED NI 43-101 TECHNICAL REPORT Dilution by Measured and Indicated blocks included material below the mining cut-off grade, but above the

resource cut-off grades. This ensures that only metal from the officially reported resources contribute to

value generation in the cash-flow analysis. This dilution totals 188,000 tonnes of material grading 1.60 g

Au/t and is part of the Proven and Probable reserves.

The stope solids include material that is classified as Inferred resource or was not estimated, and while

the tonnage of this material is included as dilution, no additional metal content is included. This dilution

totals 334,000 tonnes of material at zero grade.

On an overall tonnage basis, the total dilution included is approximately 23%.

The San Ramon deposit is a high-grade narrow vein deposit. Many of the veins are narrower than the

resource model block width. During mining, it will be important to better define these veins through

delineation drilling prior to final planning and subsequent mining for each stope. This will allow adjustment

of the stopes to a minimum mining width around the veins to get the best value out of the ore and avoid

mining areas where dilution will destroy value. Thus, no additional external dilution has been added

beyond the 2.5 m minimum mining width.

Ore loss has been accounted for by removing areas that will not be mined because either they are too

remote from other potential ore to pay for additional development, or the potential value has been diluted

to a point where the material is eliminated from consideration. No other ore loss has been considered.

15.4 Reserve Comparison with the Preliminary Economic Assessment

The Feasibility Study builds on the concepts of the PEA Technical Report by MDA (Dyer, Lindholm,

Schlitt, Defilippi, 2014) which included Inferred resources. As such, a comparison between potential

mineable material from the PEA and the resulting fully diluted reserves is presented in Table 15.4 and

shown in Figure 15.2 and Figure 15.3. The PEA carried the prescriptive wording required for a PEA that

“ The Preliminary Economic Assessment is preliminary in nature. It includes Inferred mineral

resources that are considered too speculative geologically to have the economic considerations

applied to them that would enable them to be categorized as mineral reserves, and there is no

certainty that the Preliminary Economic Assessment will be realized.” This language is repeated

here with respect to the PEA potentially mineable material stated in tables below.

The diluted Proven and Probable reserves as presented herein are based on the same resources model

that was used in the PEA, with minor changes due to subsequent additional land to the east of concession

B7560005. The changes consider the original diluted material processed from the PEA, removal of

Inferred material, changes to design, and expansion of the minimum mining width to include external

dilution.

Page 15.8




AMENDED NI 43-101 TECHNICAL REPORT Table 15.4 Changes from PEA Material Processed to Reserves

Figure 15.2 Changes to Tonnage – PEA vs Reserves

Figure 15.3 Changes to Ounces of Gold – PEA vs Reserves

Tonnes g Au/t Ozs Au

PEA Material Processed 3,642,000 4.76 557,300

Remove Inferred (832,000) 4.16 (111,200)

Change due to Design (540,000) 4.26 (73,900)

Add External Dilution 489,000 2.09 32,900

Final Diluted Reserve 2,759,000 4.57 405,100

Page 16.1




AMENDED NI 43-101 TECHNICAL REPORT 16.0 MINING METHODS

The San Ramon deposit has been planned as an underground mining operation. The advantages of

underground mining include:

Underground mining helps to reduce the footprint of the mine and its environmental

impacts;

The San Ramon deposit is a high-grade, narrow vein deposit which is ideally suited to

underground mining methods which minimise dilution from the mining process; and

Underground selectivity will help to maximize run of mine (ROM) feed grades.

Red Eagle Mining has been in discussions with contract miners for both development and underground

production. Due to the experience and capabilities found in prospective Colombian contract miners, Red

Eagle Mining elected to operate the mine using a contract mining group. The philosophy is to issue a

contract for a five year term, with an option to extend. Red Eagle Mining will retain the option to take over

the contract at any time should it prefer. Contract miners in Colombia are very experienced at

underground development mining techniques, but lack the experience in production stoping. Red Eagle

Mining intends to manage the stope production operations with close management, and with the

introduction of an experienced miner training team during the first year of ore production to train the

Colombian contractor miners.

Red Eagle Mining has progressed the contract tendering process to two very experienced Colombian

mining contractors (both with similar tender rates), and is in the process of moving forward into detailed

contract negotiations. The rates presented by the short-listed contractors have been adopted in the costs

estimates for mine development and production.

The contract miners will be required to provide infrastructure for mining activities. The infrastructure is to

include workshops, warehouse, fuel supply, and transportation and off-site accommodations for their

employees. Red Eagle Mining will provide power and water to the portal and on site facilities.

Mining will commence with construction of a portal followed by development of a decline ramp and lateral

haulage levels to access the deposit. The primary mining method for the San Ramon deposit is

mechanized shrinkage with delayed fill (MSDF).

16.1 Underground Development

The mine will be accessed through a single portal entrance. Development will include construction of a

decline, main haulage drifts, and attack ramps. Underground ventilation will require vent shafts to the

surface along with raises between levels.

Page 16.2




AMENDED NI 43-101 TECHNICAL REPORT Development has been designed on the footwall side of the deposit. Cross-cuts will connect the

development with stope production locations and, in some circumstances, may be driven through the vein

beyond the deposit and into the hanging wall so that additional exploration and delineation drilling may be

undertaken from the hanging-wall side of the deposit.

Figure 16.1 and Figure 16.2 show the underground development design. The following sections discuss

the various components of the development.

Page 16.3





Figure 16.1 Underground Development – Long View

Page 16.4





Figure 16.2 Underground Development – Plan View

Page 16.5




AMENDED NI 43-101 TECHNICAL REPORT 16.1.2 Portal Construction

The portal is located in the footwall to the south, off the central area of the deposit. The portal coordinates

are: North 1223052, East 857019 and Elevation 2,468 m.

The portal location facilitates direct access to the deposit while providing the area necessary for locating

temporary and permanent facilities near the portal area. These facilities will service the underground

mine. It also provides a convenient location for transporting mined material to the adjacent processing

plant or to temporary ore and waste stockpiles.

Geotechnical Conditions

The portal and 160m of the decline will be developed in ground with different degrees of weathering. The

main geological structures of interest are the saprolite rock, and a schistose mylonite (shear zone) rock

characterized by both ductile, in the form of mylonite, and brittle deformation, in the form of breccia and

gouge zones as described in the geotechnical report (Golder, May 2014).

Three geotechnical core holes were drilled along the centerline projection of the decline: SR-GT-06, SR-

GT-07, and SR-GT-08 and one early exploration hole further to the west: SR-008. The geology was

interpreted based on core logs of these holes and is shown below in Figure 16.3.

Figure 16.3 Portal and Ramp Area Geology

Page 16.6




AMENDED NI 43-101 TECHNICAL REPORT The three geotechnical holes also provided additional information for rock quality determination (RQD)

which was used in the design of the portal and the transition area between the portal and the decline.

Hole SR-GT-06, drilled near the top of the hill approximately 50 m east of the decline bend shows

approximately 36 m of saprolite overlaying poor-quality schist. This hole was drilled to a depth of 60 m,

and does not reach the bottom elevation of the proposed decline. SR-GT-07 was drilled approximately 40

m west of the proposed portal. This hole shows about 12 m of saprolite, followed by 26 m of weathered

and fractured quartz-micaceous schist. Fresh granodiorite is encountered below a depth of 44 m. A

portion of the proposed decline is located along schist. Schist core recoveries are generally good (above

80%), but RQD values are between 10 and 20%.

In general, the expected geotechnical conditions in the area are favourable in terms of surface stability.

Steep, unsupported cuts in saprolite have been observed in the project area with minimal erosion or

stability issues. Saprolite in the area consists of a fine-grained silty soil, grading coarser with depth with

sporadic diorite blocks (spherical weathering). Shear resistance is generally high when unsaturated, thus

the steep cuts observed, but decreases as water content increases.

Gullies and rills may trigger when the saprolite is cut at angles below 50 degrees and no surface water

diversion channels are provided near the head of the cut. No landslides were observed in the portal pad

area (Figure 16.4). Crown ditches and meshed shotcrete are recommended to mitigate erosion problems

and control runoff into the pad area. Some bolts or anchors may be necessary depending on conditions

encountered during excavation.

Figure 16.4 Portal Pad Area

Portal and Initial Ramp Construction

The construction of the portal and the ramp will consist of 3 phases: Portal Area Excavation, Portal

Excavation, and Ramp Excavation.

Page 16.7




AMENDED NI 43-101 TECHNICAL REPORT Portal Area Excavation

The portal area excavation will provide a platform and the face for the portal excavation. The approximate

volume of excavation at the portal areas is 3,270 m3. The cut is approximately 37 m long and at this

distance it will provide a face that will be 10.7 m high at an 80° face angle. The exposed face and

surroundings will be stabilized with soil bolts and shotcrete. A long section view of the estimated

excavation in this area is shown in Figure 16.5.

Figure 16.5 Portal Area Excavation Long Section

4.5m

10.7m

10cm shotcrete

Crown Ditch

PORTAL CUT3.270 cubic meters

2.7 % gradient

37.2m

MINE DEVELOPMENT ASSOCIATES

22-Sep-2014

Scale: Portal Area Excavation Long Section

San Ramon Deposit


not to scale

Date:

Portal Excavation

Once the surface of the excavated area is stabilized, portal excavation will begin using a 4.5 m wide by

4.5 m height profile with a crowned back to reduce stresses at the back of the development.

Page 16.8




AMENDED NI 43-101 TECHNICAL REPORT When collaring the portal, reinforcement of the portal structure will be required. Crown bench and crown

face anchor bolts will be installed to provide the required support. Additional ground support will require

the installation of steel ribs and possibly a canopy system such as a headwall and two wing walls and a

buttress against the portal crown. A cross section of the portal preparation is shown in Figure 16.6.

Figure 16.6 Cross-section of Portal Excavation Area

Ramp Excavation

The ramp excavation is at a maximum nominal gradient of 14% (1 in 7) in the straight development, and a

maximum 11% (1 in 9) gradient in curve development. The portal structure will be extended into the

subsurface portion of the ramp with the initial 30 m requiring the canopy structure, including the steel ribs

spaced at 1 m. Figure 16.7 shows the area where the canopy structure and steel ribs will be required.

Page 16.9




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.7 Subsurface Portion of the Ramp

Saprolite

Schist

Anchor Bolts

Canopy

Reinforced Shotcrete

Crown Ditch

Steel Ribs(spaced at 1m)

10cm Shotcrete

4.5m

10.7

50.2m

SR-GT-07

19.0m

Ramp (15% max gradient)

21.9m

MINE DEVELOPMENT ASSOCIATES

22-Sep-2014

Scale: Subsurface Portion of the Ramp

San Ramon Deposit


not to scale

Date:

Beyond the initial 30 meters, reinforced shotcrete and bolts will be required until development in more

competent rock begins approximately 180 m down the ramp.

A long section of the first 200m is shown in Figure 16.8. The denser grid lines are shown where the ramp

curves. Figure 16.9 shows the topography along with the ramp centreline.

Page 16.10




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.8 Long Section of the Initial 200 m of the Ramp

Page 16.11




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.9 Ramp Centerline Projection to Topography

Typical Cross-Section of the Ramp

A typical cross-section of the ramp showing the location of utilities (ventilation duct, water supply and

dewatering lines, communications and power) and a low-profile underground truck (20 tonne capacity) is

shown below in Figure 16.10.

Page 16.12




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.10 Typical Cross-Section

16.1.3 Primary Development

The primary development extends from the portal entrance to the deposit and is a continuation of the

ramp excavation as previously described. The primary development has been designed using a main

ramp and sublevel ramps parallel to the deposit approximately 120 m into the footwall providing access to

main haulage drifts. The primary development gradient is designed at 14% where the development is

straight. In places where curves are required, the development is designed at a slope of 11%.

Design dimensions are 4.5 m wide by 4.5 m high, with a crowned profile to reduce stresses at the back of

the development. The dimensions are designed to accommodate haulage of material using 20-tonne-

class, underground, articulated haul trucks. Primary development will be finished with utility lines,

electrical cable, and ventilation ducting as required. A total of 4,400 m of main ramp and 11,100 m of

sublevel ramps have been designed for the life of the mine.

16.1.4 Haulage Drifts

Haulage drifts are developed approximately 100 m away from the deposit and are designed to use the

same profile as the primary development (4.5 m wide and 4.5 m high, crowned) in order to accommodate

all underground mining equipment. In general the haulage drifts are developed with a zero gradient, but in

some areas the haulage access is driven at grade to get from one attack ramp to another. This is done to

reduce the overall development for the deposit.

In total, 1,800 m of main haulage drifts have been designed for the life of the mine.

Page 16.13




AMENDED NI 43-101 TECHNICAL REPORT 16.1.5 Attack Ramps

Attack ramps have been designed to access the stopes. Many of these attack ramps connect to the

sublevel development. However, where the attack ramps are on roughly the same elevations, and can be

easily connected, haulage drifts are developed. The attack ramps will be driven at gradients required to

access the different levels of stope panels, but not to exceed 15% down to, and 17% upward to the stope.

Attack ramps have been designed with dimensions of 2.5 m wide by either 2.5 m or 5.0 m high, depending

on the height of the lifts to be used in the stope. Attack ramps will be mined using single boom jumbos

and smaller 1.5 cubic meter load-haul-dump (LHD) equipment. A total of 9,100 m of attack ramps have

been designed for the life of the mine.

16.1.6 Ventilation Shafts, Raises and Drifts

Ventilation shafts and raises have been designed and will be connected to primary development. Two

ventilation shafts have been incorporated into the design, and the additional ventilation raises connect to

various levels to provide additional ventilation. Ventilation shafts have been designed to be 3 m in

diameter, though the final diameter may change during engineering design, depending on final ventilation

requirements for the mine. A total of 800 m of ventilation raises have been designed for the life of the

mine.

Ventilation drifts connect the ventilation raises with the primary development. Ventilation drifts will have

dimensions of 4.5 m by 4.5 m with crowned backs. A total of 380 m of ventilation drifts have been

designed for the life of the mine.

16.1.7 Other Development

Additional development will be required for muck-bays, workshops, and drill stations. Ten-meter-long drill

stations will be required every 25 m along the decline (on both the main decline and sublevel ramps) from

approximately elevation 2,380 m to elevation 2,220 m. Initially these will be used for delineation drilling of

the deposit. After that, they will be used for passing areas, muck bays and safety bays. Additional muck

bays will be developed near the attack ramps as required.

16.2 Stoping Methods

The mining method selected for the San Ramon deposit is Mechanized Shrinkage with Delayed Fill

(“MSDF”). The method is similar to mechanized cut and fill but uses breast blasting of the back between

lifts. Then, rather than mucking of ore and backfilling immediately, the MSDF method leaves the ore in

place in the stope, only removing enough material from the stope to remove swell (similar to shrinkage

stoping but removing the swell from the top instead of the bottom). Access to the stope is provided by

establishing an attack ramp. For each lift, the attack ramp back is blasted to establish access for the

subsequent lift. Support is placed in the stope as needed during the drilling and blasting cycle of each lift.

Once enough lifts have been drilled and blasted, the ore will be mucked out completely. The last cycle of

mining for MSDF stopes is backfilling of the stope from the bottom up, progressively backfilling the attack

ramp as well.

Page 16.14




AMENDED NI 43-101 TECHNICAL REPORT Figures 16.11 through to Figure 16.17 show the sequence of activities which are described in the following

sections.

Figure 16.11 Mechanized Shrinkage with Delayed Fill: Development in Ore

Figure 16.12 Mechanized Shrinkage with Delayed Fill: First Lift

Figure 16.13 Mechanized Shrinkage with Delayed Fill: Second Lift

Page 16.15





Figure 16.14 Mechanized Shrinkage with Delayed Fill: Subsequent Lifts

Figure 16.15 Mechanized Shrinkage with Delayed Fill: Mucking

Figure 16.16 Mechanized Shrinkage with Delayed Fill: Backfill

Page 16.16





Figure 16.17 Mechanized Shrinkage with Delayed Fill: Completed Stope

16.2.2 Attack Ramp Development

To access the ore body, an attack ramp is driven perpendicular to the ore body at a 15% decline gradient.

The attack ramp standoff distance has been designed so that the total stope height can be 30 m high,

utilizing a 15% gradient downhill to the first lift and a 17% gradient up the last lift in the stope. This

standoff distance is variable in different areas as a function of the ore height and dip of the ore body, but

averages 90 m laterally from the ore body. At the bottom of the attack ramp, turns are developed into the

ore body in both directions where development in ore commences.

The initial attack ramp is completed using typical drifting methods. The back is supported using standard

support, and may require modification as experience in different areas is gained.

Subsequent lifts within the stopes require that the back of the attack ramp be slashed down to allow

access for each lift. This is done by drilling blast holes into the back at a 45 degree angle and then

blasting the back downward. The new back that is developed is supported as required to ensure stability

and safety. Some levelling of waste is required while supporting the back. Once sufficient support has

been completed, the remaining waste is mucked out of the attack ramp so that it will not dilute any ore.

Slashing of the back in the attack ramp is done after slashing of the back within the stope. The excess

swell of ore from the stope is used to backfill the access floor in the attack ramp so that during final

mucking of the stope, all material can be sent as ore to the process facility.

Some dilution will occur when waste is blasted down onto the ore backfilled in the attack drift, and care will

be required to ensure that this dilution is not excessive. Methods of controlling dilution may include

marking the interface between the ore and the waste by using either a geo-membrane or paint.

Most importantly, when the uppermost portion of the attack ramp is developed, the supports must be

carefully installed to ensure safe operations during the final mucking of the stope. The backfilling of attack

ramps will be done in the same manner as described in the backfill Section 16.2.6.

Page 16.17




AMENDED NI 43-101 TECHNICAL REPORT 16.2.3 Development in Ore

At the bottom of each attack ramp, a curve is developed in each direction with sufficient radius and width

to allow for equipment to access the stope. The development is then driven along the trend of the ore

body either to the end or up to 100 m in length, whichever is less, in each direction. The development in

ore is driven to a minimum width of 2.5 m and a height of the lift. Lift heights are anticipated to be 5 m

where possible, but may be mined at a height of 2.5 m where the higher lift would cause excessive

dilution. Over break of 0.5 m is included in the minimum width (0.25 m per side), so some widths may

vary; however, with experience, a minimum width of 2.5 m with the over break is expected to be

maintained without difficulty.

The entire length of the ore development is supported as required to safely enter underneath the

excavation. In some areas, this may require use of mesh, straps, and rock bolts based on ongoing

experience. Safety is of key concern, and support in the ore development drifts will only be needed for a

short time. Thoughtful design of the support mechanisms must satisfy safety needs and also minimize

scrap that will have to be removed from the ore prior to processing.

All muck generated from development in ore will be hauled to underground muck bays for truck loading or

directly to the surface for processing.

16.2.4 Rib Mining

Once ore development has been completed, the resulting drift is widened, as needed, to extract ore

remaining along the sides of the development. This will be done using slashing techniques. After

slashing, the material is cleaned away, and support is placed as required. All ore that is mined by

slashing is hauled to muck bays for truck loading or directly to the surface for processing.

16.2.5 Mining Additional Lifts

Once the bottom lift of a panel has been mined, additional lifts are blasted, first by slashing the back, then

supporting the new back to ensure safe access, and then mining of the ribs. The additional lifts are mined

leaving ore in place and mucking only what is required to relieve swell from the top of the stope muckpile

so that work can continue on the subsequent lift. The required tasks to mine each lift are described below.

Each lift will be mined in the same manner, with the exceptions of using more stringent support on the top

lift.

Slashing the Back

Slashing of the back will be completed by drilling holes at a 45 degree angle into the back and up into the

stope using a single boom jumbo. The holes are oriented toward the direction of the development. Where

the shot is to be initiated, holes should be angled slightly flatter to provide a sufficient breaking point.

Blast pattern designs will use between a 0.50 kg/t and 0.30 kg/t powder factor, depending on the

requirements for fragmentation and based on experience.

Page 16.18




AMENDED NI 43-101 TECHNICAL REPORT Support

Once the back has been slashed, then the muckpile along the top of the stope is levelled off as needed

and used as a platform to install support. Ore removed to level the floor will either be placed into the

attack ramp to act as a road way, or hauled to a muck bay, or to the surface. The remainder of the ore is

left in the stope, which will provide for additional support of hanging wall and footwall until final mucking of

ore from the stope.

Rib Mining

Rib mining will be completed in the same manner as described for the initial lift.

16.2.6 Mucking

Mucking is undertaken once enough lifts have been mined. The broken ore that has been left behind is

mined out using LHD equipment. The ore is trammed to a nearby muck bay for loading onto underground

haul trucks. The muck will be removed in lifts that have similar heights as the lifts that were mined,

although some areas may be mucked on higher lifts, if appropriate. After and/or during the mucking of

each lift, additional support will be placed in the hanging wall and footwall of the stope as needed,

providing additional safety. Mucking continues until the stope is empty of ore.

Mucking is intended to be started after 15 m of the stope has been mined (after six 2.5 m lifts or three

5.0 m lifts). The height is determined by the geotechnical stability of the stope, and will come from both

experience and geotechnical analysis of core from the delineation drilling program. During mucking

operations, it will be important to monitor the stability of the stope while ore is removed. Should the stope

show signs of becoming unstable, it may be necessary to use remote-controlled LHD’s to complete the

mucking operations. The mining contractor will be required to have at least one LHD with remote control

capabilities available at all times.

16.2.7 Backfill

Backfill is to be used as required, but not all stopes will be fully backfilled. The backfill will consist of

filtered tailings and development waste rock. Tailings will be placed in a single lift extending from the far

end of the stope inward to the attack ramp. Once the lift of tailings is placed, then a lift of waste rock from

development will be placed on top of the tailings. The waste is placed starting from the attack ramp and

working back to the far end. The amount of waste placed on top of the tailings will be sufficient to allow

equipment to drive on top of the backfill. It has been assumed that 60% tailings and 40% development

waste rock will be a suitable mixture of tailings to waste to stabilize the fill.

Page 16.19




AMENDED NI 43-101 TECHNICAL REPORT MSDF is a bottom-up mining method; however, to avoid having to develop to the bottom of the ore body

before production mining can occur, some mining may be done starting at the middle of the deposit. In

order to recover material from below a stope mined in the middle of the deposit, it is planned to provide an

artificial crown pillar with a layer of cemented rock fill (CRF). CRF will be produced by adding cement

slurry to mine development waste rock in a purpose built sump, and mixing the blend using an LHD and

then trammed to the stope. Mining costs for San Ramon have assumed that 2.5 m of CRF with 5%

cement will be used in these cases so that more complete extraction of ore can be obtained. However,

tests incorporating 3% cement mixed with the tailings showed very good strengths (1,500 kPa), and

should be adequate for CRF. The CRF will become an artificial crown pillar which can be undermined by

stopes below it.

A proper mixture of CRF will depend on blending fine and course waste in proper proportion with the

cement slurry. In some cases the addition of some tailings material may take the place of the finer waste

material, however, this will require additional study.

One of the benefits of the MSDF method is that it reduces dilution and ore loss when mining subsequent

lifts. Once a stope level is completed, a new level is mined above the previous one. The ore from the

new level will be dropped on top of the backfill in the old stope level below. To control the amount of

dilution and ore loss, a firmer layer of backfill will be required in the mined-out stope. For this reason, the

upper lift of a level will have additional cement added to the backfill. This will provide a smoother mucking

surface and reduce dilution and ore loss when subsequent ore is mined above the backfill. This study

assumes that six 2.5 m lifts are mined prior to backfill, leaving the maximum height exposure within the

stope at 15 m. For costing purpose, the upper most 2.5 m of each 15 m lift has been assumed to require

1% cement to help and prevent dilution and ore loss. Thus, the addition of 1% cement is used for 16.7%

of the backfill volume. The cement will be mixed with the tailings in sumps similar to the mixing of CRF.

Recent tests to investigate the use of cement mixed with the tailings has indicated that a 1% cement

addition to the top of the tailings backfill could strengthen (to approximately 200 kPa) the surface for

equipment travel. The addition of waste rock will also improve the strength to facilitate equipment travel

and control dilution.

The final upper lift of stopes that are mined to the uppermost level of the ore deposit will likely be left

empty, as long as this does not impact the stability of the mine. Stopes that contain these voids will

require monitoring from time to time through the life of the mine to ensure that they are not becoming

unstable and that they are not causing instability to occur in other nearby mining areas.

16.2.8 Overall Stope Geometry

In general, the stopes will be developed in levels that are 30 m in height off of each attack ramp. The

height they are mined prior to mucking has been planned to be 15 m, although this is dependent on the

continuity of ore and stability of the stope. Stopes that can maintain a higher height will be mined

accordingly, particularly as experience is gained.

Page 16.20




AMENDED NI 43-101 TECHNICAL REPORT The minimum mining width used for design is 2.5 m, which accounts for a 0.5 m over break. The average

overall mining width is 3.0 m. Stope lengths will be kept to 100 m maximum on either side of the attack

ramp to reduce tramming distance within the stopes. Some areas may only allow a shorter distance

based on geotechnical considerations, or continuity of ore. Most stopes have been designed to be less

than 100 m long based on ore continuity and to optimize the use of the attack ramps.

Stope widths have been designed to encompass the width of the ore body. The current understanding of

the ore body indicates that only 4% of the stopes will need to be wider than 5 m. Stopes that are wider

than 5 m require additional support while mining (see Section 16.2.8 below).

16.2.9 Modified MSDF for Wide Stopes

In areas where the ore deposit is wider than the average 3.0 m width, a modification to the MSFD method

will be used. In this situation, the MSDF stope would be mined at a comfortable minimum mining width

leaving a portion of the ore in either the hanging wall or footwall behind as each lift is mined. Then during

the mucking process, the ore would be drilled and blasted in benches prior to the mucking. In this case,

additional support will need to be installed on the wall that is blasted.

The preference is to initially leave ore on the footwall side, which is thought to be of poorer quality in

places than rock on the hanging wall. Additionally, mining the hanging wall first allows for careful

installation of support along the hanging wall side, so that material would be less likely to fall into mined

areas.

16.2.10 Modified MSDF for Multiple Veins

Where multiple parallel veins are present, the veins will need to be mined somewhat simultaneously so

that the same attack ramp can be used. This will avoid repeated backfilling of the attack ramp that would

be required if the veins were mined separately. Where possible, it will be desirable to mine the hanging

wall vein first, to minimize the amount of stress within the overall hanging wall. When mining multiple

veins in this manner, special care should be taken to ensure stope stability. This may require that only a

couple of lifts can be mined prior to backfilling.

In order to mine veins and leave a pillar between them, the separation between the veins has to be

sufficient to maintain stability of the pillar. It has been assumed that the minimum distance between veins

is 2.0 m. In cases of less than 2.0 m of separation, the stopes will either be joined, or one or more of the

veins are excluded from the stope design and are not recovered.

16.2.11 Benefits of MSDF

The benefits to using MSDF instead of cut-and-fill mining are:

Because ore is being slashed from the stope back and dropped on ore, dilution and ore

loss during mucking are greatly reduced;

Page 16.21




AMENDED NI 43-101 TECHNICAL REPORT Because MSDF slashing techniques are used for the bulk of the stope, the powder




If development of MSDF stopes is maintained ahead of ore requirements, underground

stockpiles of ore will be available that can be delivered as required to the mill;

Most of the fill only requires enough strength for equipment to be operated on top of it to

deliver more fill, which reduces cement requirements and costs; and

The method can easily be converted to cut-and-fill techniques where ground conditions

become weak.

16.3 Ventilation

A preliminary ventilation system has been configured by dividing mine development and production into

three main phases. At the start of production, the ventilation system will consist of one intake and one

exhaust opening. Later in the mine life, ventilation will be expanded to two intakes and one exhaust

openings for the remainder of the life of the mine. The ultimate configuration relies on the portal and the

far east ventilation shaft to be used as main air ventilation intakes, and the western ventilation shaft (near

the center of the deposit) will serve as an exhaust. In the event of a fire deeper in the mine, this allows for

the main ramp to act as a safe egress into fresh air. Likewise, egress into fresh air can be made into the

most eastern ventilation shaft.

At the start of construction, air will be forced into the main ramp using auxiliary fans located on the surface

near the portal. The decline will be furnished with 1 m diameter ventilation ducting to move the air down

the decline. Ventilation Phase 1 for production will start once the first ventilation shaft is commissioned.

At that point, fans installed at the surface of the ventilation shaft exhaust air out of the mine pulling fresh

air into the mine through the portal. Two 200 kW fans with the ability to exhaust 250,000 cfm of air at a

pressure of 1.6 inches of water will be installed in parallel in the western ventilation shaft. These fans will

be operated so that one is a backup to the other to allow the ability to provide ventilation 100% of the time.

Once the primary exhaust shaft is operational, the ventilation ducting will be removed from the main

decline and salvaged for use in additional development. A 75 kW auxiliary fan will be placed near the

base of the exhaust shaft to supply air through the 1 m diameter salvaged ventilation ducting for additional

development. At the start of production, additional 75 kW auxiliary fans will be placed in the main decline

near production areas and ventilation will be forced into production areas using ventilation ducting.

Once the second ventilation shaft is commissioned, Phase 2 of the ventilation system will be in place.

This ventilation shaft will provide inflow into the eastern portion of the mine. Ventilation doors will be

placed between the Phase 1 and Phase 2 ventilation system to control the amount of air drawn in from

each of the intake (the main portal and the eastern ventilation shaft). As additional levels of the mine are

developed, ventilation raises will exhaust air from the mine.

Page 16.22




AMENDED NI 43-101 TECHNICAL REPORT Auxiliary fans will provide ventilation to development and production areas as needed through the life of

mine. Air will be blown into the production areas using 0.6 m diameter ventilation ducting. Ventilation will

also be regulated using air doors as required to provide a positive pressure in the direction of desired air

flow.

Figure 16.18, Figure 16.19, and Figure 16.20 shows the conceptual ventilation designs for Phase 1, 2,

and the end of the mine life.

Page 16.23





Figure 16.18 Ventilation Conceptual Design – Phase I

Page 16.24





Figure 16.19 Ventilation Conceptual Design – Phase II

Page 16.25





Figure 16.20 Ventilation Conceptual Design – End of Mine Life

Page 16.26




AMENDED NI 43-101 TECHNICAL REPORT 16.3.2 Equipment Loads

A design value of 0.06 m³ /s/kW was used to calculate the required airflow quantity for dilution of diesel

exhaust and other contaminants in the mine. Typically, haul trucks and LHD’s have the highest horse

power and operating hours, and therefore the greatest diesel emissions. Thus a high utilization factor was

used to calculate the requirements due do these two equipment types. An equipment load summary is

presented below in Table 16.1.

Table 16.1 Equipment Loads

16.3.3 Airflow Velocity

In the quantity of airflow required, airflow velocities were also considered. Excessive air velocities not only

create problems with dust control, but might also cause discomfort to mine personnel. Recommended

maximum airflow velocities are shown in Table 16.2.

Table 16.2 Recommended Maximum Airflow Velocities

16.3.4 Surface Fans

Fan operating requirements for the Santa Rosa Gold Project are provided in Table 16.3. Fan silencing

may be a consideration, as will baffles to precipitate wet particulate matter from the exhaust air. This can

be addressed with a 90 degree bend and the erection of cladding around the system. The fan operating

requirements include estimated pressure losses in the mine up to the ventilation raise collar. The fans will

be fitted with Variable Frequency Drives (VFD). VFD installations will allow airflow to be reduced,

depending on the ventilation needs of the mine, and thus input power can also be reduced to the fans.

Location / Phase Airflow (m3/s)

Development Production Sublevel Ramp Haulage Route Total

Phase 1 20 20 10 25 75

Phase 2 10 40 20 25 95

Phase 3 10 40 20 40 110

Airway Velocity Velocity

(m/s) (fpm)

Ventilation Shafts 20 4,000

Main Haulage Routs 6 1,200

Working Faces 4 800

Page 16.27




AMENDED NI 43-101 TECHNICAL REPORT Table 16.3 Estimated Fan Operating Points (Calculated at Ventilation Raise Collar)

Ventilation will be regulated to adjust the amount of air delivered to working areas using bulkheads or air

doors, as well as auxiliary fans, as required.

16.4 Geotechnical Studies

A study on the RQD for underground mining was undertaken by Golder in April, 2014. This study included:

Definition of rock domains based on available lithology descriptions in the core drill log

database;

Data analysis and filtering of the existing database;

o RQD values associated with core run recoveries less than 80% were removed

from the analysis;

o Drill holes more than 100 m away from the shear zone and proposed mine

development were removed from the analysis.

Statistical analysis of the RQD values for each rock domain, estimation of representative

RQDs and correlation to rock quality for generation of support and stand up time

recommendations; and Analysis of tele-viewer data for assessment of principal

discontinuity sets for each domain.

For estimation of rock support requirements and stand up times the Barton (1976) Q-method was

employed. The results from the analyses are shown in on the following Figure 16.21.

Phase Fan Description Quantity Pressure Input Power Location Fan Mode

(m3/s) (kPa)* (kW)+

Phase 1 Fan 1 75 0.21 23.7 Western Vent Shaft Exhaust



* At standard density 0.075 lb/ft3 (1.200kg/m3)+ At 65% efficiency

Page 16.28




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.21 Barton Q Characterization for San Ramon RQD

The intermediate zone is defined as parts of the shear zone that contain a more consolidated rock matrix,

as inspected in the drill core.

The results were checked by using an alternative rock mass rating system - Bieniawski (1976), and the

results are depicted in the following Figure 16.22.

0

20

40

60

80

100

120

140

160

0 20 40 60 80 100

Freq

uen

cy

RQD (%)

Hanging Wall

Shear Zone

Foot Wall

Intermediate

Page 16.29




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.22 Bieniawski RMR Classification for San Ramon RQD

The Barton results show very close similarity with the Bieniawski ratings, and confirm the classifications as

analysed by Golder.

Further, utilizing Bieniawski, the stand-up times for unsupported spans for each rock type are shown in the

following Figure 16.23.

Page 16.30




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.23 Bieniawski Span Tolerances for San Ramon

The conclusions to this study were that the quality for the hanging and foot walls is expected to be good to

very good with minimal support requirements for 20 m spans or less. Occasional support in the form of

bolting and shotcrete may be needed, particularly when the openings encounter a dyke or weak zone.

Conversely, for development along the poor quality shear zone, bolting and reinforcement will be required

at all stages of development.

16.5 Hydrological Studies

In 2013 Red Eagle Mining carried out a hydrogeological study for the Environmental Impact Assessment

(EIA) submission, which included the drilling of, and equipping of piezometers in monitor holes to provide

the necessary data for groundwater studies.

The map in Figure 16.24 shows the water wells and monitor hole locations. The hole collar locations are

shown in Table 16.4.

Figure 16.25 and Figure 16.26 show the drilling of the water wells.

ROCK MASS CLASSIFICATION

HANGING AND

FOOT WALLS

INTERMEDIATE

SHEAR ZONE

Page 16.31




AMENDED NI 43-101 TECHNICAL REPORT As an addition to the main hydrogeological program, and as part of the Feasibility Study investigations, a

site was selected to the north west of the proposed process plant site in a known fracture zone within the

granodiorite country rock (a spring is visible in that area), to establish a water well. An adjacent

piezometer hole was also drilled to obtain information on the drawdown and recharge of the water well.

The water well (W-2) was drilled through the saprolite cover (approximately 50 m deep) into the underlying

granodiorite to a final depth of 75 m, to allow extraction of water from the expected fractured bedrock. The

adjacent piezometer hole (PZ-8) was drilled to 60 m. The holes were both drilled vertically.

From inspection and flow tests made using air lift techniques, it was clearly evident that the hole had

insufficient yield to allow the installation of a pump. The decision was made to maintain the hole for

monitoring purposes use only.

In 2014 Red Eagle Mining carried out a geotechnical and hydrogeological drilling program for the

Feasibility Study. This program was designed and supervised by Golder as part of their mine wide water

balance and geotechnical study. Included in the drilling program was a water well that was drilled directly

into the shear zone with data capture planned to establish the anticipated inflows to be encountered

during mining operations. A monitoring hole was drilled adjacent to the water well and equipped with a

piezometer (Hole 5) to capture drawdown and recharge rate data.

Figure 16.24 Geotechnical and Hydrological Drill Program Locations

Page 16.32




AMENDED NI 43-101 TECHNICAL REPORT Table 16.4 Hole Collar Locations

Hole Ref: Easting Northing Elevation Depth

PZ-8 856,788 1223,504 2,482 60 m

W-2 856,764 1223,501 2,486 75 m

15 857,279 1223,205 2,470 172.60 m

5 857,250 1223,189 2,465 80.75 m

Figure 16.25 Drilling of Piezometer Pole PZ-8 (distant) and Water Well W-2 (foreground)

Page 16.33




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.26 Drilling of Water Well 15

The design of Hole 15 was based on the existing geological information obtained from exploration holes in

the vicinity and Red Eagle Mining’s detailed knowledge of the shear zone deposit. Figure 16.27 shows

the geological section for hole 15 and other drill holes from the exploration program in that area. The

Water Well (15) was drilled vertically through the saprolite cover and shear zone and into the footwall

granodiorite to a final depth of 172.6 m. This provided sufficient depth below the shear zone for a sump to

be established and a pump was planned to be installed at this depth. The hole was core drilled to HQ size

and then reamed to a final 12.25'' diameter. The hole was then cased with 8'' steel pipe casing, gravel

packed and perforated steel screened sections 6 m long, installed in the borehole through the shear zone.

This casing diameter allowed for a 6'' diameter submersible pump which was installed for flow tests.

Hole 5 was drilled vertically to a depth of 80.75 m and equipped with a piezometer.

Page 16.34




AMENDED NI 43-101 TECHNICAL REPORT Figure 16.27 Geological Section Through Hole 15

Pump tests were carried out late July 2014, a pump was selected with sufficient power to overcome the

static head of approximately 150 m, this displaced the water quickly and there was too little recharge in

the hole to enable a steady flow to be obtained and measured. Consequently the equilibrium conditions

necessary for steady flow rate measurement was not able to be achieved.

The results of several test runs established that for the water level to recharge in the hole with a 15 m

increase in elevation, took three to seven hours. This equated to extremely low inflows ranging from 2.5 to

1.0 L/min. This shows that the water in the shear zone is held as natural porosity and that the hydraulic

conductivity is very low. The pump tests were simply draining the water from the shear zone in the

immediate vicinity of the borehole over an ever increasing distance from the hole, hence the reducing

flows. Measurements taken from the adjacent piezometer Hole 5 showed no change in water levels and

confirmed the low hydraulic conductivity.

Golder concluded that the test data did not allow for a full interpretation, but values ranging from 0.2 to 2

L/sec can be expected inflow rates.

Page 16.35




AMENDED NI 43-101 TECHNICAL REPORT The monitoring of water pumped from the existing Hilo Azul shaft (40 m deep) and underground workings

was undertaken in parallel to the hydrogeological and geotechnical programs. A submersible pump

discharged at surface and flow measurements are taken using a drop test. The result was a flow of 6

L/sec, this reduces over time as the water level in the workings and shaft is reduced. The discharge water

from Hilo Azul has a neutral pH ranging from 6 to 7.

With the information gained from the two water wells and Hilo Azul, this confirms that the granodiorite

country rock yields insignificant flows and has low conductivity, and the shear zone similarly yields low

flows and has low conductivity. The mine design for pumping has been conservatively sized for an initial

inflow of 5 L/sec., rising to 10 L/sec as the operations increase with depth. The mining contractors who

have submitted bids have also used similar inflow rates in pricing the development and production costs.

16.6 Development and Production Scheduling

Mine development and production were scheduled using Geovia’s MineSchedTM software (version

8.0.1.0.132). Development and production locations and rates were input based on the anticipated ramp-

up schedule and coordinated with bidding contractors.

The schedule includes a one year pre-production period. During that time, the portal and main decline will

be developed to the bottom of the primary exhaust ventilation shaft. At the same time, sublevel ramps

and haulage drifts will be developed as areas become accessible. Development of the first few stopes is

anticipated in the 12th month of construction, which will allow stockpiling of ore so that production can

begin in the 13th month. This fits with the planned process plant construction schedule, with

commissioning in the 13th month.

Production will start in Year 1, focusing on high-grade zones. Production will ramp up relatively quickly,

allowing the processing of 1,000 tonnes per day in Year 1. The following subsections describe the mine

development and mine production schedules.

16.6.1 Mine Development Schedule

Development plan centerlines and development rates were entered into Geovia’s MineSched software.

Precedents were used to ensure that development was completed in a sequential manner. The centerline

design is shown in Figure 16.1 and Figure 16.2.

The development schedule allows for construction of the portal and the first 7.75 m of underground

development in the first month of project construction (month -12). The second month of construction

(month -11) uses an advance rate of 1.5 m/day, estimated for the slow advance of development through

the weathered saprolite and schist. As more acceptable rock is encountered the rate is increased to a

maximum of 4.0 m/day. As areas for secondary development become available (haulage drifts and attack

ramps), additional development crews and equipment will be added to increase the overall development

rate. The maximum development rate for secondary development is assumed to be 9 m/day. This is

increased later in the mine life to 12 m/day, which assume an additional crew will be available at least part

time. The advance rates are adjusted to the amount of development required and are significantly lower

than those proposed by the mining contractors.

Page 16.36




AMENDED NI 43-101 TECHNICAL REPORT The attack ramps will have a 2.5 m wide by either 2.5 m high or 5.0 m high mining profile. Since these

attack ramps are much narrower than other development, a rate of 3 m/day is used. Ventilation drifts are

also developed at a rate of 3 m/day, and ventilation raises and shafts assume a 4 m/day mining rate, and

assume development using raise bore machines or Alimak raise construction methods. The mining

contractors have the capability and experience for both methods.

Table 16.5 shows the development schedule through the life of the mine.

Table 16.5 Yearly Development Schedule

16.6.2 Mine Production Schedule

Mine production was scheduled along with the development using Geovia’s MineSched software. Attack

ramps were used as precedents to ensure that required development was completed before a mining

location was scheduled. A total of 106 mining locations were defined. Each were setup to mine from the

bottom upward to represent the proper mining sequence.

Production rates were ramped up to 1,000 tonnes per day during the pre-production Year -1. A calendar

of 360 days per year was applied, making the annual production rate 360,000 per year. Each mining

stope was limited to 200 tonnes per day production, thus requiring five stopes to be producing in order to

fulfil the 1,000 tonne per day production target. With on-going backfill operations, this means that up to

ten locations may be active at one time (five in production and up to five being backfilled). Table 16.6

shows the yearly mine production. Stopes were ranked based on overall stope gold grade. Priorities were

applied based on the ranking providing higher-grade stopes to have priority over lower-grade stopes. It

should be noted that oxide and mixed (transition) ore is included in the schedule as the categories are

drawn for the resource model. The quantities, however, are very small and will not affect the flow of

production and processing.

Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Portal (Portal costs are in Capital) m 30 - - - - - - - - 30

Sublevel Ramp m 484 1,810 2,557 576 795 1,954 370 820 1,738 11,104

Haulage Drift m - 302 155 - 704 173 153 109 218 1,813

Main Ramp m 1,106 640 365 365 365 366 730 444 - 4,381

Ventilation Drift m 8 53 106 - 96 24 8 43 46 383

Ventilation Raise m - 171 200 - 147 - 63 57 177 815

Attack Ramp m - 938 239 2,696 1,634 1,098 721 610 1,184 9,120

Total m 1,627 3,914 3,622 3,637 3,742 3,614 2,044 2,084 3,362 27,646

Sublevel Ramp Tonnes 24,494 91,643 129,465 29,143 40,247 98,918 18,753 41,489 87,999 562,152

Haulage Drift Tonnes - 15,290 7,823 - 35,633 8,739 7,746 5,538 11,030 91,798

Main Ramp Tonnes 57,497 32,400 18,478 18,478 18,478 18,529 36,956 22,486 - 223,303

Ventilation Drift Tonnes 392 2,677 5,373 - 4,878 1,209 392 2,178 2,308 19,406

Ventilation Raise Tonnes - 3,018 3,529 - 2,604 - 1,106 1,014 3,130 14,400

Attack Ramp Tonnes - 14,658 3,740 42,125 25,536 17,150 11,260 9,534 18,494 142,497

Total Tonnes 82,384 159,686 168,408 89,746 127,376 144,545 76,213 82,239 122,960 1,053,556

Page 16.37




AMENDED NI 43-101 TECHNICAL REPORT Table 16.6 Yearly Mine Production

Year -1 production is retained in a stockpile until the start of production (Year 1), during which the plant is

available. Yearly process production is shown at the bottom of Table 16.6. Recovery is varied based on

gold grade as defined by recovery methods.

16.7 Mine Equipment

Mining is expected to be contracted for the life of mine. The contractor will provide the equipment required

to achieve the development and production schedule shown in Section 16.6. Table 16.7 shows the

equipment and underground facility requirements that are anticipated to complete the schedule. This list

will be subject to change based on negotiations with the selected contractor.

Mine Production Schedule Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Oxide K Tonnes 0 15 - 0 1 21 22 8 8 75

g Au/t 2.60 5.12 - 1.75 3.47 4.07 4.20 2.68 2.29 3.96

K ozs Au 0 2 - 0 0 3 3 1 1 10

Mixed K Tonnes - 6 0 1 1 1 1 9 0 20

g Au/t - 6.55 2.40 3.63 3.59 3.58 3.40 2.86 2.13 4.16

K ozs Au - 1 0 0 0 0 0 1 0 3

Sulfide K Tonnes 9 298 320 321 318 295 293 304 193 2,351

g Au/t 3.65 6.96 7.58 5.03 4.63 5.41 3.64 3.12 5.11 5.20

K ozs Au 1.0 66.8 78.0 51.9 47.3 51.2 34.3 30.5 31.8 393

Total K Tonnes 9 319 320 322 320 317 317 321 201 2,446

g Au/t 3.63 6.87 7.58 5.02 4.62 5.31 3.68 3.10 5.00 5.15

K ozs Au 1.0 70.5 78.0 52.0 47.5 54.1 37.5 32.1 32.3 405

Internal Waste K Tonnes 1 40 40 38 40 44 43 39 27 312

Total Diluted Material K Tonnes 10 359 360 360 360 361 360 360 228 2,759

Mined g Au/t 3.13 6.11 6.74 4.49 4.11 4.67 3.24 2.77 4.42 4.57

K ozs Au 1.0 70.5 78.0 52.0 47.5 54.1 37.5 32.1 32.3 405

Total K Tonnes - 352 360 360 360 361 360 360 245 2,759

g Au/t - 6.21 6.74 4.49 4.11 4.67 3.24 2.77 4.25 4.57

K ozs Au - 70 78 52 48 54 37 32 34 405

K Ozs Rec - 68 75 50 45 52 36 30 32 388

Net Rec 0% 96% 97% 95% 95% 96% 95% 95% 96% 96%

Page 16.38




AMENDED NI 43-101 TECHNICAL REPORT Table 16.7 Equipment Requirements and Underground Facilities

Mine productivity is based on the mine operating schedule, equipment availability, and the operating

efficiency. The mine is intended to operate 24 hours per day using 2-12 hour shifts. Standby time

includes one hour for shift startup and shut down and an additional 0.25 hours for lineout and safety

meetings per shift. The safety shifts will not be daily, thus the 15 minutes per day is accumulated to allow

for safety meetings once to twice a week as needed. In addition, 0.5 hours for lunch, and 0.5 hours for

breaks has been accounted for. This gives an overall operator availability of 81.3%.

Equipment availability was started at 85% and allowed to decline by 1% each year down to 80%, where it

is retained through the life of the mine.

Overall productivity for equipment is based on first principle calculations based on the type of equipment

and its operating specifications. For instance, the size of a truck, loading time, cycle time, and spotting

times determine the theoretical productivity for a truck. However, operating equipment will experience

operational delays such as queuing at a loading unit, slowing for other traffic on the ramp, waiting for

blasting and clearing of smoke after a blast, etc. This is equivalent to using a commonly known 50 minute

hour. This is accounted for by providing an efficiency factor, which has been incorporated, based on the

type of equipment and operating characteristics.

Peak # of

Units Item Comments

4 20t Articulated Haul Truck Primary haulage for ore and waste out of the underground mine

2 Double Boom Jumbo Used for primary development

2 Single Boom Jumbo Attack ramp and ore development and ore production

2 3m3 LHD Truck load out of production ore and development mining

3 3.6t LHD Stope mucking and backfill

1 Bolter Development bolting for support

1 Service Truck Equipment maintenance

1 Lube Truck Equipment fuel and lube

1 Grader Ramp and road maintenance

1 Scissor Truck Utility and shot loading

1 Flat Bed Transporting supplies

7 4x4 Utility Vehicle Personnel transportation, surveying, sample transportation

6 25 Liter/sec Water Pumps Dewatering (3 to pump to the surface and 3 auxiliary pumps)

1 Leaky Feeder System Communications

2 High Voltage Transformers Electrical distribution

2 Refuge Chambers Emergency refuge

2 Primary Ventilation Fans Exhaust fans at primary exhaust shaft (placed in parallel)

4 Auxiliary Fans Fans along primary development ventilating active mining areas

Page 16.39




AMENDED NI 43-101 TECHNICAL REPORT Theoretical productivity for loading trucks is based on a 3 m3 LHD loading 20-tonne articulated haul trucks.

A loading rate of 250 tonnes per hour was determined. It is assumed that the truck drivers will load

themselves, which means that there will be operational delays as the driver gets on board the LHD. Thus,

an efficiency of 65% is used to reduce the theoretical productivity and calculate the operating hours

required for LHD’s.

Haul truck productivity is based on the full cycle time. Cycle times were estimated from each stope to the

surface and back. The speed of trucks was determined from performance curves based on the haulage

gradient and a 2% rolling resistance. A speed limit of 24 kph was used for haul and return. A loading time

of 5.5 minutes was used, which includes an additional 2.0 minutes per load for the operator to load

himself, and the spot and dump time of 1.00 minute was added to the overall cycle time. The productivity

was reduced by a 70% efficiency to model operational interruptions.

Stope mucking operations will be carried out by 1.5 m3 LHD’s with a total capacity of 3.0 tonnes. A fill

factor of 95% was used to reduce the haulage capacity to 2.85 tonnes per load. The LHD travel speed

was assumed to be 8 kph for both haul and return. LHD haul cycles were estimated for each stope based

on the stope length and an average distance of 85m for attack ramps. An efficiency of 80% was used to

estimate the effective productivity.

Production drilling productivity was calculated based on development in ore for the first lift in a panel of

stopes and production drilling for the remaining lifts. The remaining lifts use less drilling and explosives in

comparison to development in ore. Development in ore assumed 2.5 m rounds for an average of

approximately 42 tonnes per round. A penetration rate of 29 m per hour was used and a setup time of

30 minutes per round was used and assumed to include moving time from hole to hole. An efficiency of

83% was used to calculate operating hours from the theoretical hours. It was assumed that 17% of ore

would be generated by development in ore.

Production drilling estimates also used a 29 m per hour drilling rate and an efficiency of 83%. A 10%

addition to operating time was estimated to reflect non-drill time. A total of 83% of material sent to the

plant is assumed to be produced by production drilling.

16.8 Manpower

A preliminary list of personnel required to achieve the production schedule has been compiled. This list is

shown in Table 16.8 and specifies the positions as reporting to either the contractor or mine operations.

This list is subject to negotiations with the selected contractor for mining.

Page 16.40




AMENDED NI 43-101 TECHNICAL REPORT Table 16.8 Mine Personnel Requirements

Mining General Personnel Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8

Mine Manager Mine 1 1 1 1 1 1 1 1 1

Mine Clerk Mine 1 1 1 1 1 1 1 1 1

Chief Mine Engineer Mine - 1 1 1 1 1 1 1 1

Mine Engineer Mine - 3 3 3 3 3 3 3 3

Mine Surveyor Mine 1 1 1 1 1 1 1 1 1

Surveyor Helper Mine - 2 2 2 2 2 2 2 2

Chief Geologist Mine 1 1 1 1 1 1 1 1 1

Ore Control Geologist Mine - 1 2 2 2 2 2 2 2

Core Drillers Mine 2 2 2 2 2 2 2 2 2

Samplers Mine - 2 3 3 3 3 3 3 3

Subtotal - Mine General Personnel Mine 6 15 17 17 17 17 17 17 17

Contractor Supervision

Contractor Mine Superintendent Contract 1 1 1 1 1 1 1 1 1

Mine Clerk Contract 1 1 1 1 1 1 1 1 1

Mine Foremen Contract 4 4 4 4 4 4 4 4 4

Mine Maintenance

Mechanics Contract 2 8 16 16 16 16 16 16 8

Welders Contract 4 8 8 8 8 8 8 4

Mine Electricians Contract 1 2 2 2 2 2 2 2 2

Servicement Contract 1 2 4 4 4 4 4 4 2

Mine Operating Personnel

Lead Miner - Development Contract 2 4 4 4 4 4 4 4 4

Miner's Helper - Development Contract 4 4 4 4 4 4 4 4 4

Lead Miner - Production Contract 4 4 4 4 4 4 4 4

Miner's Helper - Production Contract 4 4 4 4 4 4 4 4

Truck Drivers Contract 2 4 4 4 4 4 4 4 4

LHD Operators Contract 4 20 20 20 20 20 20 20 20

Total Personnel

Red Eagle Mine 6 15 17 17 17 17 17 17 17

Contractor Contract 22 62 76 76 76 76 76 76 62

Total 28 77 93 93 93 93 93 93 79

Page 17.1




AMENDED NI 43-101 TECHNICAL REPORT 17.0 RECOVERY METHODS

17.1 Process Design

17.1.1 Summary




process prior to the tailings being filtered. Part of the filtered tailings will be dry stacked in a tailings and

waste rock storage facility, the balance will be used as backfill in the mine. Filtrate will be recycled back to

the process plant to minimise the raw water requirement.

17.1.2 Design Philosophy



well Proven in industry.

The key project and ore specific criteria that the plant design must meet are:

The plant is designed for an initial throughput of 360,000 tpa with provision for future

expansion to at least 720,000 tpa;

Testwork shows that the ore is of medium hardness with average head grades over the

life of the project of 4.57 g/t gold and 8.5 g/t silver;

Mechanical availability for the process plant of 91.3%;

A level of automation to reduce the technical complexity of the plant with manual

operation where practical;

Equipment selection for reliability and ease of maintenance; and

Layout for ease of access to all equipment for operating and maintenance requirements

whilst maintaining a compact footprint that will minimise construction costs.

A detailed process design has been prepared incorporating engineering design criteria and key

metallurgical design criteria derived from the results of the metallurgical testwork.

17.1.3 Key Process Design Criteria

Table 17.1 presents the detailed process design criteria and mechanical equipment list.

Page 17.2




AMENDED NI 43-101 TECHNICAL REPORT Table 17.1 Summary of Key Process Design Criteria

Units Primary Source*

Plant Capacity tpa 360,000 MDA

Head Grade g Au/t 4.57 MDA

Head Grade g Ag/t 8.5 MDA

Design Gold Recovery % 96 Testwork

Design Silver Recovery % 69 Testwork

Crushing Plant Availability % 85 Lycopodium

Plant Availability % 91.3 Lycopodium

Bond Abrasion Index (Ai) 0.139 Testwork

Drop Weight (SMC, A*b) 41 Testwork

Bond Rod Mill Work Index (RWi) kWh/t 13.0 Testwork

Bond Ball Mill Work Index (BWi) kWh/t 15.9 Testwork

SG 2.73 Testwork

Grind Size (P80) µm 125 Lycopodium

SAG Mill Grinding Media

Consumption

kg/t 0.603 OMC

Tower Mill Grinding Media

Consumption

kg/t 0.158 OMC

Flotation Gold Recovery % 95.0 Testwork

Concentrate Regrind Size (P80) µm 15-20 Testwork

Pre-leach Thickener Solids Loading t/m2.h 0.65 Assumption

Leach Circuit Residence Time hours 48 Lycopodium

Leach Slurry Density % w/w 50 Lycopodium

Number of Leach Tanks 1 Lycopodium

Number of CIL Tanks 6 Lycopodium

Cyanide Consumption kg/t 0.39 Testwork

Hydrated lime Consumption kg/t 1.93 Testwork

Elution Circuit Type AARL Lycopodium

Elution Circuit Size t 2 Lycopodium

Frequency of Elution strips / week 6 - 12 Lycopodium

Cyanide Destruction Process SO2 /Air

Process

Lycopodium

Page 17.3




AMENDED NI 43-101 TECHNICAL REPORT 17.1.4 Projected Plant Recovery

Tests conducted on the six composite samples using the selected flotation/concentrate regrind/CIL

process showed a strong linear relationship between head grade and gold extraction. This relationship

was used to predict plant performance. Gold recovery modelling is based on the plant regrind achieving

the testwork regrind product size plus it includes a CIL solution loss of 0.01 g/t with a CIL pulp density of

50% w/w solids.

Results are summarized in Table 17.2 and Figure 17.1 with the anticipated gold recoveries in the plant.

Note that one outlier (Composite #4) was removed from the dataset because of unusually high recovery.

The exclusion of Composite #4 in recovery modelling provided a more conservative estimate of plant

performance. The overall relationship between mill head grade and projected gold recovery is:

[Gold Recovery Incl. Solution Loss (%)] = 0.45 x [Gold Head Grade (g Au/t)] + 93.394

This relationship was developed for head grades of 2 to 10 g/t. For head grades above 10.5 g/t the

recovery is capped at 98% in the projected gold recovery model.

Silver assays in the composite samples ranged from 4.0 to 12.9 g/t and yielded laboratory extractions

ranging from 65 to 85%. Silver was not included in the modelling due to lack of silver reserve estimates.

An average silver recovery of 68.6% was applied in the design of the elution circuit and downstream unit

operations.

Table 17.2 Summary of Metallurgical and Anticipated Gold Recoveries

Sample ID Test No. Process

Roll Bottle Test Results Anticipated

Overall Plant

Recovery

Au %

Head Calc Tails Metallurgical

Assay Grade Assay Recovery

Au g/t Au g/t Au g/t Au %

Composite 1 CY-49 Flot/CN 2.35 2.48 0.13 94.8 94.5






Page 17.4




AMENDED NI 43-101 TECHNICAL REPORT Figure 17.1 Projected Gold Recovery vs Mill Head Grade

17.1.5 Selected Process Flowsheet

The process plant design incorporates the following unit process operations:

Primary crushing with a single toggle jaw crusher (1,100 x 700 mm) to produce a

crushed product size of 80% passing (P80) 100 - 120 mm;

A crushed ore surge bin (30 m3) with a nominal capacity of 1 hour process plant feed of

45 t;

Single stage SAG mill (5.0 diameter x 3.5 m EGL – 1,200 kW) in closed circuit with

cyclones to produce a P80 grind size of 125 µm;

Rougher scavenger flotation (6 x 8 m3 conventional cells) to produce a sulphides/gold

concentrate;

Tower mill (150 kW) for regrind of the concentrate to a P80 grind size of 15 - 20 µm;

Pre-leach thickener (16 m dia) to minimise carbon in leach (CIL) tankage and reduce

overall reagent consumption;

A hybrid CIL circuit incorporating one leach tank and six adsorption tanks (430 m3 each)

with 48 h total residence time;

90.0

91.0

92.0

93.0

94.0

95.0

96.0

97.0

98.0

99.0

100.0

0.0 1.0 2.0 3.0 4.0 5.0 6.0 7.0 8.0 9.0 10.0 11.0 12.0

Go

ld R

eco

very

(%

)

Head Grade (Au g/t)

Lab Leach Extraction Projected Gold Recovery

Page 17.5




AMENDED NI 43-101 TECHNICAL REPORT A 2 tonne AARL elution circuit with electrowinning and smelting to produce doré bars;

Cyanide destruction using the SO2 / Air process (120 min retention); and

Tailings pressure filtration (445 m2 filter area) to 16% moisture content.

A process flowsheet depicting the unit process operations is shown in Figure 17.2.

Page 17.6





Figure 17.2 Santa Rosa Gold Project Simplified Process Flowsheet

Page 17.7




AMENDED NI 43-101 TECHNICAL REPORT 17.2 Process Description

17.2.1 Run-of-Mine (ROM) Pad

Ore is delivered by truck directly from the adjacent decline to a ROM stockpile. Sufficient area has been

allocated to allow for ore blending and a total capacity of 3,000 tonnes (three days of plant feed).

A front end loader (FEL) reclaims ore from the stockpile on the ROM pad and delivers ore to the ROM bin.

17.2.2 Crushing Circuit

ROM ore enters the crushing circuit via a small ROM bin fed by the FEL. A grizzly (500 mm aperture) is

fitted to the ROM bin. The grizzly is inclined to be self cleaning as far as possible and to facilitate oversize

removal using the FEL feeding the crushing circuit. Secondary breakage, if needed, will be managed on

the ROM pad.

ROM ore is withdrawn from the ROM bin at a controlled rate by a variable speed apron feeder (1,200 mm)

and discharged into the jaw crusher.

The primary jaw crusher has been sized for the planned expansion capacity of 720,000 tpa. In practice,

maximum particle size, not throughput dictates the crusher selection and as such the unit will not be

capacity constrained.

The jaw crusher product discharges onto a 1,000 mm wide conveyor feeding directly to the crushed ore

surge bin (45 t). An electromagnet suspended above the conveyor removes magnetic tramp metal from

the primary crushed ore.

Under normal operating conditions the crushing rate into the surge bin exceeds the ore reclaim rate to the

milling circuit. Excess crushed ore overflows the surge bin and is directed on to a conveyor feeding the

overflow stockpile. When the crushing circuit is off line, for example during periods of crusher

maintenance, ore from the overflow stockpile is loaded by FEL into the surge bin to maintain mill feed.

Page 17.8




AMENDED NI 43-101 TECHNICAL REPORT Figure 17.3 Crushing Circuit Layout

17.2.3 Grinding and Classification Circuit

A single stage semi-autogenous grinding (SAG) mill in closed circuit with hydrocyclones was selected to

grind the ore and F80 of 100 mm to produce the target grind P80 size of 125 µm. The SAG mill is

equipped with a 1,200 kW variable speed drive. The mill rotation speed range is between 60 – 80% of

critical speed to handle varying ore competencies. In addition to standard charge weight and power

control algorithms, the mill is equipped with noise monitoring to alarm low mill charge levels. This circuit

provides stable and simple operation that is well understood and proven at many successful installations.

Crushed ore is withdrawn from the surge bin at a controlled rate by a variable speed apron feeder and fed

via the mill feed conveyor directly to the SAG mill. A weightometer indicates the instantaneous and

totalised mill feed tonnage.

Oversize material from the SAG mill trommel, consisting of pebbles and worn steel grinding media, reports

to the scats bunker and is returned to the mill by FEL via the surge bin. SAG mill grinding media is also

be added by FEL into the surge bin as required.

Page 17.9




AMENDED NI 43-101 TECHNICAL REPORT The SAG mill trommel undersize gravitates to the mill discharge hopper, where it is diluted with process

water and pumped to the hydrocyclone cluster for classification. Relatively large hydrocyclones (6 x

250 mm diameter) have been specified to minimise the number of cyclones in the cluster and to reduce

the potential for spigot blockages occurring from coarse SAG mill discharge material.

The combined cyclone overflow stream, with a nominal pulp density of 40% w/w solids, gravitates to the

flotation conditioning tank. The cyclone underflow is collected in the underflow launder and gravitates to

the feed chute of the SAG mill.

General maintenance lifts around the mill and cyclones are carried out by the site mobile crane. The

milling area layout accommodates crane access for all heavy lifts. A liner handler is provided for mill liner

change-outs.

The mill floor slab is sloped towards a collection sump. The mill hopper overflow and cyclone feed pump

dump lines are routed to discharge directly into the sump.

The mill shell structural design is specified to allow high ball charges. A conservative liner design will limit

grinding media trajectories such that impact will be on the toe of the charge at low mill charge levels.

The milling circuit design allows for a future pebble crusher installation in the unexpected case that dilution

of the feed with the more competent hanging wall material could reduce the SAG milling rate. In such a

scenario a pebble crusher will be required to maintain milling rates. Although not required when

considering the current mine plan, this design allowance has been made to ensure changes in ore

competency do not impact on the circuit throughput.

Page 17.10




AMENDED NI 43-101 TECHNICAL REPORT Figure 17.4 Grinding Circuit Layout

17.2.4 Flotation and Regrind

Sulphide flotation is expected to recover 95% of the gold into a concentrate representing 10-20% of the

ore mass. The concentrate is reground to 15-20 microns to expose the gold for improved leach recovery.

Cyclone overflow slurry gravitates to the flotation conditioning tank where the slurry is diluted to 35% w/w

solids and flotation collector (PAX) will be added. The conditioned slurry gravitates to six conventional

8 m3 flotation cells arranged in a rougher – scavenger configuration with total residence time of

25 minutes. Flotation concentrate reports to the regrind circuit and flotation tails flow to the pre-leach

thickener feed box.

Page 17.11




AMENDED NI 43-101 TECHNICAL REPORT The concentrate regrind circuit consists of a tower mill (150 kW) in closed circuit with hydrocyclones.

Regrind circuit cyclone overflow gravitates to the pre-leach thickener feed box where it is recombined with

the flotation tails stream ahead of the pre-leach thickener.

17.2.5 Pre-Leach Thickening

A pre-leach thickener has been included ahead of the CIL circuit to allow for adequate slurry dilution

ahead of flotation and provide high slurry density (50% solids) to the CIL circuit so both unit processes are

optimised. There are significant benefits with a higher and more constant leach feed density in terms of

reduced reagent costs and reduced overall circuit tankage volume required to achieve the target CIL

residence time.

The recombined flotation tails and re-ground concentrate slurries are de-aerated prior to entry into the pre-

leach thickener. Flocculant is added into the launder and feedwell of the thickener. Thickener underflow

is pumped to the CIL circuit. Thickener overflow reports to the grinding water tank to provide a separate

water circuit between grinding/flotation and cyanide leaching.

The pre-leach thickener has been sized to meet the throughput requirements of the expanded plant

capacity (720,000 tpa).

17.2.6 Leach and Carbon Adsorption

The leach characteristics identified during the metallurgical testwork program indicated that a 48 hour

leach is required to achieve optimal gold extraction. To obtain this retention time a circuit configuration of

one leach tank and six leach/adsorption tanks was selected. This is the minimum tankage required for

acceptable stage efficiency and for achieving target solution grades.

Thickener underflow is pumped to the leach tank via the vibrating trash screen. The trash screen removes

any misreporting coarse ore particles, wood fragments, organic material, plastics and lime slurry grits that

would otherwise accumulate in the carbon circuit and/or 'peg' the inter-tank screens.

The slurry pH is adjusted with hydrated lime added to the vibrating trash screen feed box. Sodium cyanide

solution is metered into any of the first four tanks to maintain the desired cyanide concentration in the

circuit.

Each tank is fitted with a dual impeller mechanical agitator to ensure uniform mixing. Compressed air is

sparged to the leach and CIL tanks to maintain a high dissolved oxygen profile.

The tanks are interconnected with launders and slurry flows by gravity through the tank train. Bypass

facilities allow any tank to be removed from service for agitator or screen maintenance.

Page 17.12




AMENDED NI 43-101 TECHNICAL REPORT Fresh and regenerated carbon is added to the circuit at the last adsorption tank (Tank 7) and is advanced

by air lift counter-current to the slurry flow. An air swept intertank screen in each adsorption tank retains

the carbon. Loaded carbon from the first adsorption stage (Tank 2) is air lifted to the loaded carbon

recovery screen mounted above the CIL tanks. The carbon is washed and dewatered on the recovery

screen and reports to the acid wash column. Slurry and wash water from the carbon recovery screen

underflow return to Tank 1.

Leach tails from Tank 7 gravitate to the vibrating carbon safety screen via the tails sampler. The safety

screen recovers any carbon leaking through worn inter-tank screens or overflowing the tanks. Screen

underflow gravitates to the cyanide destruction circuit.

The tanks are located in a bunded area with a sloping concrete floor. Any spillage from the circuit reports

to one of two sumps and can be returned to the circuit or to the carbon safety screen ahead of the cyanide

destruction circuit.

Figure 17.5 Leach and Carbon Adsorption Circuit Layout

17.2.7 Acid Wash

Loaded carbon is acid washed in 2 tonne batches in a rubber lined carbon steel column with 3% w/w

hydrochloric acid (HCl) solution. Washing takes place six days per week during day shift. The dilute

hydrochloric acid is pumped through the column in an up-flow direction to remove contaminants,

predominantly carbonates, from the loaded carbon. This process improves the elution efficiency and has

the beneficial effect of reducing the risk of calcium-magnesium 'slagging' within the carbon during the

regeneration process.

After a two hour acid soak period the carbon bed is rinsed with four bed volumes of raw water to displace

any residual acid from the carbon. Dilute acid and rinse water is disposed of in the tailings collection

hopper. Acid-washed carbon is transferred to the elution column for stripping using raw water.

Page 17.13




AMENDED NI 43-101 TECHNICAL REPORT 17.2.8 AARL Elution

The AARL system was selected for elution as it provides flexibility in the number of stripping cycles per

week to suit gold production.

Carbon transferred to the elution circuit has a design loading of 2,700 g Au/t and 4,000 g Ag/t. A single

stainless steel elution column is provided to hold the two tonne batch of acid washed carbon. Strip solution

consisting of 2% NaCN and 2% NaOH is pumped from the strip solution tank through inline heat

exchangers into the base of the elution column.

The loaded carbon is pre-soaked in the cyanide / caustic solution for 30 minutes to prepare the gold for

elution. The carbon is then be eluted by hot strip solution (120 °C) which exits the column to the pregnant

solution tank. Outgoing pregnant solution passes through the recovery heat exchanger to heat the

incoming strip solution.

17.2.9 Electrowinning and Doré Smelting

Gold is electrowon from the pregnant solution within the security area of the goldroom. The pregnant

solution circulates through two parallel cells of 12 cathodes each in multiple passes to ensure a minimum

gold tenor in the barren eluate. Direct current is passed through stainless steel anodes and stainless steel

wool mesh cathodes to deposit gold and silver sludge on the cathodes. A sludging cell design with in-tank

washing of the cathodes has been selected to simplify the cathode handling process. Electrowinning

continues until the solution exiting the electrowinning cells is depleted of gold. Solution discharging from

the electrowinning cells returns by gravity to the pregnant solution tank.

Rectifiers, one per cell, are located in a non-secure area below the cells allowing maintenance access

without breaching gold room security. Rectifier remote indication and controls are located adjacent to the

electrowinning cells.

In normal operations electrowinning is completed in less than a 12 hour shift.

The system is configured to allow multiple pass electrowinning for operation when two strips per day are

processed. The option also exists to operate the electrowinning circuit in open circuit with one of the

pregnant tanks operated as a lean eluate tank.

The electrowon silver and gold sludge is removed from the cathodes by washing with high pressure water.

The resulting sludge is filtered in laboratory style pressure filters and dried in an oven. The sludge is then

direct smelted with fluxes in a natural gas-fired smelting furnace to produce doré bars. Slag from smelting

operations is returned to the milling circuit.

Based on the design carbon loading (2755 g Au/t and 3995 g Ag/t) smelting will occur every second day

yielding production of 27 kg of doré. The cast doré bars will carry approximately 41% gold and 59% silver.

In the early years of production with higher mill head grades, additional elution and doré smelting cycles

may be required each week and these can be accommodated within this production schedule.

Page 17.14




AMENDED NI 43-101 TECHNICAL REPORT Fume extraction equipment is provided to remove gases from the cells, sludge drying oven and smelting

furnace.

17.2.10 Carbon Regeneration

After completion of the elution process, the barren carbon is transferred from the elution column to the

carbon dewatering screen prior to entering the feed hopper of the horizontal carbon regeneration kiln

(100 kg/h capacity). Any residual water in the feed hopper is also drained before the carbon enters the

kiln. In the kiln the carbon is heated to 650 - 750°C and held at this temperature for 20 minutes to allow

regeneration to occur. Regenerated carbon exiting the kiln is quenched by spray water on the carbon

sizing screen. The screen oversize (regenerated, sized carbon) is returned to the CIL circuit while the

quench water and fine carbon reports to the carbon safety screen. The fine carbon retained by the safety

screen is collected in a bin for eventual sale to realize any adsorbed gold values.

17.2.11 Cyanide Destruction Circuit

The SO2 / air process has been selected to reduce weak acid dissociable cyanide (CNWAD) in the plant

tailings stream to less than 1 ppm. After passing through the carbon safety screen the CIL tailings slurry

gravitates to the cyanide destruction circuit. This circuit consists of two agitated tanks which can operate in

either series (normal operation) or parallel configuration; each tank has 1 hour residence time.

The SO2 source for detoxification is sodium metabisulphite (SMBS). Oxygen is supplied by sparging air

into both tanks. The detoxification process requires the presence of a soluble copper catalyst which is

supplied by metering copper sulphate solution directly from a mixing tank. Hydrated lime slurry from the

CIL distribution ring main is added to maintain pH in the range of 8.0 to 9.0.

Probes measuring Eh and pH are used to control dosing of SMBS and hydrated lime, respectively.

17.2.12 Tailings Disposal

Tailings from the detoxification circuit are pumped to the tails surge tank at the mine portal for filtration.

The filtration plant consists of two recessed plate filter presses (one operating and one standby). Each

press has the capacity to handle 100% of the tailings.

The operating cycle consists of filter closing and clamping, filter feed, air blow, cake discharge, and finally

cloth washing. Filtrate is collected and pumped to a clarifier system. The cloth-wash discharge is

collected separately and pumped back to the clarifier. Clarified water is returned for reuse in the process

area. The small amount of sludge underflow from the clarifier is periodically pumped back to the filters via

the tails surge tank.

The filtered tailings contain nominally 16% moisture. A front end loader collects the filtered tailings which

are hauled via surface trucks to the dry waste management facility immediately east of the plant site or

backhauled via mine trucks underground as mine stope backfill.

Page 17.15




AMENDED NI 43-101 TECHNICAL REPORT 17.2.13 Future Expansion

The design provides for future expansion to at least 720,000 tpa. Critical equipment has been sized to

handle the increased throughput (jaw crusher, SAG mill, thickener, elution circuit and gold room). The

layout of the plant allows for:

Addition of a ball mill in order to maintain the target grind;

Installation of a pebble crusher;

Twinning of flotation and concentrate regrinding;

Installation of three new leach tanks. The leach tank in the initial circuit is designed to

convert to an adsorption tank when the circuit is expanded; and

Addition of a third press filter for tailings.

17.3 Reagents and Consumables

Table 17.3 summarises the reagents and consumables for the Santa Rosa Gold Project. Sufficient stocks

will be maintained on site (2-4 weeks) to ensure that supply interruptions due to transport or weather do

not impact production.

Page 17.16




AMENDED NI 43-101 TECHNICAL REPORT Table 17.3 Reagent Summary

Description Application Mode of Delivery Method of Handling

Hydrated Lime pH control in CIL and

Detox circuits

1 tonne bulk bags Mixed to 25% w/w, distributed via ring

main

Potassium Amyl

Xanthate (PAX)

Sulphide flotation

collector.

Mixed to 10% solution, meter pumped to

flotation

MIBC Flotation frother 200 L drums or isotainers Meter pumped to flotation

Copper Sulphate Catalyst in Detox 25 kg bags Mixed to 20% w/w, metered to flotation

and detox.

Sodium Cyanide leach reagent for CIL and

elution

Dustless pellets in 1 tonne

boxed bulk bags

Mixed to 20%w/v, metered to CIL and

elution

Caustic Soda pH modifier in elution 25 kg bags of pearl pellets Mixed to 20%w/w, metered to elution

circuit

Hydrochloric Acid Carbon acid wash reagent 32% concentrated acid delivered

in 1,000 L isotainers

Metered to acid wash feed line to obtain a

3%w/w solution

Activated Carbon Gold adsorption in CIL 500 kg bulk bags To carbon sizing screen (vendor package)

Sodium

Metabisulphite

Reagent for Detox 1 tonne bulk bags Mixed to 20% w/w, metered to detox

circuit

Flocculant Flocculant 25 kg bags Mixed and diluted in vendor package,

dosed to preleach thickener

Grinding Media SAG mill balls 2 tonne bags Charged via FEL to ore surge bin.

Compressed

Natural Gas

heat source Trucked and pumped into site

storage tank

Reticulated to elution heater, carbon kiln

and smelting furnace

Antiscalant Grinding water circuit

17.4 Services

17.4.1 Raw Water

Raw water for the project is diverted from the spring source and catchment area of the La Veta creek to a

three day capacity storage pond located east of the plant site. Additionally, seasonal precipitation plus

surplus water from the La Veta creek system is collected in the seepage collection pond, which will be

pumped back to process water storage tank at the plant. The seepage collection pond will have a constant

overflow to sustain the minimum flow rate requirement for the La Veta creek (7.2 m3/h).

For a year with an average rainfall the runoff entering the project site will be on an average between 22 to

44 m3/h.

Raw water is used to feed the potable and stripping water treatment plants as well as reagent mixing, and

gland water.

The estimated plant raw water make-up will be approximately 14 m3/h.

Page 17.17




AMENDED NI 43-101 TECHNICAL REPORT 17.4.2 Fire Water

Fire water for the process plant is drawn from the bottom third of the raw water tank (270 m3 reserve).

The fire water pumping system consists of:

an electric delivery pump to supply fire water at the required pressure and flowrate;

a diesel driven pump that will automatically start in the event that power is not available

for the electric pump; and

an electric jockey pump to maintain fire ring main pressure.

Fire hydrants and hose reels are placed throughout the process plant, fuel storage and plant offices at

intervals that ensure complete coverage in areas where flammable materials are present.

17.4.3 Potable Water

Water from the raw water tank supplies the potable water treatment plant. Water treatment includes

filtration, carbon contacting, and chlorination. Additional ultra-violet sterilisation units will be installed on

outgoing potable water distribution headers if required. Potable water is reticulated from the potable water

storage tank to the site ablutions, safety showers and other potable water outlets.

Potable water is also pumped to the mine services facilities.

17.4.4 Process Water

The process water is separated into two circuits: the grinding water circuit, and the process water circuit

(used in the CIL section). The two circuits are designed to operate separately to avoid cyanide and high

pH in the flotation circuit. Separate storage tanks of 80 m3 and 125 m3 are provided.

Grinding water consists primarily of recycle from the pre-leach thickener overflow with make up from raw

water and any surplus mine discharge water. The grinding water is fed to the grinding circuit, flotation

circuit, regrind circuit, and pre-leach thickener feed dilution.

The process water consists primarily of the tailings filtrate. Grinding water tank overflow also reports to the

process water tank. Seepage pond, monitoring and event pond water can also be returned to the process

water tank.

Any overflow from the process water tank discharges into the monitoring pond. If the cyanide levels are

too high in the monitoring pond the water is pumped back to the process water tank and onto the cyanide

detox circuit to adjust levels in the plant effluent.

Duty and standby pumps are provided for raw water, grinding water and process water. Antiscalant is

added to condition the grinding water and reduce fouling of pipelines, spray nozzles and screen decks.

Page 17.18




AMENDED NI 43-101 TECHNICAL REPORT 17.4.5 Plant, Instrument, Flotation, and Air Supply

Plant and instrument air is supplied from two high pressure screw compressors. The air is dried before

distribution with one air receiver supplying both plant and instrument air.

A low pressure blower is dedicated to the flotation circuit. Three positive displacement low pressure

blowers provide air to the CIL tanks (two operating one standby) and can service the flotation circuit when

that blower is offline.

17.5 Instrumentation and Control

The level of instrumentation and control has been selected to provide a basic regulatory control to

maintain steady operation with minimal process excursions. Following industry practice for similar size

plants, a supervisory control and data acquisition (SCADA) and programmable logic controller (PLC)

architecture was selected for the plant wide process control system. It is a reliable and low cost approach.

The SCADA/PLC integrates original equipment manufacturer (OEM) controllers in the field into a uniform

operator interface located in the main control room. Where vendor packaged process control systems are

not available, logic will be developed at the SCADA level for process control and monitoring.

The control room houses two PC based operator terminals which both act as the control system (SCADA)

servers as well as engineering / operator stations. The system includes a historian capability for data

analysis and reporting.

In general the status of process plant drives is reported to the SCADA and is displayed in the control

room. Local control stations are located in the field in proximity to the relevant equipment. These will, as

a minimum, contain Start and Lock-Off-Stop (LOS) pushbuttons which are hard-wired to the drive starter.

Local selection allows each drive to be started and stopped by the operator in the field via pushbutton.

Plant drives are generally started remotely from the control room. Remote selection requires the

equipment to be started from the control room. Status indication of process interlocks as well as the

selected mode of operation is displayed in the control room.

Safety interlocks such as emergency stops and thermal protection are hardwired and will apply in all

modes of operation. All software process interlocks also apply in both Local and Remote modes.

17.6 Metallurgical Accounting

A weightometer on primary crusher discharge conveyor measures the primary crushed ore tonnage. A

second weightometer on the SAG mill feed conveyor determines mill feed tonnes.

The pre-leach thickener underflow stream is pumped to a two stage cross cut feed sampler then gravitates

into the vibrating trash screen feed box. The sampler collects representative samples of the leach feed.

Leach tails from Tank 7 gravitates to a two stage cross cut feed sampler for representative samples of the

tails.

Page 17.19




AMENDED NI 43-101 TECHNICAL REPORT Density and flow meters on the leach feed allows the dry tonnage of solids to be determined as a cross

check on the mill feed tonnage determined from the mill feed weightometer. In conjunction with the leach

feed and tails samplers, the mass flow measurements allow the gold recovered in the CIL to be

calculated.

Regular ‘in circuit' surveys will allow reconciliation of precious metals in feed compared to doré production.

Weights and assays of the doré bars will be reconciled against the calculated gold recovery.

Reconciliation of the amount of reagents used over relatively long periods will be achieved by delivery

receipts and stock takes. On an instantaneous basis, reagent usage rates of cyanide, elution and

detoxification reagents to unit operations are measured (l/min) and accumulated (m³ ) using flow meters.

Page 18.1




AMENDED NI 43-101 TECHNICAL REPORT 18.0 PROJECT INFRASTRUCTURE

18.1 Project Infrastructure

18.1.1 Infrastructure Scope

The Santa Rosa Gold Project consists of an underground decline accessed mine, processing plant, and

plant infrastructure. As part of this Feasibility Study, a site plan was developed for the project site. Major

plant infrastructure consists of the following:

Main Access Road from the concession gate to portal pad;

44 kV Power transmission line from EPM substation (8.9 km long);

44 kV Switch Yard;

Main Switch Room;

Reagents Switch Room;

Mine Services and Filter Plant Switch Room;

Dry Waste Management Facility (DWMF);

Sedimentation Pond;

Seepage Pond;

Monitoring Pond;

Event Pond;

Reagent Storage Building;

Plant Administration Building;

Assay and Metallurgical Laboratory;

Plant Workshop and Main Warehouse Building;

Mine Truck shop/ Warehouse Buildings; and

Diesel fuel storage and filling station.

Page 18.2




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.1 Overall Site 3D View

18.1.2 Site Access

The Santa Rosa Gold Project is located near the town of Santa Rosa de Osos, 73 km north-west of the

city of Medellin, in the province of Antioquia. Primary access to the site is via paved Highway No. 25

north-northeast through Copacabana and Don Matías for approximately 65 km to a turn-off located 12 km

south of Santa Rosa de Osos. From the turn-off to the east, it is approximately 8 km to the proposed mine

portal via an unpaved road.

The existing Site Access road is in good condition and does not require major upgrading but some

sections will have to be improved in order to facilitate shipping larger equipment to site. The last 400 m of

the road from the existing concession gate to the Santa Rosa Site will require complete realignment in

order to minimize cut and fill quantities and connect directly to the south end of the proposed mine portal

area.

18.1.3 Site Plan

The proposed process plant site is bounded by the underground mining area to the south, the La Veta

creek valley to the east, and the ridge line separating the San Francisco valley from the La Veta creek

valley to the north and west. The topographic relief in the project area is moderate with gentle to relatively

steep-sided valleys and hills. Elevations range between 2,300 m and 2,500 m above sea level.

Page 18.3




AMENDED NI 43-101 TECHNICAL REPORT The general stratigraphic column for the site consists of a surface residual soil composed of silt and clay

deposits to a depth of 2-3 m underlain by a coarser saprolite unit of 30-50 m thickness, and fresh

granodiorite bedrock. The proposed plant site will be located on a mixed cut and fill section, with

maximum cut depth of 35 m and maximum fill depth of 15 m. The heavy equipment foundation structures

will be located on the cut platform and the ponds and lighter buildings will be located on compacted fill

material. The majority of the foundations will be reinforced concrete spread and strip footings with rafts for

the heavy equipment and other settlement sensitive structures.

The following factors were considered when developing the site plan layout:

Minimizing the environmental impact and visibility;

Reduced haul truck travel distance from the mine portal area to the ROM pad / primary

crusher;

Usage of existing access roads;

Equalizing cut and fill material to avoid fill import or excessive cut to waste;

Ensure all heavy equipment foundations are supported on native undisturbed soil with

sufficient load bearing capacity;

Optimized ore flow through the process plant;

Compact plant layout to reduce overall footprint, which minimizes electrical cabling,

piping and service roads thus reducing capital costs;

Locating the filter presses as close as possible to the mine portal and the DWMF.

The mine portal is situated in the southern extremity of the La Veta creek valley. All the mine service

buildings (truck shop, warehouse) have been located immediately north of the portal and adjacent to the

main haul road to the primary crusher. The filter presses, the mine area substation, and the diesel fuel

storage area are located along the eastern edge of the open space in front of the portal. In order to keep

the project site compact all the major process plant infrastructure is located immediately north of the mine

service buildings on a platform cut into the ridge separating the San Francisco valley from the La Veta

creek valley.

Page 18.4




AMENDED NI 43-101 TECHNICAL REPORT The crusher area has been located as close as possible to the mine portal with an elevation 9 meters

above the adjacent mine service and reclaim areas. The rest of the processing plant will be constructed

north-northwest of the crusher and reclaim areas 7 metres below the portal and reclaim areas. The heavy

equipment has been located as much as possible towards the west edge of the site on a cut platform in

order to provide sufficient bearing capacity and minimize settlement under the critical equipment and

structures. Lighter structures (gold room, reagents storage, ponds) have been sited on the east edge of

the plant site and are supported primarily on a compacted fill material. The plant administration building,

assay and metallurgical laboratory, and the plant workshop and main warehouse building are located at

the north edge of the plant site. On the east the plant site area is bounded by the la Veta creek valley,

which is the proposed location for the DWMF.

18.1.4 Site Roads

The site roads will be constructed of compacted rockfill overlying competent native soil. The design cut

slopes are 75 degrees up to 10 m slope height. For cuts more than 10 meters 1.5 m wide bench will be

required every 10 meters. The fill slopes will be 2.5H : 1.0V. The maximum vertical slope will be 10%.

Dust suppression shall be applied to the gravel surface to minimize dust generation.

Page 18.5





Figure 18.2 Overall Site Layout

Page 18.6




AMENDED NI 43-101 TECHNICAL REPORT 18.2 Plant Area Buildings

18.2.1 Buildings General

The major building structures will be made of structural steel with uninsulated roof and wall cladding. The

building foundations will consist of cast in-situ conventional reinforced concrete footings. Secondary

buildings will be of pre-engineered, pre-fabricated, or portable module type, where applicable. The

ancillary buildings will require varying degrees of air conditioning and ventilation. The process plant facility

will be entirely outdoors, and only the main control room, the electrical switch rooms, and the laboratory

building will be air conditioned. The gold room will be ventilated only. Fire protection, lightning protection

and smoke detection have been considered for various buildings.

18.2.2 Plant Workshop and Main Warehouse

The Plant Workshop and Main Warehouse building is located at the north end of the plant site area.

This building will house the main work area for equipment maintenance and repairs for the crushing

plant, mill, regrind, and recovery plant together with a warehouse containing the critical spares for the

processing plant. It will also provide an office area, tool room, and toilet facilities. The building will be a

single storey steel frame structure (32 x 12 m) with uninsulated cladding. The steel structure will be

supported on reinforced concrete spread footings.

18.2.3 Reagents Permanent Storage

The Reagents Storage building is located east of the plant workshop and warehouse building and will

house the storage areas for sodium cyanide, sodium hydroxide, hydrochloric acid, hydrated lime, carbon,

sodium metabisulphite, flocculants, and antiscalant. Most reagents are delivered on a monthly basis with

the exception of hydrated lime, which will be delivered bi-weekly. Hence, the building is designed for a

month’s worth of reagent storage. It will be a steel frame open structure (20 x 8 m) with uninsulated metal

roofing and cladding on the east side only. The steel structure will be supported on conventional

reinforced concrete spread footings.

18.2.4 Gold Room

The gold room is located east of the CIL tanks and west of the monitoring pond. This will be a steel frame

building (14 x 8 m) with metal roofing, totally enclosed by cladding and security mesh. The steel structure

will be supported on reinforced concrete spread footings. The interior and exterior of the building will be

under surveillance via cameras and closed circuit television.

18.2.5 Tails Filtration Building

The tails filtration building will be located opposite the mine portal and south of the DWMF haul road. It will

house the two filter presses producing the filter cake to be deposited in the DWMF or mixed with cement

to be used as a backfill in the mine. The building will be a steel frame (42 x 10 m) with uninsulated metal

roofing and cladding, supported on concrete walls with reinforced concrete strip footings.

Page 18.7




AMENDED NI 43-101 TECHNICAL REPORT 18.2.6 Plant Air Compressors/ Blower Shed

The plant air compressors/ blower shed is located west of the detoxification area in the south west corner

of the processing plant. The building will house two compressors and two blowers servicing the

processing plant. The shed will be constructed as a steel frame with metal roofing and partial side

cladding and will be supported on concrete spread footings.

18.2.7 Filter Area Compressors Shed

The filter area air compressors/ blower shed is located south of the filter press building in the portal pad

area. The building will house two compressors servicing the filter presses. It will be a steel frame building

with metal roofing and partial side cladding and will be supported on concrete spread footings.

18.2.8 Guard House

The guard house will be located at the main access road entrance and adjacent to the mine portal. It will

be a modular type building (4 x 4 m) with sandwich panel walls and roof. The guard house will be

supported on a concrete slab on grade.

18.2.9 Plant Administration Building

The plant administration building is located in the north west corner of the plant site, adjacent to the assay

and metallurgical laboratory and the plant workshop and main warehouse building. This building will

consist of two modified 40 foot shipping containers with a steel support structure covering the space

between the containers. It will contain the plant offices, meeting room, toilet facilities, break room, first aid

area, and a reception area.

18.2.10 Assay and Metallurgical Laboratory

The assay and metallurgical lab will be located between the plant administration building and the plant

workshop and main warehouse building. This will be a full service laboratory facility and will house the

following areas: fire assay lab area; wet lab area; sample preparation area; data input and weighing area;

metallurgical lab area; sanitary facilities; and an office area. The laboratory will be sized to process

50 samples per day, solids, solutions and carbon samples. These will be samples from the underground

and surface drilling; samples from the processing operations; and various environmental samples. The

building will consist of two modified 40 foot containers with a steel support structure spanning between the

containers.

Page 18.8




AMENDED NI 43-101 TECHNICAL REPORT 18.2.11 Plant Control Room

The control room serves as a central monitoring station of the process plant. It will be a 4.5 x 6 m

prefabricated building with sandwich panel walls and roof located on top of the CIL tanks for optimum

visibility and control. The control room will consist of an operators’ room and a server room. The

operators’ room will house two PC based operator interface terminals for operating and monitoring all

process plant and one printer for recording all key process and maintenance parameters. The server

room will house the UPS, patch panel for the server, Ethernet switches, and fibre optic breakout tray.

18.2.12 Crusher Control Room

The crusher control room will house the crusher control station and will be located in the primary crusher

area. It will be a prefabricated building with sandwich panel walls and roof.

18.2.13 Main Switch Room

The main switch room will house the LV switchgear and the motor control centres (MCCs) for the feed

preparation and milling areas, flotation and water services, and screening and air services. It will also

contain the lighting and small power distribution panels and the area PLC and marshalling panels. The

main switch room will be located in the midst of the process plant between the SAG mill and the gold room

and minimizes cable lengths and improves MV efficiency. It will be a pre-fabricated steel constructed

building with double cladding and insulation. It will also be fitted with a double door and a single door. The

building will be supported on a structural steel frame on reinforced concrete strip footings.

18.2.14 Reagents Switch Room

The reagents switch room will house the motor control centre for the leaching, desorption, goldroom, and

reagents areas. It will also contain the lighting and small power distribution panels and the dedicated area

PLC and marshalling panels. The reagents switch room will be located directly north of the reagent mixing

and storage area. The building will be a modified 40 foot shipping container supported on a structural steel

frame with reinforced concrete strip footings.

18.2.15 Filter Plant Switch Room

The filter plant switch room will house the motor control centre for the filter plant. It will also contain the

lighting and small power distribution boards, the PLC remote rack and marshalling panels. The filter plant

switch room will be located in the portal pad area north of the tails filtration building. It will be a pre-

fabricated steel constructed building with double cladding and insulation. It will also be fitted with a double

door and a single door. The filter plant switch room will be supported on a structural steel frame on

conventional reinforced concrete spread footings.

Page 18.9




AMENDED NI 43-101 TECHNICAL REPORT 18.3 Mine Area Buildings

Mining operations buildings will be located at the mine-portal area. These will include a truck shop,

warehouse, and offices, all of which are provided and maintained by the mining contractor. In addition,

the owner will provide and maintain offices for mine management, engineering, and geology personnel, in

conjunction with other Red Eagle Mining facilities which are currently on site on site.

18.3.1 Truck Shop

The truck shop will be provided by the mining contractor, and will include the necessary service bays and

an office for maintenance, supervision, and planning. The building will be a steel structure with metal

cladding and a concrete slab on grade.

18.3.2 Warehouse

A mobile equipment and spare parts warehouse will be attached to, or adjacent to, the truck shop. The

building will be a steel structure with metal cladding and a concrete slab on grade. The facility includes

inside storage for parts and supplies, an office, and a tool crib/small parts area. A fenced storage yard

located adjacent to the warehouse will be used to store large items or bulk materials which can withstand

exposure to the elements. The outside storage area will have a compacted, gravelled base and security

fencing.

18.3.3 Explosives Storage

Transportation, permits, management, and storage of explosives will be the responsibility of the mining

contractor. Only explosives sufficient for a few days of use will be stored on site. This will include

detonators, boosters, and bulk explosives or ANFO. Each of these types of explosives will be stored in

separate magazines on the surface. An approved small magazine and an adjacent large building already

exist adjacent to the old Hilo Azul shaft (a legacy of former artisanal miners on the property). These

facilities will be converted to approved magazines suitable to the contractor’s requirements and will be

surrounded with security fencing. The contractor will be required to provide security at all times while

explosive materials are stored on site.

18.4 Building Fire Protection Systems

Systems to be provided for personnel and property protection include: smoke/heat detectors and manual

pull stations, fire extinguishers, fire hydrant coverage of all process plant areas; process plant buildings,

and internal fire hose coverage for all enclosed building areas.

Fire hose cabinets and external fire hydrants will be located so that all interior areas of the buildings are

within reach of a fire hose stream.

Page 18.10




AMENDED NI 43-101 TECHNICAL REPORT A firewater header system will be provided at the site and will cover the process plant and ancillary

buildings, along with fire hose coverage throughout the facility, supplemented by hand held fire

extinguishers. A dedicated fire hydrant will be installed to provide fire hose coverage throughout the

reagent area, with hand held fire extinguishers. Fire protection coverage for the crusher area will be

provided by site fire hydrants supplemented by hand held fire extinguisher.

For electrical rooms ionization type smoke detectors will be provided, integrated with an automatic dry fire

protection system with additional hand held fire extinguishers.

Hand held fire extinguishers will be provided for the control rooms.

18.5 Sewage Treatment

A septic system will be utilized for sewage disposal. Septic tanks will be located at the process plant, and

near the mine portal for mining operations. The septic tank sludge will be removed by vacuum truck at

regular intervals.

18.6 Security System and CCTV Monitoring

Process plant and goldroom area accesses will be controlled and monitored 24 hours per day. The

goldroom will also be monitored by remote motion, vibration and temperature sensors to detect

unauthorized intrusion. High security cameras will be located in the goldroom, and at the plant guard

house.

In addition to the high security system, an independent CCTV system will monitor the crusher and ore

feeders, with the monitors located in the main control room. A video recorder will capture all relevant entry

/ exit details in high security areas and log all security alarms in chronological order. Security signals will

be transmitted via secure dedicated cables with the system backed up by dedicated UPS.

18.7 Site Services

18.7.1 First Aid

In addition to the first aid post at the existing camp / administration site, a first aid post will be located in

the plant administration building and will be staffed by a qualified registered nurse 24 hours per day. An

ambulance will be stationed on the project site 24 hours per day. The nearest full-service medical facility

is in Santa Rosa de Osos, approximately 20 km (30 minutes) from the project site (Hospital San Juan de

Dios).

18.7.2 Communication Systems

An existing communications tower installed at the Santa Rosa exploration camp will provide internet and

telephone services. This system is synchronized to the Red Eagle Medellín and Vancouver office servers.

Connections from the existing communications system to the process plant and mine-area offices will be

made by installing repeater dishes as necessary.

Page 18.11




AMENDED NI 43-101 TECHNICAL REPORT Hand-held and base-station radios will be provided for operators for on-site communications. Surface

mobile equipment will be equipped with radios.

18.7.3 Underground Communications

A leaky feeder system will be used for underground communications. The system includes a coaxial or

“radiating” cable that is run through the drifts to act as an antenna for radio communication. Line amplifiers

will be placed every 350 m to 500 m to boost signals as required. This will allow for two-way radio

communication throughout the mine. The system will be installed and maintained by the mining

contractor. Capital costs have been included with utilities for development costs.

18.7.4 Transportation

Project personnel will be transported to the site from the town of Santa Rosa de Osos and other nearby

communities. Transportation will be provided for the workers via locally hired buses and/or large vans.

There is no plan to provide camp facilities at the project site.

18.7.5 Solid Waste Disposal

Solid wastes will be disposed of in a manner complying with local regulations. Allowable products will be

disposed of in a solid-waste landfill constructed on site. Products not allowed to be disposed of in the

landfill will be transported to appropriate facilities off site.

18.8 Fuel Delivery and Storage

Diesel fuel will be delivered to the mine site via contractor-owned tanker trucks and stored in tanks on-site

for use by mine and surface mobile equipment and vehicles. Each storage tank is contained in a lined

basin to assure no fuel is leaked into the environment. The diesel fuel storage and filling station is located

opposite the mine portal in the area adjacent to the main access road.

This facility will be provided and managed by the mining contractor.

18.9 Site Fencing

The mine portal and process plant areas will be enclosed by four-strand barbed wire fencing to prevent

livestock from wandering onto the property. Security fencing (chain-link fence with barbed-wire or razor-

wire crowns) will be utilized around the 44 kV Switchyard and the electrical switch rooms.

Page 18.12




AMENDED NI 43-101 TECHNICAL REPORT 18.10 Power Supply and Distribution

18.10.1 Power Supply – Incoming

The Santa Rosa plant will be supplied by a 44 kV overhead power line coming from an existing EPM

substation located approximately 9 km south of the plant site at Rio Grande. Figure 18.3 shows the EPM

power substation and plant site with the proposed route for the new 44 kV power line. The overhead

power line and associated infrastructure will have sufficient capacity for the planned expansion (10 MW).

Figure 18.3 44 kV Overhead Power Line from EPM Substation

The 44kV line will be terminated at the main substation dead-end structure. The line is connected to the

main breaker through a motorized disconnect switch. A set of surge arresters, current transformers and

voltage transformers will be provided at the line side of the main disconnect switch.

18.10.2 Main Transformer and Medium Voltage Switchgear

Power will be stepped down from 44 kV to 4.16 kV by means of a 44/4.16 kV, 7.5/10 MVA oil filled

ONAN/ONAF transformer with delta configured primary and wye configured secondary. The main

transformer is grounded via a resistor transformer. The 4.16 kV facilities include switchgear and two

300 kVAR shunt capacitor banks, station services, protection and control. The switchgear will be a single

bus arrangement.

Page 18.13




AMENDED NI 43-101 TECHNICAL REPORT 18.10.3 Power Distribution

The stepped down 4.16 kV power will be distributed to the plant switchrooms (load centres) using 5 kV

cables installed as a combination of cable trays and direct buried. The main switch room will supply power

for equipment located in the milling, feed preparation, and flotation areas. The reagents switch room will

supply power to leaching, desorption, goldroom, and reagents areas. The plant administration building and

the plant workshop and main warehouse will be fed from separate 4.16 kV lines and their dedicated 4,160

/ 480 V transformers.

A 4.16 kV direct buried cable will supply power to the 4.16 kV mine services switchgear, which will feed

the mine area operation, ventilation fans, and the filter plant switchroom. The mine area operation and

ventilation fans will be fed via underground 4.16 kV lines. Each line will be protected by a 4.16 kV feeder

breaker. The step down transformers and their protection at the destination will be supplied by the mining

contractor and ventilation supplier. A dedicated feeder breaker will supply power via an oil filled outdoor

4,160 / 480 V, 500 kVA transformer to the filter plant switch room.

18.10.4 Power Demand

18.10.5 Process Plant Power Demand

The estimated load for the Santa Rosa Process Plant is as follows:

Connected load: 5.16MW

Name plate load: 4.93MW

Maximum demand: 3.94MW

Average(operating)load: 2.58MW

Largest size motors: 1.2MW SAG MILL

Overall Power Factor 0.89

4.16kV system neutral grounding resistance grounded

480V neutral grounding solidly grounded

A large proportion of the electrical load will be due to the process plant. The process plant is expected to

run continuously for 24 hours per day. The load list summary is shown in Table 18.1.

Page 18.14




AMENDED NI 43-101 TECHNICAL REPORT Table 18.1 Process Plant Power Requirements

In general, equipment will not be loaded above 80% of its rated capacity. The peak demand load was

calculated from the load list taking load factors into account.

Demand current will not exceed 80% of the MCC bus ratings

Demand current will not exceed 80% of the feeder capacity

18.10.6 Mine Area Power Requirements

It is estimated that a peak of 1.25 MW will be required for the underground mine and infrastructure. Of

this, 400 kW is required at the ventilation shaft collar (400 kW demand at the shaft collar would be very

rare as these ventilation fans are intended to be run one at a time), leaving a peak requirement of

approximately 900 kW underground. Table 18.2 shows the demand by unit for each type of equipment

and Table 18.3 shows the quantity of equipment that will be operating each year.

Process Area

Nameplate

Load

kW

Connected

Load

kW

Max Load

Factor

LF

Peak

Demand Load

kW

Power

Factor

Cosϕ

Peak

Demand

kVA

Utilisation

Factor

UF

Operating

Load

kW

SAG Mill 1,200 1,288 0.9 1,159 1 1,165 0.91 873

Feed

Preparation &

Milling

661 693 0.75 523 0.88 607 0.52 322

Screening / Air

Services 671 701 0.65 459 0.86 543 0.69 320

Floatation &

Water System 816 858 0.72 614 0.82 750 0.60 403

Leaching,

Desorption,

Goldroom &

Reagents

743 769 0.79 605 0.81 742 0.53 319

Tailing Filters 731 745 0.66 491 0.86 573 0.54 268

Facilities and

Buildings 108 108 0.85 92 0.80 115 0.91 84

Plant Total 4,930 5,162 0.76 3,943 0.89 4,394 0.66 2,584

Page 18.15




AMENDED NI 43-101 TECHNICAL REPORT Table 18.2 Maximum Mine Area Power Demand

Note that dewatering Pumps 1 through 4 are staged in overtime based on the location and pumping

needs. See Section 21.2.1.9.

Table 18.3 Mine Area Electrical Equipment in Operation by Year

Of note, the primary and secondary ventilation fans would be placed in parallel within the same ventilation

shaft, and the secondary fan would be acting only as a backup to the primary fan. Thus, the electrical

demand from the two fans shouldn’t exceed the 200 kW rating for a single fan. Also, the single and

double boom jumbos and the bolters would be powered by diesel for transportation, and only running off

of electricity during drilling and bolting operations.

Mobile Equipment Demand Units Comments

Single Boom Jumbo 45 kW Note that Mobile Equipment travels on Diesel and then operates on electrical

Double Boom Jumbos 135 kW

Development Bolters 55 kW

Core Drills 45 kW

Ventilation Equipment

Primary Ventilation Fan 200 kW Note that the Secondary fan is backup for the Primary, thus there should be

Secondary Ventilation Fan 200 kW 24/7 utilization of the 2 together, but only one at a time would operate.

Auxiliary Fans 75 kW

Dewatering

Dewatering Pump #1 47 kW Pumps differ in requirements based on head to be pumped.

Dewatering Pump #2 63 kW



Auxiliary Pumps 15 kW

Other

Compressors 93 kW Compressed air for underground services and drilling

Mobil Equipment Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8

Single Boom Jumbo # 1 2 2 2 2 2 2 2 2

Double Boom Jumbos # 2 2 2 2 2 2 2 2 2

Development Bolters # 1 1 1 1 1 1 1 1 1

Core Drills # 1 1 1 1 1 1 1 1 1

Ventilation Equipment

Primary Ventilation Fan # 1 1 1 1 1 1 1 1 1

Secondary Ventilation Fan # 1 1 1 1 1 1 1 1 1

Auxilary Fans # 2 4 4 4 4 4 4 4 4

Dewatering

Dewatering Pump #1 # 1 1

Dewatering Pump #2 # 1 1 1 1 1 1 1 1

Dewatering Pump #3 # 1 1 1 1 1

Dewatering Pump #4 # 1 1 1

Auxilary Pumps # 1 3 3 3 3 3 3 3 3

Other

Compressors # 1 2 2 2 2 2 2 2 2

Page 18.16




AMENDED NI 43-101 TECHNICAL REPORT Lycopodium has estimated costs for electrical distribution on site, which includes two 5 kV drops and

transformers to convert power to 480 V for underground equipment. One of the lines and transformers

would be placed on the surface near the furthest western ventilation shaft. That transformer will be used

for the two surface exhaust ventilation fans. An allowance for cables and switching gear from the

transformer to the surface ventilation equipment has been made in the capital cost estimate.

The other transformer will be placed near the portal during the start of operations. This transformer will be

used to operate development equipment during the construction period. Once development has been

completed to the bottom of the first ventilation shaft, the underground transformer will be moved to a

location nearby. The 5 kV power input to the transformer will be placed in the ventilation shaft to reduce

the amount of cable required. Power will be distributed to production and development areas from the

transformer via a 480 V cable hung from the back. Switch gear will be maintained near the production

areas.

During the life of the mine, the transformers will be moved as required to prevent efficiency losses. Two

additional locations will be used, one at the bottom of the second ventilation shaft to service deeper mining

on the eastern side of the deposit, and another at the main decline turn off to reach the western portion of

the deposit.

Capital requirements for the underground electrical distribution are discussed in the capital cost section.

Operating costs will consist primarily of salaries for electricians to maintain power to the production and

development areas. Maintenance of the underground electrical distribution is included in the contractor’s

responsibility; thus, the operating cost related to the electrical distribution system is included in the

contractor costs.

18.10.7 Voltage Selection

Table 18.4 describes the designed voltage levels for the project. These may vary slightly during the course

of the project based on specific requirements of equipment manufacturers.

Table 18.4 Voltage Levels

Category Condition Voltage (V) Phase Frequency

Primary Supply (High Voltage) Nominal 44,000 3 60

Primary Distribution (Medium Voltage) Nominal 4,160 3 60

Low Voltage Utilization Nominal 480 3 60

Low Voltage Lighting Nominal 380/220 3 60

Switchgear Control Voltage Nominal 125 DC -

Page 18.17




AMENDED NI 43-101 TECHNICAL REPORT 18.10.8 Voltage Drops Limit

Voltage drop limit will not exceed the following:

Steady-States:

Distribution feeders (from 4.16 kV switchgear to transformer primary): 1%

Motor feeder cables: 3%

Lighting fixtures (from distribution boards): 3%

Motor Starting

Motor starting (of nominal bus voltage at point of connection): 15%

18.10.9 Supply Voltage

The supply voltage to different equipment in the process plant will be as per table 18.3 below:

Table 18.5 Supply Voltage

Equipment Type Rating Voltage (V) Phase

Motors below 0.5 kW 220 1

From 0.5 kW to 200 kW 460 3

above 200kW if not soft start 4,160 3

SAG mill 1200 4,160 3

Heaters Less than or equal to 1 kW 220 1

Greater than 1 kW 480 3

Lighting and receptacles 220 1

Outdoor floodlights 380 3

Instrumentation and control panels 120 1

18.10.10 Emergency Power Supply - Process Plant

A 600 kVA and two 200 kVA diesel generating units are included to supply emergency power for

process plant and administration buildings. The emergency power is not meant to be used for

sustaining the operations of the plant. The purpose of the emergency power supply is to provide

power during plant power outage to the following consumers:

Plant administration building;

Guard house;

Page 18.18




AMENDED NI 43-101 TECHNICAL REPORT 30% of area lighting;

Control room power;

Thickener rake system

Thickeners underflow pumps (50%);

Fire-detection system and dry-pipe fire-fighting system (main fire loop has diesel pump).

18.10.11 Emergency Power Supply – Mine Area

A 200, kVA emergency generator will supply power for essential mine area loads during power outage,

primarily to maintain ventilation and pumping in critical areas.

18.10.12 Construction Power

The construction power demand is estimated to be about 1,000 kVA. One (1) 600 kVA and 2X 200 kVA

diesel (natural gas) generator sets will supply power to the different areas of the process plant and mine

site during construction.

These diesel generator sets will be used as emergency power supplies for the process plant and mine site

upon the completion of the construction. All mining and construction contractors have based their quotes

on construction site power supplied by generator sets.

18.10.13 Power Quality

Power factor correction calculations will be used to determine the requirements for leading reactive power.

Multi-stage capacitor banks will be provided to raise the overall site power factor. The capacitor banks will

be connected to the main 4.16 kV switchgear bus.

18.10.14 Lightning Protection System (LPS)

The designed lightning protection system shall be able to identify the most suitable location(s) for the air

terminal(s) to provide the zones of protection to the structure/building against direct lightning strikes. The

method to determine the zones of protection will be based on the ‘Rolling Sphere’ technique or equivalent

proven method.

Air terminal(s) shall be installed at any salient point (e.g. corners and highest point) on the structure /

building / roof to achieve the high probability interception point.

In addition all building columns and lighting poles will be grounded. This provides extra risk mitigation on

top of the air terminals installed for the LPS

Page 18.19




AMENDED NI 43-101 TECHNICAL REPORT 18.10.15 Electrical Switch Rooms

Electrical equipment such as switchgear, secondary substations, MCCs, panel boards, UPS, process

control system, I/O cabinets, etc, will be housed in designated electrical switch rooms within and around

the process plant in a climate controlled environment.

Air-conditioning units will be provided to control the humidity and room temperature. Electrical switch

rooms will be raised 1.8 m above standard floor levels to prevent flooding and provide room for cable pull.

Man-doors and double doors will be equipped with panic bars and will open outwards. The electrical

switch rooms will be sealed to provide the required fire rating.

18.11 Transportation and Logistics

18.11.1 General

Colombia is serviced directly, or through trans-shipment, by ample carriers of all modes, and it has the

supporting infrastructure to receive major project cargo from offshore. Santa Rosa de Osos is located

70 km northeast of Medellin, on Highway No. 25, which is a major transport route to Caribbean ports.

Figure 18.4 Santa Rosa Mine Site Location

Santa Rosa Gold Project

Page 18.20




AMENDED NI 43-101 TECHNICAL REPORT The Santa Rosa project involves the delivery of major equipment from various parts of the world, including

Asia, North America and Europe. For this project, marine and truck transportation services will be utilized,

individually or in combination. Off-site laydown areas, marshalling areas and project warehousing have

been identified in the Medellin and Santa Rosa de Osos areas.

Ground transportation will determine the maximum size loads to be delivered to site. There is adequate

road access to the Project site from the ports of Barranquilla and Buenaventura.

18.11.2 Port of Barranquilla

The preferred port for the Santa Rosa project is the port of Barranquilla situated 630 km north of the

project site on the Caribbean Sea. The port is connected to the site via two lane national Highway No. 25,

which is part of the Pan-American Highway. The 630 km represents an average of 13 hours driving time

between the port and site.

Figure 18.5 Port of Barranquilla

Barranquilla is the largest port in Colombia and includes storage for over 6,000 TEUs of containers and

warehouses with a capacity of over 140,000 metric tons of bulks and 100,000 metric tons of general

cargo.

The port of Barranquilla is a natural river port considered to be a medium sized with excellent shelter

characteristics accommodating vessel sizes of over 150 metres in length. The port has a total 2,100 linear

metres of river frontage, including one dock of 1,050 metres with six berthing positions for ships. The

water depth varies between 9 to 10.5 metres with a cargo pier of 9 metres. The port provides anchorage

of 8 to 9 metres.

Barranquilla port has lifts and cranes ranging up to 100 tonnes including mobile and floating cranes.

Page 18.21




AMENDED NI 43-101 TECHNICAL REPORT 18.11.3 Port of Buenaventura

Buenaventura Port is located on the Pacific Ocean close to the Panama Canal and is one of the closest

ports on the American Continent to the Far East. The port is situated approximately 560 km southwest of

the project site linked by the two-lane national Highway Nos. 25 and 29. The 560 km represents an

average of 11 hours driving time between the port and site.

Figure 18.6 Port of Buenaventura

The port of Buenaventura is a natural river port considered to be a small sized with good shelter

characteristics accommodating vessel sizes of over 150 metres in length. The Port access canal is

31.5 kilometers, or 17 nautical miles long. At low tide, it is 10.3 metres deep in the outside bay and in the

interior bay the depth is 12.5 metres at low tide. The access canal width is 200 metres.

Buenaventura port has lifts and cranes ranging up to 80 tonnes including mobile and floating cranes.

18.11.4 Containerized Cargo and Legal Load Trucking

Standard trucking equipment is available within Colombia to meet the general trucking and container

transportation requirements of the Project. Larger trucking companies could potentially mobilize the

necessary equipment at selected ports, according to arrival schedules.

Page 18.22




AMENDED NI 43-101 TECHNICAL REPORT 18.12 Waste Rock and Tailings Facility

18.12.1 Design Concept

The proposed facility is comprised of mine development rock (granodiorite) and filtered tailings blended

together to form the Dry Waste Management Facility (DWMF), located to the north of the project site

(Figure 18.7). The concept of filtered tailings has been adopted to minimize the footprint of the deposit,

reduce the amount of water treatment demand and have a stable mine waste deposit during and after

operations. Water will not be stored within the DWMF, eliminating the risk of a tailings dam breach and

downstream flood wave.

Figure 18.7 General Configuration of the DWMF

The low moisture content tailings (16%) will be transported via truck and deposited along with mine

development rock. The material will be spread and compacted to form an unsaturated mass which will not

require a downstream containment structure. This design was adopted because it offers the following

advantages:

Allows maximum water recovery and minimum fresh water use for the Santa Rosa Gold

Project;

Eliminates the risk of dam breach and tailings runout as associated with conventional

storage facilities;

Suited to areas where seismic activity may occur because the tailings are compacted to

produce a stable material;

Minimizes the construction material required for the facility;

Allows progressive rehabilitation and revegetation;

Page 18.23




AMENDED NI 43-101 TECHNICAL REPORT Incorporates a drainage system beneath the DMWF that collects seepage from the

waste materials and significantly reduces the potential for groundwater contamination;

and

Allows increase in capacity by raising the elevation of the DWMF.

18.12.2 Geotechnical Investigation

To provide the geotechnical information required for the DWMF design, a series of geotechnical drillholes

were carried out between April and May 2014 (Figure 18.8 and Table 18.6).

Table 18.6 Summary of geotechnical investigation

ID Total depth (m) Area Comments

BH-GA-02 30.0 Portal pad Samples for laboratory analysis

recovered

BH-GA-03 45.7 Portal area Samples for laboratory analysis

recovered

BH-GA-04 28.0 Filter press Samples for laboratory analysis

recovered.

BH-GA-05 80.2 Monitoring well near

La Veta creek

installation of open pipe (Casagrande)

piezometer

BH-GA-06 15.1 Process plant Samples for laboratory analysis

recovered.

BH-GA-07 71 Process plant

Standard Penetration Tests and

samples for laboratory analysis

recovered, installation of open pipe

(Casagrande) piezometer

BH-GA-08 16.8 Process plant

Standard Penetration Tests and

samples for laboratory analysis

recovered

BH-GA-10 15.0 Process plant Samples for laboratory analysis

recovered

BH-GA-11 30.0 DWMF Along right abutment of proposed

DMWF deposit

BH-GA-15 172.8 Pump well

Pump tests performed for assessment of

hydrogeologic conditions within shear

zone

Page 18.24




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.8 Geotechnical Investigation Carried Out in 2014 (Golder, 2014a)

The drill holes completed in the project area had the following objectives:

At the portal and process plant area (GA-03, -04, -06, -07 and -08), estimate soil profiles

for subsequent bearing capacity and slope stability analyses;

At the DWMF area (GA-11), estimate an approximate depth to bedrock and soil

conditions for DWMF foundation preparation;

At the decline area (GA-02, -03, -04), assess depth to bedrock and quality of saprolite

and bedrock to be crossed by the initial decline segment; and

Along the shear zone (GA-05, -15), assess the hydrogeological behaviour of the shear

zone via a deep pump test (170 m deep).

N

Shear zone

Plant and

portal areas

DWMF

area

Waste soil

area

Page 18.25




AMENDED NI 43-101 TECHNICAL REPORT Results of the drilling program, detailed field and laboratory tasks are summarized in the field report by

Golder (2014a). The drilling program was carried out by L.T. Geoperforaciones Ltd. and supervised by

Golder and Red Eagle Mining personnel.

Standard Penetration Tests (SPT) were performed along saprolite and ‘undisturbed’ Shelby samples

collected at different depths for estimation of index properties (moisture content, grain size distribution,

consistency limits) and other engineering properties (shear strength, compaction and consolidation). The

results of the laboratory tests are described by Golder (2014b).

Geotechnical tests on disturbed and undisturbed samples were conducted at the soils laboratory of the

Universidad Nacional de Colombia in Medellín. The testing programme included the following:

38 natural moisture content determination (ASTM D2216-98);

38 grain size analyses (ASTM D422-07);

38 Atterberg limits (ASTM D4318-05);

3 one-dimensional soil consolidation tests (ASTM D2435-04);

2 Triaxial shear strength tests (ASTM D4767-04);

2 Standard compaction (Proctor) tests (ASTM D 698-07); and

2 Direct shear tests (ASTM D3080-04).

The testing programme focused on estimating the behaviour of the surface soils (saprolite and residual

soil) for the design of foundation systems and earthworks associated with the process plant, portal area

and DWMF. A geotechnical characterization was based on the results of these tests and results of the

field investigation campaigns. Results of the laboratory tests are summarized in the following tables and

values used in design summarized in Table 18.6. Details of the geotechnical characterization are

presented by Golder (2014b).

The general soil profile in the project area is characterized by a thin topsoil layer (0.3 to 0.5 m thick), a

clayey residual soil layer (generally 1-2 m thick), overlying a 5 to 40 m thick saprolite layer (Figure 18.9).

The saprolite grades from fine-grained (clay and silt) near the surface to coarse-grained (sand and small

rock fragments) with depth. A transition zone comprising fractured rock (sap-rock) separates the fresh

bedrock from saprolite.

Natural moisture contents in the saprolite range between 20 and 40%, liquid limits between 35 and 60%

and plasticity indexes between 15 and 20%. Fines content in the saprolite range between 30 and 80%

with some lower values likely due to washing of the fines during sampling.

Page 18.26




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.9 Saprolite Thickness in Project Area (CRA, 2013)

The DWMF, portal access and process plant infrastructure will be founded on saprolite; therefore the

geotechnical characterization is focused solely on the behaviour of this layer. The underlying fractured and

fresh rock are considered stable and are sufficiently deep to not have an influence on the stability of the

DWMF, portal and process plant areas.

Saprolite is a transition geological unit between fresh rock and completely weathered soil. It is created

from the slow weathering processes degrading the underlying igneous and metamorphic bedrock, and it is

an in-situ soil (i.e. no mass transport has occurred). Saprolites often exhibit continuation of discontinuities

or fractures remaining from the underlying rock and can be preferential seepage paths or potential failure

planes.

Generally, saprolite in the project area has good stability conditions, evidenced by nearly vertical cuts for

roads (Figure 18.10). Except for the bottom of the valleys, these materials are constantly above water

table (unsaturated) thus their shear resistance is high and allows steep cuts (between 55 and 75 degrees)

with little or no support requirements. The main issue these materials exhibit is erosion susceptibility.

When exposed to runoff or direct precipitation erosion and rilling is rapidly triggered, thus runoff control is

designed for all cuts and project roads.

Page 18.27




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.10 Near-Vertical Cuts in Saprolite

Table 18.7 summarizes the results of shear strengths correlated from SPT tests as a function of depth.

The average values of friction angle obtained from this correlation are 33 degrees for the upper 3 meters

and 38 degrees for depths greater than 3 m.

Page 18.28




AMENDED NI 43-101 TECHNICAL REPORT Table 18.7 Shear strength estimates for saprolite soil

Boring Depth (m)

NFIELD (N1)60 ' (°) From To

BH-GA-07

1.5 1.95 6 10 33

3.5 3.95 20 29 41

5.5 5.95 20 26 40

7.55 8 13 16 36

9.5 9.95 14 16 36

12.5 12.95 28 29 41

15.5 15.95 27 26 40

18.5 18.95 45 41 45

21.5 21.95 48 41 45

24.5 24.95 48 39 44

28.55 29 33 25 40

33.5 33.6 - - -

34 34.45 90 63 51

BH-GA-08

0.5 0.95 2 4 28

2 2.45 3 5 29

4 4.45 7 10 32

5.95 6.4 6 8 31

7.9 8.35 11 13 34

9.85 10.3 31 35 43

11.8 12.25 4 4 28

13.4 13.85 18 18 37

15.35 15.8 22 21 38

15.8 16.25 30 29 41

Additional shear strength values for in-situ drained and undrained conditions were estimated using triaxial

tests (Table 18.8) for bearing capacity and slope stability analyses. Drained friction angles (’) range from

31 to 38 degrees and undrained shear strengths (Su) between 100 and 520 kPa. To estimate the stability

behaviour of saprolite to be used as fill material at the plant site and portal pad areas, direct shear tests on

reconstituted samples after standard Proctor tests were done to estimate maximum dry density and

optimum moisture content (Table 18.9).

Page 18.29




AMENDED NI 43-101 TECHNICAL REPORT Table 18.8 Summary of triaxial strength tests on 'undisturbed' samples

Drillhole Depth

(m) USCS ' (°)

Undrained Shear

Strength (kPa)

BH-GA-07 0.70 MH

33.6 118

36.1 231

31.1 401

BH-GA-07 4.70 SM

38.0 199

35.2 340

32.7 522

Table 18.9 Summary of direct shear strength tests on remoulded samples (to be used as

compacted fill)

Sample Code Dry Density

(KN/m3)

Optimum Moisture Content

(%)

(°)

Apparent Cohesion Intercept

(kPa)

PC-GA-01 14.7 26 30 51

PC-GA-02 13.9 31 31 29

Table 18.10 summarizes the results of geotechnical characterization of in-situ and remoulded saprolite

soils for use in slope stability analyses. These were estimated from direct shear tests on reconstituted

samples to near optimum water content for the compacted saprolite and ‘undisturbed’ samples for in-situ

saprolite.

Table 18.10 Summary of strength values for in-situ and compacted saprolite

Soil Unit Weight

(kN/m3)

Apparent

cohesion

intercept

(kPa)

Drained

Friction

Angle, ' (deg)

In-Situ saprolite 17 25 30

Compacted saprolite 18.5 30 30

Page 18.30




AMENDED NI 43-101 TECHNICAL REPORT 18.12.3 Dry Waste Management Facility Design

The proposed San Ramón deposit DWMF stores a combination of filtered tailings and mine development

rock. The design concept involves the separation of contact and non-contact waters in the DWMF footprint

to minimize the amount of water infiltrating into the deposit and thus flowing into the seepage collection

pond. The tailings filtration process will produce a filter cake with a moisture content of 16%. This percent

of water in the tailings stream will allow for proper compaction (95+% of standard compaction test) of the

tailings mass within the deposit so a stable, self-supporting mass is constructed.

The proposed DWMF area is a 0.11 km2 sub basin located north of the process plant. It is approximately

450 m long and 300 m wide, with relief of about 60 m.

Geo-Logic Associates performed geotechnical tests on the project tailings for the design of the DWMF.

The objective of the testing program was to characterize the strength and hydraulic conductivity of the

compacted tailings mass for use in infiltration and slope stability analyses. The following tests were

performed:

Standard compaction (ASTM D-698) tests to estimate the maximum dry density and

optimum moisture content achievable by compaction;

Grain size analysis (ASTM D-422-07) tests to assess grain size distribution;

Direct Shear (ASTM D-3080-04) tests to estimate shear resistance of the compacted

tailings. These tests were performed on tailings samples at optimum moisture content

and to a moisture content 3% above optimum moisture content;

Unconfined Compression Tests (ASTM D-6236) to estimate cured strength of tailings

mixed with cement. These results will be employed in underground mine backfill designs

part of future activities; and

Permeability tests (ASTM D-2434) to estimate saturated hydraulic conductivity of

compacted tailings.

Figure 18.11 shows the results of the standard compaction test. The maximum dry density achieved

was 19.3 kN/m3 at a moisture content of 14%. The particle size distribution for the tailings is shown on

Figure 18.12. The tailings have a P80 of 106 microns, P50 of 20 microns and P10 of 2 microns.

Page 18.31




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.11 Standard Proctor Test Results (GeoLogic Associates, 2014)

Figure 18.12 Tailings Particle Size Distribution. Data from GeoLogics Associates, 2014

Hydraulic conductivity tests performed on reconstituted samples to dry densities above 95% of the

maximum dry density showed an average saturated conductivity of 6x10-7 cm/s Direct shear tests

performed on samples at optimum moisture content gave strength parameters of and c =

100 kPa, and samples tested at 3% moisture content above optimum gave and c = 20 kPa. For

design, a drained friction angle of 30 and no cohesion were employed to account for the long-term loss of

apparent cohesion due to saturation of the deposit.

0

10

20

30

40

50

60

70

80

90

100

1 10 100 1000

PER

CEN

T FI

NER

PARTICLE DIAMETER (mm)

Page 18.32





18.12.4 DWMF Slope Stability

The strength values for tailings and saprolite soils were employed in static and pseudo-static slope

stability analyses using the theory of limit equilibrium. An analysis section through the maximum deposit

height was used, consisting of about 50 m of compacted tailings and about 35 m of saprolite overlying

bedrock. Phreatic surface was simulated along the tailings/foundation contact as a drain will be

constructed along the bottom of the DWMF. Figure 18.13 shows the critical slip surface for a static failure,

a non/linear, relatively shallow surface extending from the crown of the DWMF to about 30 downstream of

the toe of the deposit. Table 18.11 summarizes the results of the DWMF slope stability analyses. A

pseudo-static horizontal coefficient of 0.1 g was used in the analyses.

Figure 18.13 Critical slip surface for static conditions (Golder, 2014b)

Table 18.11 Results of slope stability analyses (Golder, 2014b)

Case Factor of Safety

Static 1.58

Pseudo-static 1.14

The calculated slope stability factors of safety are above local regulations and typical values employed in

mining projects (1.5 minimum for long term conditions and 1.0 for pseudo-static conditions).

Page 18.33




AMENDED NI 43-101 TECHNICAL REPORT 18.12.5 Deposition Plan

Based on the Santa Rosa mine plan, there will be 1.05 Mt of mine development rock produced during the

mine life. Approximately 40% of this quantity is expected to be used for underground stope fill, and the

remainder will be placed in the DWMF. A portion of the mine development rock will be used for

construction, foundations and road maintenance.

The tailings stream from the process plant will be 1,000 tonnes per day (tpd). A total of 2,759,000 tonnes

of tailings will be produced during the mine life. Table 18.12 below summarizes the expected production of

mine rock and tailings:

Table 18.12 San Ramón project waste production plan

Units Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Total

Mine Rock kt 4 139 135 135 136 139 141 138 88 1,055

Mine Rock to DWMF kt 4 55 41 41 41 44 41 42 11 320

Mine Rock Volume at

1.9 t/m3

m3x

1000 2 29 22 22 22 22 22 22 6 169

Tailings kt 0 360 360 360 360 360 360 360 239 2,759

Tailings to DWMF kt 0 235 219 219 219 219 211 216 125 1,663

Tailings to DWMF

Volume at 1.9 t/m3

m3x

1000 0 124 115 115 115 115 111 114 66 875

Material to DWMF kt 4 290 260 260 260 260 253 259 137 1,983

Material to DWMF

Volume

m3x

1000 2 153 137 137 137 137 133 136 72 1,044

A total of 1.04 million cubic meters of mine rock and tailings will be placed in the DWMF over the life of the

mine.

The development of the tailings management plan and yearly configurations was developed using

GoldTail®, developed by Golder Associates. The model consisted on developing yearly configurations of

mine waste within the designated footprint north of the process plant.

Tailings and waste rock will be transported in trucks from the process plant and from underground to the

DWMF area and spread and compacted to create a self-supporting structure. Figure 18.14 to Figure 18.18

show the yearly configurations of the DWMF. The growth of the deposit is expected to entail shallowly

sloped platforms from years 1 to 4 of the mine life. After year 4, the tailings will be stacked to the northeast

of the deposit. Mine rock will be placed preferentially near the outer slopes to improve erosion resistance

of the stack. The slope and platforms will be revegetated as completed.

Page 18.34




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.14 DWMF Deposit Configuration, Pre-mining and Year 1

Figure 18.15 DWMF Deposit Configuration, Years 1 and 2

Process plantand portal

El. 2460 m


El. 2460 m

El. 2437 m

3H:1V slope

PRE-MINING Year -1


El. 2460 m


El. 2460 m

El. 2453 m

3H:1V slope

Year 1 Year 2

El. 2447 m

3H:1V slope

Page 18.35




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.16 DWMF Deposit Configuration, Years 3 and 4

Figure 18.17 DWMF Deposit Configuration, Years 5 and 6


El. 2460 m


El. 2460 m

El. 2462 m

3H:1V slope

Year 3 Year 4

El. 2459 m

3H:1V slope

El. 2459 m


El. 2460 m


El. 2460 m

El. 2470m

3H:1V slope

Year 5 Year 6

El. 2459 m

3H:1V slope

El. 2466 mEl. 2459 m

Page 18.36




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.18 DWMF Deposit Configuration, Years 7 and 8

18.12.6 DWMF Water Management

Water management in the DWMF will consist of minimizing the amount of water infiltrating into the deposit

by intercepting the runoff from upstream of the basin and diverting it downstream through the use of two

HDPE pipes. The pipes will discharge into an 8,000 m3 capacity sediment control pond downstream of the

DWMF toe. The sediment pond will have an outlet to allow surplus water to be discharged to the La Veta

creek.

The surface of the DWMF will be sloped so any runoff flows off the surface before it infiltrates. Water that

infiltrates the DWMF is considered contact water and cannot be released downstream to La Veta creek.

Prior to construction of the DWMF, a rockfill underdrain will be placed along the main natural drainage The

underdrain is a 0.7 m deep, 2 m wide rockfill drain wrapped in non-woven geotextile to prevent particle

migration and clogging of the drain. Contact water collected by the underdrain will flow into a 500 m3

capacity seepage pond located just downstream of the DWMF toe. A pumping system will recycle the

collected seepage to the process water storage tank. Figure 18.19 shows the water management system.


El. 2460 m


El. 2460 m

El. 2480m

3H:1V slope

Year 7 Year 8

El. 2459 m

3H:1V slope

El. 2459 m

El. 2475m

Page 18.37




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.19 Water Management for the DWMF

A preliminary water balance by Golder (2013) estimated runoff flows for the project area for wet, average

and dry years (Figure 18.20). The catchment area between the portal pad and the La Veta creek basin

head is expected to produce approximately 6 liters per second; this discharge will be used for makeup

water for the process plant. Estimated runoff on the DWMF is approximately 2 liters per second and will be

conveyed to the downstream sediment control pond. Infiltration through the DWMF will be conveyed to the

seepage collection pond and pumped back to the process plant.

Non-contact waterdiversion pipes

Rockfill underdrain

Non-contact waterdiversion pipes

Sedimentcontrol pond

Rockfill underdrain

Seepagecollection

pond

Outflowto La Veta

Page 18.38




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.20 Runoff discharges for wet, average and dry years (Golder, 2013)

The makeup water demand for the process plant is estimated as 3.9 litres per second (14.3 m3/hr). The

makeup water will be fed from the seepage collection pond and the upper catchment area. The

preliminary water balance numbers indicate that there is a positive balance for wet, average and dry

years. The excess discharge from upstream of the portal pad will be conveyed to the sediment control

pond using diversion pipes and eventually discharged back to La Veta creek. Figure 18.21 shows the

water flows for average meteorologic conditions.

Page 18.39




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.21 Water Flows for Average Year Conditions

18.12.7 Soil and Topsoil Storage

After construction of the fill pads, approximately 120,000 m3 of saprolite material and 20,000 m3 of topsoil

material will have to be disposed of in the “Los Guapos” batición (Figure 18.22), an old hydraulic mining

area to the north-east of the DWMF area. This saprolite deposit will be a 40 m high fill with a 3H to 1V

slope partially filling the batición. Topsoil will be placed over the compacted saprolite platform and in a

small area just downstream of the saprolite deposit. An underdrain similar to the DWMF drain will be

placed along the main drainage of the deposit to avoid saturation. The discharge from this drain will be

allowed to freely flow downstream as it is not considered contacted water. Revegetation and sediment

control measures will be applied to ensure the adequate performance of the saprolite deposit.

Page 18.40




AMENDED NI 43-101 TECHNICAL REPORT Figure 18.22 Spoil Saprolite Storage Area

18.12.8 Preliminary Acid Drainage Considerations

Modified Acid Base Accounting (ABA) tests were performed by SGS Canada in 2013 to estimate the

potential for Acid Rock Drainage (ARD) from waste rock and tailings to be placed in the DWMF. Details of

the fifteen tests performed can be found in SGS (2013). The neutralization/acid ratios indicate low

potential for generation of ARD, with only three schist samples producing NP/AP ratios below 3.0.

Sulphide content of the samples with NP/AP < 3.0 was about 5%, whereas the remainder of the samples

had less than 1% sulphide.

Based on these results, it has been assumed that acid management in the project will consist of closely

monitoring the chemistry of the tailings stream, and if required, a small amount of lime could be added to

the filtered tailings prior to transporting them to the DWMF. Future work will entail additional testing of the

final tailings grind, kinetic testing and geochemical modelling for a more detailed assessment of long term

ARD potential.

Page 19.1




AMENDED NI 43-101 TECHNICAL REPORT 19.0 MARKET STUDIES AND CONTRACTS

19.1 Marketing Studies

No market studies have been undertaken for this project; however, the commercial product will be gold

silver bullion. During 2014 the price of gold has fluctuated in the $1,380 to $1,220 per ounce range. As of

this writing, the three-year average gold price is $1,450 per ounce. Thus, a price of $1,300 per ounce is

reasonable with respect to the current market.

A selling price of $1,300/oz Au has been used for this Feasibility Study.

The project has obtained pricing from major refineries and a quotation from Johnson Matthey has been

used for the refining costs. Red Eagle Mining plans to contract the transportation, security, insurance, and

refining to one refinery. This ensures the security of the gold on leaving the mine gate is the responsibility

of one party.

Page 20.1




AMENDED NI 43-101 TECHNICAL REPORT 20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR

COMMUNITY IMPACT

20.1 Introduction

This section is a summary of the Environmental Impact Assessment (EIA) which was submitted to

‘Corantioquia Tahamies’ (Environmental Authority) in February 2014. The EIA was based on the

technical content of the Preliminary Economic Assessment (PEA) issued by Red Eagle Mining in

October 2013. Upon the advice of the Environmental Authority an update to reflect the results of this

Feasibility Study, will be issued after the granting of the Environmental License. Sub section 20.14 of

this summary shows the key changes from the PEA to this Feasibility Study.

The (EIA) has been developed by Tetra Tech Colombia SAS, with the full support and participation of

the Universidad de Antioquia (University of Antioquia) and Fundación Universitaria Católica del Norte

(Foundation Catholic University of the North), who carried out the environmental baselines studies in

the disciplines of their respective expertise. The EIA is in compliance with requirements and

specifications provided in the Terms of Reference for Gold Mining Operation Projects issued by the

Regional Autonomous Corporation of Central Antioquia, ‘Corantioquia Tahamies’ (Environmental

Authority), and by the terms for submission of Environmental Studies issued by the Ministry of

Environment, Housing and Territorial Development – MAVDT, and International Standards.

The EIA has been prepared from the results of baseline studies conducted during 2012 and 2013, and

technical studies undertaken during 2013. A Preliminary Economic Assessment (PEA) was undertaken

by Mine Development Associates of Reno, Nevada, and published in October, 2013. At the same time,

a Plan de Trabajo y Obras (PTO), was completed in house, which followed the technical content of the

PEA, and was submitted to the competent mining authority, the Secretaria de Minas de Antioquia

(Secreminas, or Mines Department) in November, 2013. This PTO was subsequently approved in

August, 2014. Approval of the PTO is the first stage of the environmental licensing process.

20.1.1 Requirement For EIA

The EIA is a public document necessary for all projects that have potential impact on the environment

and local communities and stakeholders. The EIA is a necessary requirement as part of the process to

obtain an Environmental License to proceed with a project (Decrees 2811 of 1974, and 2820 of 2010)

Any mining operation that plans to excavate more than 2 million tonnes per year of rock must apply to

the National Environmental Agency, who rely initially on the local Environmental Authority, this adds to

the complexity of the process.

As the Santa Rosa Gold Project will excavate less than 500,000 tonnes per year, the EIA for the Santa

Rosa Gold Project will be reviewed and approved by ‘Corantioquia’ the local Environmental Authority,

which is the local government agency and competent authority in the Santa Rosa Municipality and the

Department of Antioquia, with a base in the town of Santa Rosa de Osos from where the approval will

be granted.

Page 20.2




AMENDED NI 43-101 TECHNICAL REPORT The EIA was submitted to the Environmental Authority on February 20, 2014

20.1.2 Objectives of the EIA

The EIA is a detailed technical document prepared using best practices and principles, to enable the

Environmental Authority and stakeholders to evaluate the feasibility of a project by identifying the

principal impacts and mitigation factors that may apply. It is also a requirement by international project

finance banks that the EIA be compliant to Equator Principles, and the applicable IFC Standards.

Red Eagle Mining requested Tetra Tech Colombia SAS to undertake the preparation of the EIA, based

on the environmental baseline studies conducted 2012 and 2013 by the two local universities noted

above. These institutions are highly regarded in the region, and by the Environmental Authority.

The EIA for the Santa Rosa Gold Project contains:

All technical aspects of the project, including mining and processing operations,

infrastructure and economics;

Baseline studies of the physical, biological, socio-economic and cultural

environments of the areas of influence;

Identification of specific natural resources that will be developed and / or impacted;

Identification and evaluation of the environmental impacts;

Establishment of an Environmental Management Plan (EMP), to prevent, mitigate,

correct, or compensate potential impacts of the project; and

Establishment of contingency, closure, monitoring, and follow-up plans.

20.2 Environmental Licensing

The Environmental License (EL) is the final stage in the mining project permitting process. The license

will contain two final parts:

Environmental Guidelines which are issued by the local Environmental Authority

within the Department of Antioquia, who are based in the nearby town of Santa Rosa

de Osos.

An EMP which is included in the EIA, and may require amendment to include any

further additions the Environmental Authority may deem necessary.

The final approval and issuance of the EL allows the Company to proceed with project construction

followed by operations, without any further permitting processes being required.

Page 20.3




AMENDED NI 43-101 TECHNICAL REPORT The construction period and operations during the life of the mine will be regularly inspected and

monitored by the Environmental Authority.

All statistics and technical aspects of the project must be submitted to Secreminas on an annual basis;

changes in the technical aspects must be approved by Secreminas, and in turn submitted to the

Environmental Authority for their approval, which may or may not warrant further changes to the EMP.

20.2.1 Process Stages of the Santa Rosa Gold Project

The following table and diagram show the necessary steps to obtain approval of the EIA, and the

granting of the EL:

Steps in the Environmental Licensing Process

Process Result

Submit a letter to the Antioquia Environmental

Authority (‘CORANTIOQUIA’) requesting Terms of

Reference for the EIA

Official Terms of reference for the area to be developed - Terms of

Reference were issued and accompanied by a Letter of

Confirmation from the Environmental Authority in July, 2013

Submit a letter to the Ministry of the Interior

(‘MININTERIOR’) requesting presence of ethnic

communities in the area

A certificate regarding ethnic communities presence/absence in the area

(black or indigenous communities) - Official Letter informing the

Company that there were no ethnic communities present in the tenement areas issued in June, 2012

Submit a letter to the Institute of Rural Development

(‘INCODER’)

An official reply about ethnic community property in the area - Official

Letter informing the Company that there were no ethnic communities present in the tenement areas issued in June, 2012

Submit a letter to the Institute of Anthropology and

History (‘ICANH’)

An official reply concerning any archaeological research requirements in

the area - Official Letter informing the Company that the archaeological authorization N° 3425 is granted on May 22, 2013.

Official Letter informing the Company that the final report and the

Archaeological Management Plan have been approved in January, 2014.

Submit a Technical Study (PTO) describing the

technical aspects of the planned mining and

processing development and operations to the

Secretary of Mines (‘SECREMINAS’)

An official approval of the PTO, which initiates the environmental

approvals process - Submission of the official document in

November, 2013, with official approval of the PTO issued by Secreminas in August, 2014.

Commence Environmental Base Line-EBL studies

Preliminary baseline assessment - Work commenced under contract

with Universidad de Antioquia (UdeA), and Fundacion Universitaria

Catolica del Norte (FUCN) in early September 2012 and was completed in May 2013.

Consolidate Environmental Impact Assessment-EIA

A final EIA document - Contract to complete the EIA was awarded to

Tetra Tech Colombia SAS in early December 2012, and was

completed and submitted to the Environmental Authority in February, 2014.

Submit EIA to Environmental Authority

An official submission receipt on February 20, 2014

An official letter informing the Company that the Initial Act for the

licensing process is issued on May 6, 2014

Evaluate the final EIA and make resolution of

approval

An official Environmental License for the Company allowing it to proceed,

without any additional permits, for construction and operation of the

project.

Page 20.4





The process to grant the EL includes a site visit to the project area following an initial review of the EIA

document. This process may be followed by any recommendations from the Environmental Authority

for incorporation into the EIA document as Addenda, prior to granting the EL. Recommendations have

been received and addressed by Red Eagle Mining during 2014 since submission of the EIA.

The Santa Rosa Gold Project is located in the mining concession contract B7560005. The aim of the

project is to extract and process the identified gold resource through a safe, efficient, and economical

underground mining operation, with minimal environmental impact. The EL will be specifically issued

for this concession contract.

This EIA Summary Report contains the key information and findings of the EIA for the Santa Rosa

Gold Project, providing information regarding the environmental impacts that the project may cause,

and most importantly, the methodology that the Company will use to control or mitigate these impacts

at all stages of the project.

20.2.2 Community Participation

The most significant issue in the EIA approval process is the recognition of a total social acceptance of

the project (which is not established by law, but of absolute necessity), such that all communities and

stakeholders have a full understanding of the Company’s intentions, and of the development of the

project. As a result it is important that the local communities and other stakeholders have been

involved in the preparation of the Santa Rosa Gold Project EIA.

The EIA was developed with the participation of communities and stakeholders of the Santa Rosa de

Osos Municipality, both in rural and urban areas. Information workshops were presented to explain the

project at the various stages of exploration, construction, operation, closure, and post-closure. These

information workshops were held in each village of the of the project's Area of Direct Influence (ADI),

and Area of Indirect Influence (AII).

Page 20.5




AMENDED NI 43-101 TECHNICAL REPORT In addition to these information sessions, participatory workshops were also held, where the

communities and other stakeholders including local government, community leaders, institutions,

environmental authorities and NGOs actively collaborated with the Company and their environmental

consultants, to identify the project impacts and assisted in formulating measures to mitigate and

manage the identified impacts.

Part of this process included visits to the project site area, to enable the participants to understand the

scale of the project and the minimal impact of the planned surface facilities to the surrounding area.

20.3 The Company

Red Eagle Mining is listed on the Toronto Stock Exchange Market with the ticker (TSX-RD), their

website address is: www.redeaglemining.com

The Company is engaged in the development of the Santa Rosa Gold Project, near to the town of

Santa Rosa de Osos in the Antioquia Department of Colombia. It operates with the corporate policy of

“Responsible Mining” carried out with a duty of care to the environmental and community.

This policy ensures that the Company conducts their operations and activities in harmony with the

environment and society, contributing to the development, welfare, and improvement of the quality of

life of their employees, their families and the communities local to their operations.

This is achieved through joint participation with communities and local government agencies in

decisions and actions that seek to provide effective, efficient, and sustainable solutions to the benefit

the communities and other stakeholders.

20.4 The Santa Rosa Gold Project

The Santa Rosa Gold Project is a gold deposit located approximately 70 km from Medellin, in the

Municipality of Santa Rosa de Osos. The municipality has over 400 years of mining history. The

project, including the Feasibility Study, PTO and EIA has focused on the San Ramon deposit located

within concession contract B7560005, which has a total area of 500 hectares (ha) and is situated

between the villages of La Cejita, Playa Larga, Rio Negrito, San Felipe, San Francisco, San Jose de la

Ahumada, San Ramon and Ventiadero (areas of possible direct or indirect influence).

The PTO and EIA reported the project surface facilities will occupy less than 60 ha of the concession,

comprising just 12% of the total concession contract area. The following maps show the location of the

Santa Rosa Gold Project in detail. The Project site can be accessed from Medellin via approximately

65 kilometers (km) paved road to within 8 km of the project area; the remaining distance is by an

unpaved road which is in good condition.

http://www.redeaglemining.com/

Page 20.6





Location of the concession contract B7560005, San Ramon deposit

Page 20.7




AMENDED NI 43-101 TECHNICAL REPORT 20.5 Baseline Studies

20.5.1 Baseline Study Consultants

The Company tendered the various aspects/disciplines of the baseline study work to local universities

in the Department of Antioquia. Two universities were selected based on their expertise in the required

disciplines.

The University of Antioquia, located in Medellin was selected to undertake surveys on the atmosphere

(air quality and noise), water quality and uses, hydrology (including detailed modelling), and

hydrogeology.

The Foundation Catholic University of the North, located in Santa Rosa de Osos was selected to

undertake surveys on flora and fauna, geotechnics and geology, soil and landscape, communities

including public health and epidemiology, and archeology.

These two universities are very highly regarded by the Environmental Authority in their fields; the

FUCN was particularly important due to its very local location and knowledge of the region and culture.

The baseline survey work was conducted over a period of 9 months for the EIA, but has continued

since the EIA study completion.

The baseline work was audited in detail by Tetra Tech Colombia SAS, who have been responsible for

the EIA, as well as the authors of all relevant documentation.

20.5.2 Areas of Influence

For the preparation of the EIA, the areas of influence have been defined as those geographic areas

where the project activities will be carried out, and which will have direct and indirect impacts on abiotic

(water, air and soil), biotic (flora and fauna) and socio-economic environments (social, cultural and

economic activities).

The influence areas established for abiotic and biotic environment are:

Area of Direct Influence (ADI): refers to sites that could be affected directly by

impacts resulting from the project activities; and

Area of Indirect Influence (AII): refers to sites that could be affected by indirect or

secondary impacts, occurring in distinct geographic areas, outside the ADI and occur

on a mid to long-term time frame.

Page 20.8




AMENDED NI 43-101 TECHNICAL REPORT For the abiotic environment an ADI with a total of 300 hectares was defined (more than 5 times the

project defined footprint). The criteria used to identify the area were watersheds, river basins, soil

usage, air quality modelling, and was based on an analysis of direct impacts on these aspects.

Additionally an AII with an area of 240 hectares was defined using the hydrological component

parameters and soil and air quality data.

For the biotic environment an ADI with an area of 75 ha was defined. This area was determined by

using the eco-systemic landscape and elements representing natural or artificial barriers such as

roads, basins, and watersheds. Additionally, an AII with an area of 268 ha was defined. This area was

determined by using the same eco-systemic landscape and elements as for the ADI.

Page 20.9




AMENDED NI 43-101 TECHNICAL REPORT The abiotic and biotic ADI and AII environments are shown in the following map:

Page 20.10




AMENDED NI 43-101 TECHNICAL REPORT The areas of influence defined for the socio-economic environment are:

Area of Direct Influence (ADI): refers to the presence of visible or perceptible impacts

directly from project implementation at the various stages. The locations concerned

are: the villages of San Ramon, Ventiadero, San Francisco, the Cejita, Rio Negrito,

San Felipe, Playa Larga and San Jose de la Ahumada and the rural jurisdiction of

Santa Rosa de Osos Municipality;

Area of Indirect Influence (AII): refers to those territorial jurisdictions (districts and

villages) associated with those communities that will be indirectly affected, in a

cumulative or in a residual way. The locations concerned are: Hoyorrico Villages,

San Isidro, the Malambo village and the urban center of Santa Rosa de Osos

Municipality; and

Specific areas of influence (Project Footprint): related to the 60 ha site where the

footprint of the mining project is located, and also where the existing Company camp

is located. The locations concerned are those that are part of the territorial jurisdiction

of the villages of San Ramon and Ventiadero.

Page 20.11





20.5.3 Abiotic Environment

Climate

The regional climate is classified as cold-very humid with an average temperature of 15°C and a

relative humidity of 85%. The project is located at an approximate altitude ranging from 2,300 to

2,500 m ASL and the total annual rainfall lies in the range of 1,800 mm and 2,300 mm.

Page 20.12




AMENDED NI 43-101 TECHNICAL REPORT Rainfall characteristics of the area establishes there are two periods of high rainfall in the region;

during the first half of the year the months with most rain are April, May and June, during the second

half of the year they are September and October.

Regional and Local Geology

Regional Geology

The project is located in the highlands of Santa Rosa de Osos on the Central Colombian Andes

Mountain Range in the northern center of the Antioquia department. It corresponds to an ancient

erosion surface, shaped by the actions of climate and tectonic uplift, to present now a relief of hills and

valleys colluvial-alluvial carved into a very thick blanket of alterations (80 - 90 m), produced by the

weathering of Antioquia Batholith granodiorites. The Antioquia Batholith has an area of 1600 km2 the

highest areas are located between 2,900 - 2,950 m.

The ADI of the project is located entirely in the highland area, which are alluvial plains formed with little

or no consolidated materials.

Local Geology

The local geology is characterized by a monotonous sequence of gradiorite to diorite rocks belonging

to Antioquia Batholith. Methamorphic rocks including amphibolite and metasediments are shown as

isolated hanging walls; diorite, dacite, felsite, aplite and pegmatite dykes are also present. A brownish-

red saprolite is well distributed along the area up to 40 m in depth, it is represented by granite altered

to clays. Soils are often 50 cm thick and rarely up to 2 m. Mylonite in a schistose shear-zone was

detected along some outcrops and adits.

Soils

The ADI of the project soils are mainly cold thermal soil, where the relief presents slight steep slopes

(25%-50%), with susceptibility to erosion and landslides. These are very deep soils with medium

texture, good natural drainage, highly acid with low fertility, high aluminum saturation, high phosphorus

fixation capacity, with varying pH.

The current land use in the ADI corresponds to upper secondary vegetation for forestry production-

protection land (5 ha) and land devoted to livestock ranching that are in weedy grass cover and low

secondary vegetation (14 ha).

Water

The AII of the project is located within the basin of the Guadalupe River which is the recipient of major

streams, such as San Francisco, San Ramon, and La Veta. However, the proposed mining activities

do not directly affect the Guadalupe basin.

Page 20.13




AMENDED NI 43-101 TECHNICAL REPORT The total ADI area established for the project is 3 km² , of which 76% is the San Ramon micro-basin

and the remainder belongs to the micro-basin of the San Francisco at 24%.

Page 20.14




AMENDED NI 43-101 TECHNICAL REPORT Water in the three streams in the vicinity of the project area has been monitored over three campaign

periods during 2012 and 2013 to cover seasonal variations.

The monitoring programme covered both water flow rates and water quality.

The following charts show the results from the three campaigns (Campana 1, 2 and 3), for water flow

rates (Cauda) expressed in litres/sec, and also a water quality graph:

La Veta Creek

Page 20.15




AMENDED NI 43-101 TECHNICAL REPORT San Francisco Stream

San Ramón Stream

Page 20.16




AMENDED NI 43-101 TECHNICAL REPORT Guadalupe River

Water quality graph

Page 20.17




AMENDED NI 43-101 TECHNICAL REPORT Campaign 1 (Campana 1) = 24-25th Sept 2012

Campaign 2 (Campana 2) = 28-29th Oct 2012

Campaign 3 (Campana 3) = 28-29th Jan 2013

Water quality values (as shown on graph above):

ICA Value (Valor ICA) Water Quality

0 - 20 Poor

20 - 50 Bad

50 - 70 Good

70 - 90 Very Good

90 - 120 Excellent

ICA VALUE = ICOMI, ICOMO, ICOSUS, ICOpH, ICOTRO

MI = Minerals

MO = Organic Materials

SUS = Suspended Solids

pH = Acidity

TRO = Trophic Pollution

Air

In order to establish the initial conditions, levels of gas concentration and particulate matter were

recorded at four different sampling stations (see map below) to establish the project baseline.

Air quality monitoring was conducted in the area of influence of the project, which determined that all

concentrations are below the maximum allowed levels under the reference conditions provided in the

Article 2 of Resolution 610 of 2010 MAVDT.

Page 20.18





The air quality index for certain pollutants monitored (PST, PM10, SO2, NO2 and CO), show that in the

project area air quality is good, indicating that the concentration of any pollutants in the atmosphere

presents a low level of air pollution and very low risk to the exposed population's health, also

anthropogenic activities in the area represent low impact on air quality according to the contaminants

monitored.

Page 20.19




AMENDED NI 43-101 TECHNICAL REPORT An important related aspect for the future operations is the wind direction. This was undertaken over

extended periods of time, and demonstrates a relatively continual dominance in the north westerly

direction, as shown by the wind rosette example from the survey:

This rosette shows the wind direction from source.

This result is important in that virtually all air movement past the operations will travel towards an area

(North West), where no communities or infrastructure are situated for many km.

Landscape

The project area has a landscape of highland and a relief type of ridges, hills, terraces and valleys,

where small watercourses form a network, draining the area’s rainfall to the major collector which is the

Guadalupe River. Four types of coverage in total are identified: forest plantations, stubble, pastures,

and agricultural crops.

NORTH

SOUTH

WEST EAST

5%

10%

15%

20%

25%

WIND SPEED

(m/s)

>= 10,8

8,1 - 10,8

5,3 - 8,1

3,1 - 5,3

1,7 - 3,1

0,6 - 1,7

Calms: 19,59%

Page 20.20




AMENDED NI 43-101 TECHNICAL REPORT The following map shows the respective areas of vegetation, agricultural cover, and types of soil cover

over the landscape surveyed for the baseline studies.

Page 20.21





Page 20.22




AMENDED NI 43-101 TECHNICAL REPORT Noise

Sound pressure levels were determined in the project area, by monitoring various points of the four

zones of the influence area.

These measurements indicated that the sound pressure levels managed to exceed standard maximum

levels but only in places where the main source of the cause is the activities of the local population.

The property poses no noise issues as it is located in a secluded area. This is also borne out by the

wind direction survey which shows that virtually all air movement past the operations will travel towards

an area of no communities or infrastructure for many Kilometers along that direction.

In addition, detailed noise and vibration models of the planned operations were carried out by a

consultant, which showed that although the mining and processing operations will have an impact, the

study provides mitigation measures to ensure minimal impact. The wind direction data has been of

benefit in this modelling process.

20.5.4 Biotic Environment

Regional Description

The project area is characterized by being in the bio-geographical region of the Andes, in the North

Andean province, which has a total area of 212,227 km2 of the country and has been extensively

transformed by human activity, an analysis confirmed in the study area of the project, where most of

the area has been used for the development of livestock, agriculture and to a lesser extent, for mining

activities.

As a result of these activities, the transformation of natural ecosystems in the area is clear. The once

large areas of dense gallery forests and lower mainland, were gradually replaced by crops such

Cyphomandra Betacea (tree tomato or tamarillo) and others less representative such as Zea mays L.

(maize). Most of the original intervention is mainly due to the clearing of natural areas to establish and

improve extensive livestock grazing.

Classification of the area of direct and indirect influence is located in the settlements in the Lower

Montane Wet Forest (with average temperatures between 12°C and 18°C, annual precipitation

average between 2,000 mm to 4,000 mm and a potential evapotranspiration of 0.5 to 0.25 mm/month)

which is located in the temperate latitudinal subtropical region.

The Santa Rosa Gold Project is inside the “Selva Andina" formation, which corresponds to areas

located in the higher parts of the mountain ranges. In general this vegetation is characterized by lower

variety of species compared with the formations at lower elevations. The amount of vegetation is also

comparatively less.

The ADI and AII are within the Great Tropical Rainforest Biome, which occupies an area of 105,632 ha

of the country and characteristically includes two main types of climates - warm wet and moist warm.

Page 20.23




AMENDED NI 43-101 TECHNICAL REPORT Flora

The representative vegetation of the Santa Rosa Gold Project is common in Colombian

agroecosystems and is mainly represented by the clean pastures and weed types such as the tree

tomato. Regarding the ecosystem importance, the most representative ground covers are low dense

upland forests, gallery forests, peatlands and secondary vegetation.

The following tables present the types of land cover for the respective types for both AII and ADI

categories:

Land coverage present in the biotic AII

Ground Cover Area (ha) AII

Gallery Forest and riparian 5.2

Mainland lower dense forest 1.3

Corn 0.3

Weedy pastures 54.8

Clean pastures 73.8

Conifers plantation 32.6

Infrastructure 5.9

Rivers 0.3

Barren Land and degraded 1.9

High secondary vegetation 23.6

Lower secondary vegetation 43.8

Mining extraction sites 2.3

Wetlands 4.6

Total 250.4

Land coverage present in the biotic ADI

Ground Cover Area (ha) ADI

Corn 0.2

Weedy Pastures 37.2

Clean Pastures 8.1

Road network, railways & associated land 4.0

Barren Land and degraded 2.6

High secondary vegetation 16.3

Lower secondary vegetation 22.0

Wetlands 2.1

Total 92.5

Page 20.24




AMENDED NI 43-101 TECHNICAL REPORT Two endemic and considered ecologically valuable species are located in the project area. These are

roble (type of oak), and helecho sarro (type of fern) – see below.

Permission to remove a small identified quantity of these species has been submitted at the National

Environmental Authority (Ministry of Environment) and a specific compensation plan has been

developed according to the national forestry compensation regulations, outlining a tree replanting

program.

Oak (Roble - vulnerable) Fern (Helecho sarro – forbidden)

Fauna

In the baseline study area, indirect sampling involved a review of available information about the

species potentially present in the (ADI and AII) study area. The review included scientific papers

regarding wildlife herbs, birds and mammals present in the Antioquia Department. Another indirect

method was finding traces of fauna such as faeces, skins and footprints; these were complemented by

community surveys.

Amphibians

Investigations show the presence of three amphibian species belonging to two distinct families, and

one common frog in the ADI. It is noteworthy that all species are endemic, and two species were found

to be in the threatened category.

The main habitat of this wildlife is found in is wetlands or areas with some water accumulation, peat

areas, forests and low secondary vegetation.

Reptiles

Four species of reptiles were identified in the different habitats in the baseline study area and showed

a preference for the secondary vegetation and forest coverage. This is due to the fact that these are

the habitats in which the greatest amount of food supply (insects, birds, small mammals and

amphibians) were observed, because of the varied vegetation structure.

Given the wide range of habitats and microhabitats the species were easily observable in many

different locations particularly under dead vegetation, and soil.

Page 20.25




AMENDED NI 43-101 TECHNICAL REPORT Birds

Within the ADI of the project, a total of 69 bird species corresponding to 35 families and 13 orders were

recorded, the most diverse group were of passerine birds or songbirds. Hummingbirds and swifts were

commonly recorded. The flight condition of the birds and their constant movement allowed registry

throughout the area of study and also a large number of active nests suggesting reproductive

conditions of many species at this time of year were recorded.

With reference to the International Trade in Endangered Species of Wild Fauna and Flora Convention,

it was found that 11 species were listed that are not necessarily threatened with extinction, but in which

trade must be controlled.

In the area directly affected by the project area, some of the best mulches to establish some bird

communities such as passerine and hummingbirds are found. Forested areas are well represented,

allowing birds to be constant movement using riparian vegetation and canopy for movement corridors.

The forest canopy provides more resources for feeding and protection from predators for many of the

species recorded and have aided bird settlement

Secondary vegetation and forests were the habitats that presented greater species diversity. Typically

there are rivers, wetlands and grasslands with a similar richness of species. The least favourable

coverage for birds are crops, conifer plantations, bare land and the urban fabric, typical of the project

area.

Mammals

In the area directly affected, seventeen species of mammals, distributed in eleven families and six

taxonomic orders were identified.

Within these mammals are: fara, chucha, opossum, armadillo, gurre, several species of bats, fox,

weasel, squirrel, house mouse, mouse bush, hedgehog and rabbit.

20.5.5 Socio-Economic Environment

Population Dynamics

In the ADI of the project, there is currently a population of 1,077 individuals, of these, 56% are female

and 46% are male.

The prevailing trend of female over male population is evident in almost all the villages of the ADI,

except in the village of San Jose de la Ahumada.

Page 20.26





Infrastructure and Sanitation

Public Services

Public service coverage in the ADI are as follows:

water supply 36%;

sewerage 3%;

electricity 97%;

gas 74%;

telephone landlines 1%;

cellular 98%;

sanitation and collection services 31%.

Page 20.27




AMENDED NI 43-101 TECHNICAL REPORT Solid Waste Disposal

The predominant mechanism for the disposal of solid waste in the ADI is by burning, this represents

49% of the population in the ADI.

Collection is then followed with 17%, followed by disposal on open ground and burial which correspond

to 15% and 16%, respectively.

The sanitary landfill is the least used mechanism for solid waste disposal with only 6%.

Wastewater Disposal

The predominant type of disposal of wastewater is in the open field or open pit with 85% coverage, the

septic tank system represents 15% of use in management of wastewater discharges.

Domestic Water Catchment

There are two predominant systems in the ADI water catchment area, the predominant one is the use

of springs with 74%, followed by piped system with 26%.

In the village communities, San Ramon, Playa Larga, Rio Negrito, there is only spring water collection.

Page 20.28




AMENDED NI 43-101 TECHNICAL REPORT In San Francisco and the Cejita village, there are three types of available systems (springs, piped and

rainwater).

Transportation

The means of transport used by the inhabitants of ADI are predominantly public motor transport and to

a lesser extent motorcycles, bicycles and horses.

Social Services

Social security coverage is a subsidized system where 74% of the inhabitants of ADI are enrolled. The

remaining members of the population obtain cover through a contributory scheme.

According to the population survey and health records the most prevalent diseases within the

Community are respiratory, digestive and cardiovascular.

Settlements

Types

There are two types of settlements present in the communities located in the ADI, these correspond to

widespread and semi-urban types. The San Ramón village is only the community which can be

regarded as a semi-urban settlement.

Page 20.29





Properties

In the ADI small farm type properties are predominant, these being small farms of short and medium

land area (3 ha to approximately 600 ha).

The following map shows the distribution of properties with respect to the project area.

The remoteness of properties to the Project area should be noted:

Page 20.30





The predominant types of homes are privately owned houses, characterized by family legacy, which is

developed from the main economic activities of agriculture and small scale artisanal mining.

Page 20.31




AMENDED NI 43-101 TECHNICAL REPORT The photographs below show views from the south looking up and across the San Francisco valley,

which is the immediate west of the Project area. Careful design and attention was made to the Project

layout, to ensure this beautiful valley remained totally undisturbed. A high ridgeline separates this

valley from the project area, completely screening it from view.

Most of the properties are relatively old, with a majority built over 20 years ago:

50% of homes in the ADI were built over 20 years ago;

8% of homes were constructed 16 – 20 years ago;

9% between 11 to 15 years ago;

15% have been built between 5 – 10 years ago;

9% were built less than 5 years ago.

The residents of the ADI are native inhabitants of the municipality and the Antioquia region, and they

have deep roots in the area.

In the ADI, 23% of the population has lived over 25 years in the region and 20% of the population has

lived there between 11 to 25 years.

Page 20.32




AMENDED NI 43-101 TECHNICAL REPORT The predominant wall material in the ADI houses is brick with 71%, followed by mud walls with 22%,

the precast type represents 3%.

Social Organization

Industry

The principal industries in the Municipality of Santa Rosa de Osos are primary and include:

Dairy cattle;

Forestry;

Poultry farming;

Production of tomato trees (tamarillos);

Pig farming;

Socio-Cultural Identities

The communities in the ADI are defined as rural, using the land for subsistence. Some families work

full-time for farmers/land owners, some for their own production.

Historically, a small part of the settled community has engaged in artisanal mining activities. The

community considers themselves predominantly as farmers and only a small number are artisanal

miners.

Sociocultural System and Practices

A basic sociocultural system is widespread throughout the population of the Municipality of Santa Rosa

de Osos. They have strong religious beliefs (in which the Catholic faith is predominant), and visiting a

local church on a regular basis is a common activity.

It is a rural and urban population with good manners, friendly and cordial. They have patriarchal

relations with their family and although being relatively dispersed from each other, (especially in the

rural area), they maintain strong communication with neighbours from their community.

Archeological Heritage

The archaeological study had the support of the Colombian Institute of Anthropology and History. This

resulted in a report and management plan which was approved by the Institute in January, 2014.

Page 20.33




AMENDED NI 43-101 TECHNICAL REPORT The archaeological study was based on the analysis of landscape and its components, seeking to

establish the basic units on which humans developed their actions in the past, changes produced by

them over time and the potential that the environment offered. The result of the archaeological study

fieldwork led to the conclusion that certain remains found in the region, do not affect the area of the

footprint of the Project.

As a preventive measure an Archaeological Management Plan was formulated for implementation

during the stages of construction, and operation of the project and will be used if necessary.

20.6 Corporate Social Responsibility

20.6.1 Introduction

The Company has a key corporate strategy of social responsibility, this is to ensure the Company´s

reputation, social license to operate, and long term sustainability. This is achieved by adopting at all

times a transparent and ethical behaviour and maintaining harmony and balance with the local

communities.

This corporate strategy includes maintaining relations with all relevant stakeholders and communities,

by the implementation of the social management plan.

It is supported by the following objectives:

By undertaking all activities with the highest regard for social and environmental

responsibility;

By strengthening relationships through shared responsibility, based on dialogue and

participation with all stakeholders;

By promoting collective processes of regional development; and

Acting within a framework of respect for human rights.

20.6.2 Relations with Stakeholders

During 2012, Red Eagle Mining commenced the identification and implementation of a stakeholder

relations plan. This was designed to generate communication channels to establish, maintain and

improve relations with the various stakeholders, as well as timely and quality responses to their

requests and concerns.

The following actions and programs have been established as part of this stakeholder relations plan.

Page 20.34




AMENDED NI 43-101 TECHNICAL REPORT Information Office

An information office has been established in the town of Santa Rosa de Osos with the purpose of

allowing ease of communication between the local population and the Company.

The principal functions of the office are:

Receive, address and/or direct with the Company any queries, suggestions, opinions,

and/or claims from citizens;

Monitor and control these requests in order to ensure they are resolved quickly and

adequately;

Serve as an information and guidance centre for the Santa Rosa Gold Project and

the Company´s activities; and

Serve as a future recruitment centre for advertising job opportunities, receiving

applications and also to invite interest from suppliers and contractors for the Project

needs.

Education

Communicate and educate stakeholders about responsible mining activities and modern mining. It has

been a key strategy to seek community and stakeholders’ trust in the Company’s present and future

activities. This process has been facilitated by the following activities:

Red Eagle Mining provides a monthly contribution to a regional environmental and agricultural

newspaper called “Antioquia Si Tiene Norte”. This is a publication lead by Santa Rosa de Osos, where

all municipalities of the north region of Antioquia show their efforts in protecting and preserving nature

and agricultural best practices. It has a circulation of approximately 50,000 copies monthly. The page

addresses and answers the most common concerns stakeholders have regarding the project;

The Company has also commenced publication of a corporate newsletter with the aim to regularly

inform all of the stakeholders of the development and progress of the Santa Rosa Gold Project, and

includes activities related to social and environmental initiatives and best practices; and

A 20 page activity and colouring book aimed at educating children about the world of modern and

responsible mining, including for example: What is mining? How is the ore processed? and other

related social and environmental practices.

Page 20.35




AMENDED NI 43-101 TECHNICAL REPORT 20.6.3 Consultation

Since commencing the Environmental Baseline Study (EBL) in September, 2012, the Company has

undertaken 67 information and consultation meetings and workshops with all stakeholders of the Santa

Rosa Gold Project including: local government, dioceses and churches, economic sectors,

environmental and mining authorities, NGOs, municipality teachers, youth movements, university

students and each community of the ADI. Many of the consultations included site visits. The aim of the

consultation and information processes is to seek input from key stakeholders on the proposed mine

plans and discuss any questions or concerns they have about the planned future mining operations.

The table below summarises the meetings and workshops conducted, with each stakeholder and

associated sub-groups.

Page 20.36




AMENDED NI 43-101 TECHNICAL REPORT Group Stakeholder Information and Consultation Meetings

National Government Ministry of Environment Project Information and Consultation

Meetings

National Institutions Chamber Of Mines Project Information and Consultation

Meetings, including a Site Visit

International Institutions Canadian Embassy Project Information Meetings

Central Government Secretary of Mines

Project Information and Consultation

Meetings and Site Visit

Environmental Baseline results

information meetings and feedback.

EIA Summary Information Meetings and

feedback

Central Environmental

Entity

Environmental Authority

(Corantioquia)


information meetings and feedback

Local Government and

Entities

Santa Rosa De Osos Mayorality

Exploration activities information

meetings






feedback

Councillors






feedback

Presidents Community Action Boards Project Information and Consultation


Economic Sector




information meetings and Feedback

Local Environmental

Entity Corantioquia Tahamies






feedback

Church Diocesis of Santa Rosa De Osos




feedback

Education

Universities Project Information and Site Visit

Local Teachers Project Information and Consultation


Page 20.37




AMENDED NI 43-101 TECHNICAL REPORT Group Stakeholder Information and Consultation Meetings

NGOS

Central NGOs Project Information and Consultation


Local NGOs




feedback

Protected areas meeting

Communities

Area of Direct Influence

Exploration activities information and

education meetings and feedback


Meetings

“My Dream Territory”, Social Baseline

identification workshop

Impacts Identification and Management

Plans workshops




feedback

Area Indirect Influence






feedback

Young People Of Santa Rosa De

Osos





Mining workshops


feedback

20.6.4 Social Management Plan

Red Eagle Mining has implemented a social management plan, which includes practices and initiatives

involving local communities and other stakeholders, and is summarized as follows:

Schools Assistance Workshops

Regular monthly workshops are conducted in the five ADI schools to address different topics

such as care of the environment, water use and management, waste management and to

promote positive social attitudes, values and behaviour, through the use of video and other

media where the pupils are able to actively participate.

Adopt A Tree Program

Page 20.38




AMENDED NI 43-101 TECHNICAL REPORT Caring for the environment is the social responsibility of every individual. The “Plant-A-Tree

Program” has been developed to teach children of our ADI to actively participate in the

improvement of their villages by planting trees around their schools and adopting them.

Approximately 200 native trees have been planted to date.

Mining for Non Miners Course

In association with the Mining Association of ANDI (National Industrial Association), one day

courses are held to provide the community and stakeholders from a non-mining background,

with a basic, but comprehensive understanding of the mining industry and associated best

environmental and social practices.

Computer Education Program

The Company has established an alliance involving the education department of the local

government, the local Catholic University Foundation and the La Cejita School (located

closest to the project site), to implement a virtual learning education program to assist

interested parties from nearby villages (La Cejita, Ventiadero, San Francisco, La Cabaña,

Aguaditas and San Isidro).

The programs will comprise courses at three levels of computer literacy and competency.

Infrastructure Investment

The Company is also committed to assist development and the improvement of the quality of

life of the communities through sustainable infrastructure projects. To date, these have

included:

o Sports Facility in San Jose de la Ahumada School

In conjunction with central and local government and the local community, the

Company provided some of the materials necessary for the construction of the sports

facility at a local village school. This facility will be used by twelve villages, benefiting

more than 1,200 people.

o Sports court in San Francisco village

Together with community the Company has contributed construction materials for the

sports court in San Francisco village, benefiting 55 families.

20.7 Natural Resource Requirements

The following table sets out an estimate of the requirements for natural resources that the project may

need during construction and operations of the Santa Rosa Gold Project:

Page 20.39




AMENDED NI 43-101 TECHNICAL REPORT Natural Resource Use Technical Specifications

Surface Water

Water

The water supply project for the camp will be held

through two water concessions granted by

Corantioquia from the Monte San Isidro and Quebrada

La Negra streams.

The camp has a treatment system to purify water for

human consumption.

Additionally, a by-pass pipe system will be established

to collect water from the La Veta spring, which will be

used, for process plant make-up water.

Underground Water Domestic and industrial

water

The project will use groundwater that is available at

surface from the Hilo Azul mine shaft developed

during the former mining activities. In addition water

inflow resulting from underground operations will be

collected, clarified in settlement traps, and used for the

mine and process plant raw water supply.

Waste Disposal

Accommodation Camp

Waste discharges from sanitation, washing of clothes,

kitchen use and cutting drill core, will be processed

through independent treatment systems comprising

grease traps, settling tanks, septic tanks and

anaerobic up-flow filter, with discharges directly to

several points in the San Ramon stream bed.

Sanitation from the

operations

During project operation two integral sanitary facilities

will be installed in the area near to the mine site and

processing plant with sufficient capacity to meet the

wastewater generation for 100 people per day (50 per

location).

Industrial waste water from

the mining activities

Implementation of compact wastewater treatment

systems comprising three stages: primary

sedimentation; biological treatment and secondary

sedimentation.

Additional exploration activities will use a drilling mud

treatment system consisting of three in-line tanks:

2000, 3000, and 5000 litres capacity, to settle and de-

silt the water.

Water from the underground mine will be pumped into

a clarifier and will be re-used for underground drilling

and other processes.

These industrial discharges will be fed to the La Veta

creek after treatment complying with Colombian

legislation.

Riverbed Occupation

The project will partially occupy the La Veta creek, where dry and barren tailings will be

deposited.

Minor construction works will be conducted on necessary access road sections and

drainage channels.

Page 20.40





Building Materials

Cement and rebar Most will be sourced and transported from Medellín

city.

Aggregates

There are no identified quarries in proximity to the

project site for construction aggregate and other rock

materials. Therefore the waste material

(granodiorite),resulting from the excavation and

construction of the main access tunnel during pre-

production development, will be used as a source of

raw material for aggregate and other rock construction

requirements.

Waste rock from underground not used for

construction materials or underground backfill, will be

disposed in the tailings storage facility located near to

the processing plant. This rock will also be available

for crushing and use on road maintenance around the

ADI villages.

If suitable construction material is not available from

the project for any reason, this will be sourced from a

regional provider who has the proper mining and

environmental permissions.

Excavation and

Earthworks

Topsoil removal and

storage, for re-use later for

progressive reclamation and

landscaping through to final

closure.

Excavations and storage will take place within the

footprint of the project. It is estimated that 10,000 m3 of

soil will be removed and retained for future use.

A borrow material area has been defined, located

south east of the dry tailing deposit and will make use

of an existing depression in the surface topography

caused by historical mining. It has sufficient volume to

contain the volume planned to be stored.

Forestry Timber market

The site consist of clean pastures, weedy pastures,

high and low secondary vegetation, no logging

operations are necessary or planned in the project

footprint.

Atmospheric

Emissions

Mobile and stationary

sources during construction

and operation phases.

No specific permit is required for air emissions.

Page 20.41





Solid Waste

Building stage. Estimate = 25 kg/day (50 persons)

Operational Stage. Estimate = 60 Kg/day (135 persons)

Domestic

For integrated solid waste management it is planned

to identify, separate and sort the waste, with the

various types being stored in containers as follows:

- grey trash can: plastic, paper, cardboard, cans and

glasses;

- green trash can: wrappings, napkins, sweeping

debris, or anything that cannot be re-used again;

- red trash can: hazardous waste (grease, cotton

waste, oils, contaminated containers);

- organic waste will be given away to workers for use

as compost;

- ordinary waste will be managed by the municipal

collection services for disposal at the local nearby

landfill.

Industrial

Hazardous and special waste will be stored in

temporary locations preventing contamination of soil

and water bodies for subsequent delivery or collection

to a hazardous waste management company properly

certified by the environmental authority.

The dry and barren tailings material will be deposited

in the dry waste management facility (DWMF) located

east of the processing plant, and will be progressively

covered with topsoil and revegetated from the northern

end retreating to the south of the DWMF.

20.8 Environmental Assessment

20.8.1 Assessment Process

The Environmental Impact Assessment is a process and tool used to identify any positive or negative

impacts on the environment and social framework caused by a Project.

The process involves an assessment of environmental impacts for the Santa Rosa Gold Project and

considers the inter-relationship between the biotic (flora and fauna), abiotic (water, air, soil), social

(economic and cultural) and the project activities, in all of the development stages; exploration,

construction, operation, closure and post-closure.

The table below identifies the impacts.

Page 20.42




AMENDED NI 43-101 TECHNICAL REPORT Component Element Environmental Impact

Abiotic

Groundwater

Physical, chemical and microbiological alteration of water

quality;

Alteration in the availability of water

Surface Water

Physical-chemical and microbiological changes in the water

properties’

Sedimentation in natural bodies of water;

Change in surface drainage network;

Change in the available water supply.

Air and Noise

Change in particulate matter concentration in the air;

Change in local measured gas concentrations;

Change in sound pressure levels.

Soil

Change in soil usage;

Alteration in the physical-chemical and biological

properties;

Geotechnical stability alteration.

Biotic

Fauna

Changes in the composition and structure of animal

communities;

Change of habitat;

Modification of biological paths;

Temporary dispersion of animal species.

Flora

Alteration of distribution, structure and floristic composition

of natural vegetation cover;

Fragmentation of natural vegetation cover;

Alteration of the forest reserve areas.

Page 20.43





Socio Economic

Community Engagement

Compliance level in the population;

Generation of employment expectations and social

benefits;

Change in community management;

Change in institutional management;

Change in community empowerment;

Changes to the environmental understanding of the

community;

Presence of guidelines with land use planning;

Change of perception about mining exploitation;

Alteration in power relations;

Change in institutional community engagement;

Changing roles of responsibility by local governments;

Conflicts generated by the resistance to changes;

Conflicts generated by expectations.

Landscape Alteration of scenic beauty and landscape quality;

Changes in visual and aesthetic perception.

Public Health

Change in Accident rates;

Health alterations caused by particulate matter;

Alterations in accidents and incidents rate (road, social and

environmental)

Public

and

Social Services

Changes in availability and accessibility of services;

Changes in connectivity and population mobility;

Alteration of existing road infrastructure;

Changes to private property due to traffic on existing roads.

Archaeology Changes in archaeological heritage.

Culture

Sociocultural changes in community patterns;

Social changes in security of the population;

Changes in education, training and technology transfer;

Changes in traditional cultural symbols of the population;

Changes in the traditional conception of the world.

Population Dynamics

Changes in the rural population;

Changes in population dynamics by displacement and

resettlement;

Changes in population dynamics of ethnic communities;

Variation of the floating population.

Page 20.44





Economy

Expectations generated by income from royalties;

Change in gross domestic product;

Variation of local and national tax revenue;

Change of the artisanal mining;

Change in supply, demand of goods, services and

production chains;

Changes to the traditional economic model;

Labour migration;

Change in labour supply;

Change in unemployment rate;

Variation in quality of life;

Change in the average income of the community;

Change in land tenure;

Change in land market value;

Change in living costs.

20.9 Impact Assessment

These potential impacts are quantified by applying qualifying criteria to define and classify the impacts

with a significance of High, Medium and Low.

These impacts, when considered with project development activities (exploration, construction,

operations and closure), resulted in 451 interactions, which are evaluated under certain qualification

criteria with the following results:

Page 20.45




AMENDED NI 43-101 TECHNICAL REPORT The analysis shows the major impacts which are identified by a negative score and shows areas of the

environment that may be impacted by the project and where mitigating measures may be necessary.

The positive scores show impacts of a positive nature that will improve the environment and socio-

economic conditions in the region.

The conclusions of the impacts to the socio-economic environment are summarized below:

The assessment was carried out with the participation of the local communities and

other stakeholders, who were actively involved in the identification of the possible

impacts. Some communities could have the potential for changes in their lifestyle,

these have been identified and recognized and can be managed effectively with no

negative impacts;

Culturally the local communities are willing to accept change and have good potential

to benefit themselves, through the positive impacts of the project;

The project footprint lies within an area owned by a single landowner, therefore

limited effect on land or property tenure; and

There are no impacts related to displacement of the population and any impacts on

their way of life are minimal. The remote location of the project in relation to the local

communities and population, assists in minimizing any potential impacts.

It can be concluded by the quantitative and qualitative analysis of the identified potential impacts, that

these can be managed through comprehensive, effective and well managed environmental strategies.

There is a high potential for the social environment to be developed and benefit from sustainable

programs in the Project execution area.

For every negative impact identified in the EIA, environmental management measures are detailed in

the complimentary EMP.

20.10 Environmental Zoning

For the EIA, an environmental zoning process is carried out, using methods prescribed by the

Environmental Authority, to determine the areas of influence and the identification of possible

environmental impacts for the development of the project in all of the future stages.

The results of the environmental zoning for Santa Rosa Gold Project are as follows:

Page 20.46





20.11 Environmental Management Plan

20.11.1 Introduction

An EMP is an obligatory part of the process required by the Environmental Authority when granting the

Environmental License. The EMP details the proposed management policies, strategies, monitoring

programs and actions to be implemented by the Company to prevent, mitigate, correct, or compensate

any impacts that may occur during the entire lifespan of the project.

Each of the identified impacts for the project will be managed through effective environmental and

social action programs.

20.11.2 Management Programs

The management programs shown below have been developed to mitigate any negative effects of the

project and strengthen the positive ones in each of the development stages of exploration,

construction, operation, closure, and project post-closure.

The project EMP includes 22 programs for environmental management of the abiotic (water, air and

soil), biotic (flora and fauna) and socio-economic (social, cultural, economic and archaeological)

environments affected by the project. Each of these programs is explained in a practical manner and

clearly identifies their objective and at which stage of project development they occur. An action plan

for each program will include the necessary human and financial resources, technical details of the

measures to be used, logistical requirements, scheduling and reporting functions.

Page 20.47




AMENDED NI 43-101 TECHNICAL REPORT Tracking and Monitoring

Included in the EMP is a Tracking and Monitoring program (TMP) to evaluate the efficiency and

effectiveness of management measures adopted in the EMP.

These programs allow the tracking of the commitments and obligations established in the EMP in order

to verify compliance. They serve as a tool for the Environmental Authority to monitor the commitments

throughout the lifetime of the project, to assess the efficiency and effectiveness of the measures and

make any corrective actions should any be necessary.

The TMP seeks to design and establish the tools to check the status of compliance programs proposed

in the EMP, to analyze trends in the quality of the environment in which the project is developed, to

validate the anticipated environmental impacts and verify the effectiveness of EMP measures.

The TMP allows an ongoing assessment of the measures through monitoring indicators, to identify the

need for any adjustment, if required, to any changing conditions as they arise during the development

of the project.

The TMP ensures that the objectives of the programs defined in the EMP are effective and continue to

be valid for the life of the project.

The tables below show a summary of the programs with the corresponding monitoring program for the

Santa Rosa Gold Project:

Page 20.48




AMENDED NI 43-101 TECHNICAL REPORT Abiotic System

Environment Program Format Monitoring and Tracking

Abiotic (Physical)

EM soil management Management of soil

resources

Monitoring of erosion and

sediment yield

EM dry and barren tailings Dealing with dry, sterile

tails

Monitoring the

morphological restoration of

mining areas

EM of water resources and

erosion control works

Management of domestic

wastewater from camps

and mining areas

Management of domestic

wastewater from camps

and mining areas

Management of industrial

wastewater from the

access portal

Management of industrial

wastewater from mining

operations

Management of drilling

wastewater (aqueous

slurry)

Management of drilling

wastewater (aqueous

slurry)

Water management of

the deposit of dry and

barren tailings

Tracking and Monitoring of

runoff water and tailings -

sterile

Management of runoff

water (rainfall)

Management of

watercourse crossings

Monitoring of watercourse

crossings

Management of water

catchments (domestic and

industrial)

Monitoring catchment

management

EM of air resources

Management and control

of gases and particles

Monitoring gases and

particles

Management and noise

control

Monitoring of decibel level

Dealing with noise and

vibration from the use of

explosives in mining

production process

Alarm system and

prevention with the use of

explosives on site

Solid waste EM

Domestic and Industrial

Solid Waste

Management Monitoring of Domestic and

Industrial Solid Waste EM fuel and chemicals Fuel and chemicals

management

EM blast and explosives Blast and explosives

management

Page 20.49




AMENDED NI 43-101 TECHNICAL REPORT Biotic System


Biotic

Wildlife management

programs

Wildlife management,

forestry and revegetation

of disturbed areas

Monitoring the flora and

fauna (endemic, critically

endangered or vulnerable,

species, etc.)

Fauna and flora

preservation program

Conservation of

vegetable and animal

species with some level

of threat either, endemic,

forbidden, unregistered

or unidentified

Program and protocols for

wildlife species

management and rescue

Protocols for

management and rescue

of wildlife species.

Program for developing and

foster of ecosystems and

flora and fauna species

affected by the project

Development of

ecosystems and flora

and fauna species

affected by the project

Education and training

program for staff involved in

the project


staff involved in the

project with emphasis on

ecosystems and species

of flora and fauna of

particular concern

Compensation Program

Use of topsoil covers

such as ligneous plants

Monitoring of Compensation

Program

Use of vegetation cover

that do not involve tree

planting

Affectation of plant

species with some level

of vulnerability and/or

widespread use

Impact on wildlife

Page 20.50




AMENDED NI 43-101 TECHNICAL REPORT Socio-Economic System


Socio-Económic

Community information and

involvement program

Information and

involvement program

Monitoring the attention of

any requests or claims and

invite participation and

timely information to

communities


program for staff involved in

the project


staff involved in the

project

Monitoring the effectiveness

of the Social Management

Plan programs

Support program for the

institutional management

capacity.

Support for institutional

management capacity

and strengthening of

economic activities

Local labour hiring program.

Employment hiring

labour, local services

and general goods

Program of training,

education and awareness of

the community surrounding

the project

Training, education and

awareness of the

community surrounding

the project

Impact program for third

parties

Management of impacts

and social risks

Monitoring the impacts and

social conflicts generated

by the project during

different stages

Social compensation

program Social compensation

Monitoring the population

that could be affected by

the project

Preventive Archaeology

Program

Archaeological findings

Archaeological monitoring Archaeological

Monitoring

Promotion

Page 20.51




AMENDED NI 43-101 TECHNICAL REPORT Management Measures

Abiotic System

Name Objective Activities

Soil Management

Protection measures to prevent

erosion, loss of organic layers and

topsoil, pollution prevention and

any instability.

Effective disposal of products

generated during the mining

operations (waste rock) and

processing (tailings), to prevent

slippage and erosion.

Providing adequate providing

drainage for runoff water.

Minimizing visual impact, by

effective storage planning and

control.

Adequacy of access roads

Perform maintenance and

improvement of roads and drains

Construction of drains

Management of water runoff and erosion

control

Avoid steep slopes

Flow dissipation

Silt traps

Drainage channels

Grease traps

Collection sumps

Interceptor channels

Soil protection

Careful layer control

Minimise passage of mobile

equipment

Control soil with the lowest moisture

content

Waste Management

During the construction phase of the

portal and main underground access

ramp waste material will be used as

material for the construction of the

project infrastructure, roads, tailings

dam embankment, etc.

During mining operations waste rock

may be used for backfilling stopes to

allow safe and effective mining

operations

Transport of any materials on public

roads will be covered

The tails generated in the gold recovery

process will be filtered to remove

moisture and it is planned for

approximately 50% be returned

underground for backfilling of stopes

and 50% in the tailings storage facility

Domestic and Wastewater

Management

Mitigate pollution of streams to the

San Ramon and La Veta creeks by

domestic discharges

To maintain the water quality of the

San Ramon and La Veta streams

The domestic wastewater from the

domestic and operational infrastructure

will be treated in dedicated in WTP’s,

with a controlled and managed

discharge of the treated water to the

existing streams

Page 20.52




AMENDED NI 43-101 TECHNICAL REPORT Name Objective Activities

Management and Control of

Underground Water

Reduce the potential environmental

impacts caused by the lowering of

groundwater levels due to

underground activities and also

provide safe operating conditions

during mining operations.

Where necessary sealing of tunnel

walls with the use of shotcrete will

be undertaken; drainage channels

to channel water for storage in

sumps will be pumped to surface for

re-use in mining and processing

operations

Drill Water Management

(aqueous slurry)

Prevent any possible pollution from

drilling mud reporting to the existing

water courses and maintain the

quality any discharges

Treatment of wastewater resulting from

drilling operations, which will aim to

remove at least 80% of the solids in

water and make it pH neutral before

recirculation in the process

Runoff Water Management

(Rainfall)

Establish measures to ensure the

management, treatment, and

delivery of rainfall runoff from the

project infrastructure

Avoid erosion, pollution and any

sediment being deposited to natural

watercourses.

Build open drainage channels in the

design of the tailings storage facility to

intercept and divert runoff flows entering

the stored tailings and waste.

Surface water diversion channels to

prevent water ponding in any areas

A system of water collection with a water

sediment collection sump located below

the collection sump to control runoff

flows and reduce sediment load before

final discharge to the environment

Management of

Water Bodies Crossings

Introduce management measures

for adequate control of drainage at

crossings by existing access roads

and constructing any new accesses

where exploratory prospects may

be located and other infrastructure

required for the project

Maintenance of existing hydraulic

structures such as culverts (single and

double), which permit the passage of the

streams, in order to retain the drainage

pattern of the micro-basins and thus

ensure the functionality of the existing

reticulation system

Adapt bridges across the streambeds

where necessary

Construction of sediment retention

control structures in the form of sumps

and/or dams

Management and Control of

Gas and Particles

To mitigate the increase of gas

concentration on the atmosphere

(CO, O3, NOx y SOx).

Roads and transport

Watering of roads and material

discharge locations

Covering material during transport

Appropriate vehicles for material

transport

Mining Operations

Progressive reclamation of waste

areas to avoid dispersion and

erosion

Exhaust ventilation system in the

underground mining system

Bag filters to be installed to trap

particles from the exhaust gases

Equipment and Machinery

Regular maintenance

Page 20.53





Management of Noise Control Mitigate changes in the decibel

level

Proper preventive maintenance

(greasing of moving parts, regular

cleaning, replacement of damaged

parts, etc.)

Use electrically powered equipment

where possible

Construction of perimeter tree

curtains as noise barriers

Management of Vibration and

Noise from Explosives Use

Prevent potential effects caused by

surface vibration and noise on the

communities of the directly affected

area.

Minimising amount of explosives

per charge

Compensation procedure for any

damage caused

Limit operations whenever possible

to one blast a day, at the end of the

daily shift

Provide hearing protection for

operational staff

Domestic and Industrial Solid

Waste Management

Control the domestic and industrial

waste generation in the project.

Household Waste Separation

Separation procedure at source

contributes to reducing the volume

of solid waste reaching landfills or

disposal sites

Industrial Waste Separation

Correspond to the Hazardous and

Special Waste (hereinafter

HASWAS) categories which

comprise those classified as

hazardous chemicals, oil residues

and fuels

Final disposal

Organic, will be given away to

employees and workers as compost

Ordinary will be managed by the

municipal collection services

companies

Recycled will be marketed in the

town of Santa Rosa de Osos

HASWAS to be delivered to

companies certified by the

Environmental Authority to manage

and dispose HASWAS

Adoption of the ICMC (International

Cyanide Management Code)

Page 20.54





Fuel and Chemicals

Management

Prevent and take remedial action

for any accidental spills of fuels or

chemicals, controlling these on-site

to reduce the risk of affecting the

environment and communities.

Fuels

Will be stored in hermetically sealed

suitably pressure rated containers

Stationary fuel tanks have a

secondary containment system with

a volume equivalent to 110% of the

storage capacity

Tanks will be installed with an

impermeable cover on the base to

prevent leakage to the natural

ground, in compliance with the

relevant regulations

Chemical substances

Will be stored indoors, off ground

minimizing the possibility of contact

with water

Secondary containment systems to

be built into the storage tanks and

cyanide mixing facilities, using

materials that provide an effective

barrier against leakage

Page 20.55




AMENDED NI 43-101 TECHNICAL REPORT Biotic System

Name Objective Activities

Flora and Forestry

Management

Reduce the impact and effects on

the flora species local to the project

activities

Mitigate and/or compensate impacts

on areas where vegetation was

removed during project construction

Minimize the impacts of the activity

of removal of ligneous vegetation

and carry out appropriate forestry

wherever possible in any areas of

disturbance

Flora management

The use of, or felling of vegetation

that lies outside the footprint of the

project infrastructure shall be

prohibited

Progressive rehabilitation and re-

vegetation activities will be

conducted whenever possible prior

to final closure

To identify areas considered of high

environmental relevance and limit

any operations accordingly

Forestry management

Ensure any felling permits required

are issued by the relevant

environmental authority and are

valid

Demarcation will be done by taping

the area and marking all trees to be

felled and trees to be re-located

Take into account the disposal of

wood residues and identify specific

handling and storage areas

Revegetation management

Re-use the topsoil layer that was

stored after removal prior to project

construction

To re-plant with the most

appropriate sowing system

according to soil conditions

An ongoing maintenance program

to be implemented after re-

vegetation.

Wildlife and Flora Conservation

Protect fauna and flora with priority

to endemic species, catalogued as

endangered or forbidden

Reduce the impact and on plant

species critically endangered,

unregistered or unidentified in the

area of the project activities

Having established the presence of

these species, removal, blocking,

transport, replanting

and maintenance of the species to a

previously prepared and properly

selected site will be required to reduce

the chances of loss of these species

Page 20.56





Education and Training on

Ecosystems and Wildlife

Encourage the appreciation of the

natural surroundings and the

technical aspects of the project by

the communities of the area directly

socially affected by the project,

including the personnel hired for the

development.

Facilitate the understanding of the

environment and create a sense of

ownership of the surrounding

environment to the project

personnel

Education and training for the staff

involved in the project with emphasis on

ecosystems and flora and fauna species

of special concern through regular

workshops and discussions regarding:

HSE policies

Environmental regulations and

environmental commitments related

to company activities

Prohibition of capture, handling and

unnecessary disturbances to

wildlife

Friendly work practices with regard

to the environment

Efficient use of natural resources

Compensation for Forested

Areas

Perform compensation for forest

exploitation being conducted on

areas affected by the project.

Assuming that is required to remove

existing ligneous vegetation in areas to

intervene it is required to:

Identify potential sites to be

recovered, specify compensation

with the Environmental Authority

and disclose the information with

the community

Prepare the area to be recovered,

then replant with the recommended

species, implement a maintenance

program for a period of 18 months

and effect a final handover to the

Environmental Authority

It is proposed to plant 3,165 native trees

as compensation for the disturbance by

the project footprint

Compensation for Affected

Flora and Fauna

Perform compensation for the

impacts on the fauna and flora,

which may occur on the areas

affected by the project.

Support for research projects or

repopulation of diverse species that

needs to be carried out in the area

of direct or indirect influence of the

project

Enrichment of habitats for the

continuous growth of wildlife

populations through the planting of

suitable species attractive to wildlife

(key species that provide flowers

and fruits)

Socio-Economic System

The socio-economic environment is very important to Red Eagle Mining and the Project. As a result,

ten management programs were developed to ensure a socio-economic license was introduced and

maintained throughout the life of the project. The socio-economic license has been in the process of

development since late 2011 and Red Eagle Mining continues to work with the local communities and

stakeholders in the region.

Page 20.57




AMENDED NI 43-101 TECHNICAL REPORT A key component of the program covering socio-economic environment is the Community Information

and Participation program. This important aspect is presented below, and shows the level of initial

commitment developed by Red Eagle Mining for the program and its on-going commitment in the

future:

Program Objective Activities

Information and

Community

Involvement

Inform local authorities, communities and

inhabitants of the area of impact, the technical,

social and environmental characteristics of the

project and the management measures being

adopted by the Red Eagle Mining

Communication and engagement

strategies:

Newsletters regarding the project

description, the EMP, actions and

progress.

Information brochures containing

company news announcing progress

of the project and reports

regarding the implementation of the

EMP

Press releases about the project's

progress and implementation of the

EMP, including local and regional

cover.

Information should cover objectives,

steps and actions to develop the

project, with emphasis on those

activities from the technical and

environmental perspective

Community workshops and information

meetings:

Information meetings will be

developed prior to each stage of the

project and presented to local

communities and stakeholders

through the representatives of the

Community Action Boards

Community care office

Community relations office will be

maintained in the town centre of

Santa Rosa de Osos to receive and

answer requests, concerns and

complaints

This office will also provide

information for employment,

contracting for goods and services

and receive resumes and other

proposals from the community

Page 20.58




AMENDED NI 43-101 TECHNICAL REPORT Program Objective Activities

Support for

Institutional

Management and

Economic Activities

To promote closer relations and economic

activities between the communities, Municipal

Administration of Santa Rosa de Osos,

Antioquia and NGOs, with the support of Red

Eagle Mining

Development and project management

workshops:

During the implementation stage

"construction and assembly" of the

project, six workshops will be

conducted with representatives of

the Community Action Boards

Citizen participation workshops

Community organization

Analysis and identification of social

and environmental issues

Transparency:

A transparency commitment by THE

COMPANY will ensure information is

provided to the municipality and the

local community.

An important part of this action is

linking the Santa Rosa Gold Project

to the Transparency Initiative.

Agricultural and livestock fostering:

With the support of UMATA and

programs sponsored by the

Canadian Embassy to Colombia,

develop the technical capacity for

sustainable production as an

economic alternative in the region

(e.g. gardening, raising small

animals, etc.) in the ADI community

during operations

Traditional economic activities:

Consider the promotion of a gold

jewellery training project during the

operational phase

Page 20.59





Employment and

Contracting

Program

Give priority to local communities in the

process of hiring labour, and the supply of

goods and services, wherever possible

Requirements for employment

relationships:

There will largely be performed by

an unskilled work force;

Required personnel;

Legal requirements: age, health

status, Documents, type and

contract period, payment terms

Disciplinary regulations

Disclosure and implementation of a

code of conduct and ethics for

employees, contractors,

subcontractors and suppliers

The hiring of goods and services will be

determined by the following general

guidelines:

Demand for purchase of goods and

services, will be given priority to

Santa Rosa de Osos municipality

The goods and services that do not

exist or do not meet the

requirements of the project, shall be

obtained in regions adjacent to the

municipality

Virtual education program:

This is planned to be a program that

seeks to strengthen the level of the

middle and basic school education

of the ADI population, by providing

internet based training. This program

has the strong support of the

Catholic Diocese of Santa Rosa de

Osos

Page 20.60





Education, Training

and Awareness

Empowering the communities by informing

them about the actions and measures of the

EMP that Red Eagle Mining has implemented

for the project execution. It is intended that the

ADI population adopt these references and

make them part of their everyday life

Strengthen the basic, middle and high school

education level, of the ADI population.

Workshop scope:

Empowering the community with the

plans in compliance with the EMP

and environmental protection

activities

“Adopt a tree” project:

The project aims to engage students

in the planting and on-going tree

care. This allows them to appreciate

their natural resources, and the

benefits it brings to their school by

providing a living barrier

Assisted Virtual Education Program:

This program is aimed at improving

the standard of education in the

communities that are the subject of

ADI, since most of the population

has not finished basic school

education

Participation program guidelines:

The objective is to educate young

people in the ADI about the

participation process in regional

activities and provides the tools so

that the youth engage themselves in

active participation, control and

decision making processes

Page 20.61





Impact and Social

Risk Management

Program

Identify and compensate any potential impacts

caused by Red Eagle Mining or Contractors in

the areas of influence of the project

Minimize social risks that the project can

generate at different stages of activities

Road safety program:

Road safety program will be

implemented that includes a safety

policy, workshops and awareness

campaigns for employees,

contractors and communities near

the roads used by project transport

Signing sensitive and strategic

places within the community and

project areas will be important

Health awareness:

Annual health campaigns will be

conducted in the project ADI to

monitor the epidemiological and

health status of local communities

The campaigns will take place

during one day in each village and

will include the involvement of

entities of the health sector who

provide medical services

National army and police security

agreements:

Agreements are in place and will

continue during both the

construction phase ongoing

operations of the project

The security agreement with the

National Army is part of the

Voluntary Principles on Security and

Human Rights

20.12 Emergency Plans

An emergency plan identifies hazards and describes their prevention measures that may apply to the

Project area of influence. The plan includes procedures to address any emergencies that may occur

during the life of the Project, with the aim of ensuring the protection of human and natural resources,

assets and infrastructure that could be affected.

The main possible risks identified for the project include:

Collapse of the main decline access to the mine;

Underground flooding;

Roof failure of underground workings;

Underground fire;

Page 20.62




AMENDED NI 43-101 TECHNICAL REPORT Process plant fires, leaks or spills of fuels or chemical substances;

Soil and water contamination by spillage of toxic substances; and

Accidents due to improper or inexperienced handling of machinery and mining

equipment.

If any of the risks above occur during the execution of the project, The Company will implement the

following action plan:

Page 20.63




AMENDED NI 43-101 TECHNICAL REPORT Emergency plans are currently posted in all areas of the existing project site. These will be updated

after the commencement of Project construction activities.

Safety programs will incorporate drills to simulate emergency procedures as the construction and

operations evolve.

Regular safety training programmes are an on-going commitment of the Company.

Safety and health of all employees and contractors during project construction and on-going

operations, as well as all members of the local communities and stakeholders shall be of prime

importance. To ensure this is achieved, Red Eagle Mining will provide the necessary human and

financial resources to establish an effective HSES department.

A training centre will be established to assist in the integration of local personnel into the construction

and future operations work force; Red Eagle Mining is committed to maximize local employment

wherever possible.

20.13 Closure Plan

A closure and recovery plan (CRP) has been prepared for the project, which will be refined during the

detailed design and planning phases.

The CRP contains the actions that will be implemented to ensure that socio-environmental

management programs resulting from the cessation of project activities are determined and addressed.

These actions are regularly reviewed and re-assessed as necessary throughout the project operational

lifetime.

The main objective of the plan is to establish a strategy to effectively decommission each part of the

project and to restore those which may have been affected to a state equal to, or better than the

conditions existing before the project commenced.

The CRP is formulated at the initial design concepts of the facilities and is considered through the

engineering, construction, operations, and post-operation phases of the project.

The closure design for the project shall be a systematic implementation through the following steps:

20.13.1 Partial Closure

This will commence early in the construction period and it generally includes the works associated with

the earthworks and civil engineering.

These tasks will be designed and executed to mitigate environmental impacts, by the re-vegetation or

protection of exposed areas of the project site, as soon as practically possible.

Page 20.64




AMENDED NI 43-101 TECHNICAL REPORT 20.13.2 Gradual Closure

This will be an ongoing process over the operational life of the mine. This program consists of the

gradual rehabilitation of the areas that are not in the direct vicinity of mining activities during project

operation.

It aims to ensure geotechnical stability of the site, through progressive backfilling of mining areas,

geomorphological reformation of the terrain and progressive re-vegetation of the tailings and water

storage facilities with native species from the area.

20.13.3 Temporary Closure

Temporary Closure is defined as a period of inactivity of the operation of the mine as a result of force

majeure, such as emergencies, economic, political and/or labour disputes. In the event that any of

these conditions force the temporary closure of the operation, then care and maintenance measures

necessary to protect public health, public safety and the environment during the period of cessation will

be enforced. Red Eagle Mining will ensure the availability of operational staff to implement this plan.

20.13.4 Final Closure and Post-closure

Final closure commences after the mining and processing activities have ceased due to exhaustion of

the ore resources. It will include the dismantling and clearing of the infrastructure and processing

facilities, and the closing and sealing of the underground mine accesses.

A program of slope stabilization, contouring of the surface features, and replacement of stored topsoil

will be developed and undertaken. This will be followed by re-vegetation of the previously disturbed

areas, including final surface and groundwater management stabilization.

The final closure plan also includes social closure, with the aim to provide detailed information to the

public related to the reclamation activities.

The procedure will be planned to ensure compliance with environmental and social commitments in

accordance with the EMP and the approval of the Environmental Authority.

The following outlines the development of Final Closure and Post-Closure Plan. 3-D views of the

resulting closure plans are also shown below:

Physical removal of all equipment and machinery from the project site (this is

typically a cash positive exercise, and the salvage and resale value of a majority of

the equipment will outweigh the cost of dismantling and transportation off site);

The sealing of all mine openings;

The breaking up and removal of all concrete foundations with a rock hammer and

excavator;

Page 20.65




AMENDED NI 43-101 TECHNICAL REPORT The re-contouring of the entire occupied footprint of the mine, process plant and

waste management facility;

The reclaiming of stored topsoil, and spreading over the recontoured old footprint

area;

The revegetation of the entire area with native plants and grasses, along with a large

number of local trees.

It is anticipated that the Final Closure will return the area to a state equal to, or better than the

conditions that existed before the project commenced.

Decommissioning of Mining Areas

Page 20.66




AMENDED NI 43-101 TECHNICAL REPORT Topography of the Land After Closure of the Project

Page 20.67




AMENDED NI 43-101 TECHNICAL REPORT 20.14 Investment Plan

20.14.1 Introduction

According to Colombian legislation, the owners of all projects requiring an Environmental License are

committed to invest 1% of the total budget calculated, prior to operations, of the capital cost of the

installed equipment and machinery of the project. This is in order to contribute to the recovery,

conservation, preservation and monitoring of the basin where the water resources will be drawn from.

20.14.2 Investment Formula

The formula for calculating the commitment towards a local investment has been established by

Decree 1900 of 2006, which characterizes a project with regards to the project Land Management Plan

(LMP) and Municipal Development in the Regional Environmental Management Plan (REMP) and

Water Basins Management Plan (WBMP) as set out by the Environmental Authority.

This budget is based on the costs of earthworks and civil works, and the purchase of machinery and

equipment used in the construction of fixed plant. The budget cost for fixed plant installation was

estimated in the PEA and reported in the PTO to be $US48 million. The actual 1% value was hence

calculated as $US480,000.

20.14.3 Investment Proposals

This 1% allocation must be used for the improvements to the respective water basin where the project

is to extract water from.

Two investment alternatives for the Santa Rosa Gold Project have been proposed:

The installation of domestic sewage treatment systems in the area of influence of the

project to service one of the local communities; and

The reforestation of disturbed areas within the water basin.

The selection of the final alternative will be decided by the Environmental Authority and will be

addressed in the granted Environmental License as a commitment.

20.14.4 Additional Investments

The Company is committed to working closely, and supporting the local communities. As a result, there

are a number of areas where the Company has made commitments and established significant budget

allocations in the course of project studies. Some priority commitments include the following:

A continued environmental monitoring plan which will include representatives of the

local communities to ensure total transparency in the Company’s pursuit to preserve

the environment and mitigate all impacts resulting from the project;

Page 20.68




AMENDED NI 43-101 TECHNICAL REPORT Education and training plans for locals for both employment and furthering education;

Numerous social programs, including community workshops, broadcast of regular

news updates, open forums to discuss issues and question;

Infrastructure projects in the region (e.g. road maintenance, assistance in sewage

systems development, agriculture and small industry workshops).

20.15 Feasibility Study, September, 2014 Update

20.15.1 Technical Differences

There are a number of technical changes incorporated in this Feasibility Study. All of these changes

demonstrate positive impacts on the environment when compared to the PTO and EIA design and

description. It is also important to note that the Feasibility Study only incorporates Measured and

Indicated ore resource classification, not the Inferred classification as reported in the PTO and EIA. As

a result, the tonnages mined and gold produced are less (PTO: 3.6 million tonnes at 4.76g/t gold,

producing 516,000 ounces of recovered gold over 10 years of mine life; Feasibility Study: 2.7 million

tonnes at 4.57g/t gold, producing 388,000 ounces of gold over 8 years of mine life).

Mining

The PTO and EIA described two mining methods – long hole stoping, and mechanized cut-and-fill.

The Feasibility Study describes a variation, which has been named “Mechanized Shrinkage with

Delayed Fill” (MSDF). This method has many advantages over the EIA methods, and allows the

flexibility to revert to cut-and-fill stoping if rock conditions deteriorate.

The advantages are:

Because ore is being slashed from the stope back and dropped on ore, dilution and

ore loss during mucking are greatly reduced;

Because MSDF slashing techniques are used for the bulk of the stope, the powder




If development of MSDF stopes is maintained ahead of ore requirements,

underground stockpiles of ore will be available that can be delivered as required to

the mill;

Most of the fill only requires enough strength for equipment to be operated on top of it

to deliver more fill, which reduces cement requirements and costs; and

Page 20.69




AMENDED NI 43-101 TECHNICAL REPORT The method can easily be converted to cut-and-fill techniques where ground

conditions become weak.

Processing

Comminution

Comminution includes crushing and grinding of the ore.

The PTO and EIA described the use of a primary and a secondary crusher, crushing

in two stages to around minus 75 mm size.

The Feasibility Study describes just one primary crusher which will crush the ore to

around minus 120 mm in size. This has the advantage of crushing to a larger size,

hence reduced equipment and power requirements, with half as much impact.

The PTO and EIA also describe a ball mill to grind down to minus 75 micron in

particle size.

The Feasibility Study describes a semi-autogenous mill (SAG), which uses grinding

balls and the ore as grinding media. The grind size has been determined to be much

higher at 125 micron. This has significant advantages in the dry filtration process, as

the coarser the grind the better and more efficient the filtration process.

Pre-leach Flotation

The PEA assumed that the finer ground ore would be fed directly into the carbon-in-

leach (CIL) circuit.

The Feasibility Study introduces a flotation process after the SAG mill grind to extract

a concentrate containing over 95% of the ore in approximately 10% of feed. This

concentrate is then reground in a vertical mill to reduce the particle size to around 15

to 20 microns.

Leaching

The PEA had direct CIL leaching containing one pre-leach and five leach tanks.

The Feasibility Study has one pre-leach tank and six leach tanks with the flotation

tailings and the reground concentrate being mixed and thickened prior to proceeding

to the CIL circuit.

Elution

The PEA assumed the Zadra elution process.

Page 20.70




AMENDED NI 43-101 TECHNICAL REPORT The Feasibility Study has incorporated the AARL process. The advantage of the

AARL process is the reduced size of the equipment and the ability to run multiple

gold solution strips during a day.

Site Layout

Significant changes have been made to the site layout – all of which significantly minimise the short

and long term impact on the environment.

The PEA design had two separate areas for the process plant and the mine decline

access portal and associated facilities. The Feasibility Study design brings the two

together which allows greater efficiency in ore movement, and reduces the overall

footprint from approximately 60 ha to less than 30 ha. This also has the added

advantages in minimizing impact by reducing noise levels, dust, less visual impacts,

much reduced earthworks and impact on the existing topography.

The PEA design incorporated an “in line” arrangement for the leach tanks, which

was quite inefficient – particularly for the extent of the footprint, and also for operating

and maintenance activities. The Feasibility Study design incorporates a “cluster”

configuration, which reduces the footprint, and is far more efficient.

The PEA design had an “in-line” configuration of crusher feed, crushed ore stockpile

and mill feed, which called for a large footprint.

The Feasibility Study layout is much more compact with the stockpile at 90 degrees.

The tailings will be detoxified and filtered to approximately 16% moisture. 60% of the

filer-cake tailings are planned to be trucked underground and used as stope backfill.

The remaining 40% will be dry stacked on surface in the designed waste

management facility, totalling 1 million tonnes to be deposited on surface over the life

of the mining operations.

Additionally, the mining operations will produce 1 million tonnes of development

waste rock (granodiorite). 40% of this will remain underground as stope backfill; the

remaining 60% (600,000 tonnes) will be blended with the dry stacked process

tailings. Approximately 10% of this product is planned to be crushed and used as

road maintenance sheeting of roads in the local region.

The granodiorite also has a significant advantage in that it is acid consuming.

Testwork has successfully proven that the process tailings are not acid generating,

but the addition of the granodiorite is an added factor of safety.

Page 20.71




AMENDED NI 43-101 TECHNICAL REPORT The process tailings design in the Feasibility Study is also over a much reduced

footprint, by elevating the DWMF above the level of the plant and decline sites. This

reduces run-off from the surface, and enables improved overall management of the

facility.

For comparison the figures below show changes from the PEA to Feasibility Study flow sheet and site

layout.

PEA FLOWSHEET

Page 20.72




AMENDED NI 43-101 TECHNICAL REPORT FS FLOWSHEET

PEA SITE LAYOUT

Page 20.73




AMENDED NI 43-101 TECHNICAL REPORT FS SITE LAYOUT

Page 21.1




AMENDED NI 43-101 TECHNICAL REPORT 21.0 CAPITAL AND OPERATING COSTS

This section outlines the capital and operating costs for the Santa Rosa Gold Project. The capital and

operating cost estimates related to the mine were prepared by MDA. The estimates for the process

plant and the plant site infrastructure were developed by Lycopodium. The costs related to the DWMF

were estimated by Golder. Owner’s costs were provided by Red Eagle Mining. All costs for the Santa

Rosa Gold Project have been expressed in Q3 2014 US Dollars.


contingencies for all the facilities required to bring the Santa Rosa Gold Project into production, as

defined by this feasibility study. All equipment and material are assumed to be new. The labour rate

build up is based on the statutory laws governing benefits to workers in effect in Colombia at the time

of the estimate. Colombian import tariffs have been applied. The estimate does not include any

allowances for escalation, exchange rate fluctuations or project risks. The capital cost estimate has a

predicted accuracy of +/- 15%.

The total estimated cost of the overall project (mine plus process plant) is $ 69.90 million. This amount

excludes market forces and currency hedging. This total has been compiled as shown in Table 21.1.

Table 21.1 Santa Rosa Gold Project Overall Capital Cost Estimate ($M)

Main Area Supply Cost Installation Cost +Freight & Taxes

Total Cost





Mining 9.64 0.04 9.68


Owner’s Costs 8.50 0.00 8.50


Subtotal 62.90

Contingencies 7.00

Grand Total 69.90

21.1 Mine Capital Costs

21.1.1 Mine Capital Cost

The largest portion of the capital cost estimate is attributed to development costs, which have been

based on contractor quotations. Ventilation equipment quotations have been received from vendors.

Other minor equipment capital costs have been assumed based on InfoMine estimation guides. The

capital costs are assumed to be current as of the third quarter 2014 and within an accuracy of ±15%.

Page 21.2




AMENDED NI 43-101 TECHNICAL REPORT Mining capital has been minimized by employing a mining contractor for all mining activity. The mining

contractor will be required to provide the mining equipment, and the cost of the equipment will be

amortized into the mining cost. Mining capital includes development capital, pre-production mining

costs, and other mine capital that is comprised of portal collar work, contractor mobilization, and mine

surface facilities.

Mine development capital costs are shown in Table 21.2.

Page 21.3




AMENDED NI 43-101 TECHNICAL REPORT Table 21.2 Mine Development Capital Cost

Item Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Sublevel Ramps KUSD 797 2,981 4,211 948 1,309 3,218 610 1,350 2,862 18,286

Haulage Drifts KUSD 0 497 254 0 1,159 284 252 180 359 2,986

Main Ramps KUSD 2,539 1,054 601 601 601 603 1,202 731 0 7,933

Ventilation Drifts KUSD 13 87 175 0 159 39 13 71 75 631

Ventilation Raise KUSD 0 363 424 0 313 0 133 122 376 1,732

Total Development Cost KUSD 3,349 4,982 5,666 1,549 3,541 4,144 2,210 2,454 3,673 31,568

Cost of Utilities K USD 149 278 318 155 211 273 125 138 235 1,881

Total Development Capital K USD 3,497 5,261 5,984 1,704 3,752 4,417 2,335 2,592 3,908 33,449

Page 21.4




AMENDED NI 43-101 TECHNICAL REPORT The development capital has been estimated based on the meters of development, and the contractor

proposal for cost of development per meter, for the different development profiles. Contractor

proposed rates were given in Colombian Pesos per meter for development in saprolite, granodiorite,

and transition zones in between. A conversion rate of $1,900 Columbian Pesos per US dollar was

used. Ventilation raise development was quoted to be completed using sub-contracted Alimak raise

equipment rates. The quoted rates are shown in Table 21.3.

Table 21.3 Contractor Quotation for Development ($/m)

Item Contractor Quote

($/m)

Saprolite 3,061

Transition Zone 1,995

Granodiorite 1,578

Ventilation Raises 2,019

The contractor quotations for development costs do not include ground support or utilities, however the

unit costs for installed utilities, bolts, mesh, and shotcrete were provided in the quotation. These were

used to estimate the overall cost for development and added to the capital cost estimate Table 21.2 by

type of development.

In addition to the development mining and support cost, the contractor will supply and install utilities.

The supplies for the utilities will be reimbursed by Red Eagle Mining. The estimated utility cost for

development is also shown in Table 21.2 and includes supplies and hangers for electrical cable,

ventilation tubing, communication lines, water lines, and compressed air lines. Where some of the

utilities can be salvaged from one area to another, they are assumed to be reused. The percentage of

salvaged ventilation tubing reused is between 25% and 30% depending on size. The percentage of

pipe reused is 70%, the percentage of 480 V cable reused is 80%, and the percentage of

communication lines reused is 70%.

Attack ramps would be mined at a 2.5 m wide by 2.5 m high profile, and the unit cost for attack ramps

is $1,426 per metre. Since the attack ramps have a short life related to the stoping areas, this

development is expensed and accounted in the operating cost, with the exception of the pre-production

period.

Mine departmental capital (Owner’s cost) is shown in Table 21.4.

This was estimated as the owner capital cost required to establish mine general, engineering, and

geology services, along with the purchase of specific underground equipment to be purchased by the

owner. Mine general services includes materials for capitalized safety supplies, plotters, printers,

computers, software, and light vehicles. The plotters, printers, computers, and software would be used

by the mine department, including engineering and geology, and the bulk of the software cost is for

mine planning and geology software for design and ore control.

Page 21.5




AMENDED NI 43-101 TECHNICAL REPORT Engineering cost includes surveying systems. In addition, $600,000 was added for production training,

which is the cost of having three expatriate engineers and geologists to establish proper planning,

design, and reconciliation procedures for the mine.

Geology capital is primarily for underground drilling equipment to be used for delineation drilling, to

better define stopes prior to production. Light vehicle capital is also included under each department.

Most of the equipment and supplies that supports the mine will be purchased and installed by the

contractor with reimbursement by Red Eagle Mining. The exception is some equipment that Red

Eagle Mining will purchase from other vendors. This includes primary and auxiliary ventilation fans,

dewatering pumps, compressors, substation, and electrical switching gear. These estimated costs are

also shown in Table 21.4.

The total capital is shown in Table 21.5. This includes items from Table 21.2 and Table 21.4 along

with pre-production capital and pre-production expensed costs. Pre-production capital of $9.44 million

includes cost of ore mined and stockpiled prior to the start of production. Pre-production expensed

includes development cost for mining attack ramps to stoping areas during the pre-production period

along with delineation drilling.

Page 21.6




AMENDED NI 43-101 TECHNICAL REPORT Table 21.4 Owner’s Cost - Mine Departmental Capital


Mine General

Safety Supplies and Inventory K USD 10 10 0 0 0 0 0 0 0 20

Plotters, Printers, and Computers K USD 25 0 0 0 0 0 0 0 0 25

Software K USD 130 50 0 0 0 0 0 0 0 180

Light Vehicles K USD 60 0 15 0 45 0 0 0 0 120

Engineering Capital

Total Station Instrument K USD 80 0 0 0 0 0 0 0 0 80

Production Training K USD 0 600 0 0 0 0 0 0 0 600


Geology Capital

Diamond Core Drill K USD 90 90 0 0 0 0 0 0 0 180


Mine Equipment

Primary Ventilation Fans K USD 918 0 0 0 0 0 0 0 0 918

Ventilation Shaft Collars K USD 166 82 0 0 0 0 0 0 0 249

Auxiliary Ventilation Fans K USD 20 20 40 0 40 40 20 0 0 180

Dewatering Pumps K USD 68 96 28 22 0 0 22 22 0 259

Compressors K USD 35 35 0 0 35 35 0 0 0 139

Substation K USD 0 120 0 0 120 0 0 0 0 240

Electrical Switching Gear K USD 20 25 10 5 15 10 10 5 0 100

Radios K USD 27 11 7 6 6 6 6 7 6 82

Mine Refuge Chambers K USD 0 25 25 0 0 25 0 0 0 74

Total Departmental Capital K USD 1,769 1,209 155 33 351 161 58 34 6 3,776

Page 21.7




AMENDED NI 43-101 TECHNICAL REPORT Table 21.5 Total Mine Capital


Mine Capital Summary

Mine Development K USD 3,497 5,261 5,984 1,704 3,752 4,417 2,335 2,592 3,908 33,449

Initial Delineation Drilling K USD 428 0 0 0 0 0 0 0 0 428

Pre-Production K USD 2,965 0 0 0 0 0 0 0 0 2,965

Contractor Capital K USD 779 0 0 0 0 0 0 0 20 799

Departmental Capital K USD 1,769 1,209 336 214 532 341 239 34 6 4,680

Total Mining Capital K USD 9,439 6,470 6,319 1,917 4,284 4,758 2,574 2,626 3,934 42,322

Total Mining Capital

Initial Capital K USD 9,439 0 0 0 0 0 0 0 0 9,439

Sustaining Capital K USD 0 6,470 6,319 1,917 4,284 4,758 2,574 2,626 3,934 32,882

Total Capital K USD 9,439 6,470 6,319 1,917 4,284 4,758 2,574 2,626 3,934 42,322

Page 21.8




AMENDED NI 43-101 TECHNICAL REPORT 21.2 Mine Operating Cost Estimate

21.2.1 Mine Operating Cost

Mine operating costs have been estimated based on first principle operating parameters and costing

parameters using hourly equipment and personnel rates provided by contractor quotations. Mine

operating costs are based on the MSDF method and are shown in Table 21.6. The cost per tonne is

calculated based on ore tonnes mined per year and are shown in Table 21.7. Note that operating

costs estimated during pre-production have been included into capital costs, but have been estimated

in the same manner as described in this section. Pre-production costs and production tonnage have

been removed from Table 21.6 and Table 21.7.

Electrical and fuel consumption rates were determined using InfoMine estimation guides. Owner

operated equipment cost estimates included ventilation and dewatering costs, which were estimated

using hourly rates from InfoMine. Electrical and fuel costs have been assumed to be $0.11/kWH and

$1.10/L, respectively. This is based on Red Eagle Mining’s input and research.

The overall operating cost is estimated to be $37.36 per tonne.

The accuracy of the mine operating cost estimate is expected to be within ± 15%.

Page 21.9




AMENDED NI 43-101 TECHNICAL REPORT Table 21.6 Mine Operating Cost

Mine Production Costs Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Drilling K USD 960 962 962 962 965 962 962 609 7,346

Blasting K USD 1,615 1,619 1,619 1,619 1,623 1,619 1,619 1,038 12,369


LHD Loading K USD 170 170 170 170 171 170 170 141 1,332


Support K USD 1,196 1,199 1,199 1,199 1,202 1,199 1,199 759 9,153

Backfill K USD 796 914 930 909 898 924 889 685 6,945

Expensed Development K USD 1,636 418 4,703 2,851 1,915 1,257 1,064 2,065 15,908

Mine Support K USD 358 357 357 357 358 357 357 298 2,801

Delineation Sampling K USD 1,483 992 726 727 726 726 726 309 6,414

Ventilation K USD 579 527 533 526 528 525 518 402 4,138

Dewatering K USD 80 86 85 104 104 104 102 59 725

Mine General Services K USD 982 982 982 982 982 982 982 818 7,688

Subtotal Mine Cost K USD 11,318 9,953 14,078 12,082 11,144 10,448 10,331 8,573 87,927

Contractor Fixed Cost K USD 1,931 1,931 1,931 1,931 1,931 1,931 1,931 1,609 15,128

Total Mine Cost K USD 13,249 11,884 16,010 14,013 13,075 12,379 12,263 10,182 103,055

Page 21.10




AMENDED NI 43-101 TECHNICAL REPORT Table 21.7 Mine Operating Cost per Tonne

Mine Production Costs - $/t Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Drilling $/t 2.66 2.67 2.67 2.67 2.67 2.67 2.67 2.63 2.66

Blasting $/t 4.48 4.50 4.50 4.50 4.50 4.50 4.50 4.48 4.48

LHD Haulage to Muck Bays $/t 2.18 2.26 2.32 2.17 2.17 2.07 1.99 1.95 2.14

LHD Loading $/t 0.47 0.47 0.47 0.47 0.47 0.47 0.47 0.61 0.48

Haulage to Surface $/t 1.87 2.54 2.72 2.48 2.47 2.44 2.86 4.05 2.61

Support $/t 3.31 3.33 3.33 3.33 3.33 3.33 3.33 3.27 3.32

Backfill $/t 2.21 2.54 2.58 2.52 2.49 2.57 2.47 2.95 2.52

Expensed Development $/t 4.53 1.16 13.06 7.92 5.30 3.49 2.96 8.91 5.77

Mine Support $/t 0.99 0.99 0.99 0.99 0.99 0.99 0.99 1.28 1.02

Delineation Sampling $/t 4.11 2.76 2.02 2.02 2.01 2.02 2.02 1.33 2.33

Ventilation $/t 1.60 1.46 1.48 1.46 1.46 1.46 1.44 1.73 1.50

Dewatering $/t 0.22 0.24 0.24 0.29 0.29 0.29 0.28 0.26 0.26

Mine General Services $/t 2.72 2.73 2.73 2.73 2.72 2.73 2.73 3.53 2.79

Subtotal Mine Cost $/t 31.35 27.65 39.11 33.56 30.87 29.02 28.70 36.99 31.87

Contractor Overhead / Profit $/t 5.35 5.36 5.36 5.36 5.35 5.36 5.36 6.94 5.48

Total Mine Cost $/t 36.70 33.01 44.47 38.93 36.22 34.39 34.06 43.94 37.36

Page 21.11




AMENDED NI 43-101 TECHNICAL REPORT Drilling and Blasting Costs

The drilling and blasting requirements have been estimated based on the production schedule. The

drilling and blasting cost is a function of the equipment hours, personnel, and consumables required for

production drilling. Table 21.8 shows production drilling costs and Table 21.9 shows the production

blasting cost. Drilling and blasting costs are estimated to be $2.66/tonne and $4.48/tonne of ore

mined, respectively.

Page 21.12




AMENDED NI 43-101 TECHNICAL REPORT Table 21.8 Annual Drilling Costs

Drilling Costs Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Total Production Drill Hours Op Hours 9,732 9,753 9,753 9,753 9,779 9,753 9,753 6,173 74,447

Number of Drills # 2 2 2 2 2 2 2 2 2

Drill Availability % 85% 84% 83% 82% 81% 80% 80% 80% 82%

Schedule Hours per Period AvailOpHours 8,664 8,640 8,640 8,640 8,664 8,640 8,640 7,224 67,752

Available Drill Operating Hours AvailDrillHours 14,714 14,501 14,328 14,155 14,021 13,824 13,824 11,558 110,926

Use of Available Hours % 66% 67% 68% 69% 70% 71% 71% 53% 67%

Power Consumption MWH 506 507 507 507 508 507 507 321 3,869

Power Cost K USD 61 61 61 61 61 61 61 39 466

Fuel Consumption KL 9 9 9 9 9 9 9 6 71

Fuel Cost K USD 9 9 9 9 9 9 9 6 72

Equipment Costs K USD 651 652 652 652 654 652 652 413 4,978

Maintenance Costs K USD 113 113 113 113 114 113 113 72 866

Labour Costs K USD 126 126 126 126 127 126 126 80 965

Total Costs K USD 960 962 962 962 965 962 962 609 7,346

Total Costs $/t 2.67 2.67 2.67 2.67 2.67 2.67 2.67 2.67 2.66

Page 21.13




AMENDED NI 43-101 TECHNICAL REPORT Table 21.9 Annual Blasting Costs

Blasting Costs Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Total Explosives Used Tonnes 257 258 258 258 259 258 258 163 1,969

Explosives Cost K USD 603 604 604 604 606 604 604 382 4,610

Accessory Cost K USD 744 746 746 746 748 746 746 472 5,695

Total Consumables K USD 1,347 1,350 1,350 1,350 1,354 1,350 1,350 854 10,305

Labour Costs K USD 203 204 204 204 204 204 204 129 1,554

Equipment Costs K USD 65 65 65 65 65 65 65 54 511

Total Costs K USD 1,615 1,619 1,619 1,619 1,623 1,619 1,619 1,038 12,369

Total Costs $/t 4.50 4.50 4.50 4.50 4.50 4.50 4.50 4.55 4.48

Page 21.14




AMENDED NI 43-101 TECHNICAL REPORT Loading Cost

Loading costs includes the cost of loaders used to load 20-tonne haul trucks. The haul trucks would

haul ore to the surface and return either empty or with a backhaul of waste/tails for use in backfilling.

Due to the longer cycle times and the number of haul trucks required, the utilization of the loaders is

very low. For this reason, the operator cost has been zeroed based on the assumption that truck

operators would be cross-trained to operate the loaders and load trucks. This reduces the amount of

operators required. The loading cost is shown in Table 21.10 and estimated to be $0.48 per tonne of

ore mined.

Page 21.15




AMENDED NI 43-101 TECHNICAL REPORT Table 21.10 Annual Loading Costs

Production LHD

Loading Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

LHD Loading Hours

Used Prod Hrs

1,437

1,440

1,440

1,440

1,444

1,440

1,440

911

10,992

Operating Hours Op. Hrs

2,211

2,215

2,215

2,215

2,221

2,215

2,215

1,402

16,911

Number of LHD's #

1

1

1

1

1

1

1

1

1

LHD Availability % 85% 84% 83% 82% 81% 80% 80% 80% 82%

Schedule Hours per

Period

Available

Op. Hrs

8,664

8,640

8,640

8,640

8,664

8,640

8,640

7,224

67,752

Available LHD

Operating Hours

Available

LHD Hrs

7,357

7,250

7,164

7,078

7,011

6,912

6,912

5,779

55,463


Fuel Consumption (KL) KL

35

35

35

35

35

35

35

22

267

Fuel Cost K USD 25

25 25 25 25 25 25 21 197



Labour Costs K USD 20 20 20 20 20 20 20 17 159

Total Costs K USD 170 170 170 170 171 170 170 141 1,332

Total Costs $/t 0.47 0.47 0.47

0.47 0.47 0.47 0.47 0.62 0.48

Page 21.16




AMENDED NI 43-101 TECHNICAL REPORT Haulage Cost

Haulage costs have been estimated based on truck requirements to haul ore from underground to the

stockpile outside of the portal using 20-tonne haul trucks. The trucks will also back haul waste/tailings

material into the mine for use in backfilling. The cost of the return trip is included in the haulage cost.

The incremental cost of the truck returning from the backfill dumping location to the production loading

location has been included in the backfill costs.

The haulage cost varies based on the cycle time required. Thus, the cost is time dependent, based on

locations being mined. Life of mine average haulage costs have been estimated to be $4.75 per tonne

mined and are shown in Table 21.11.

Page 21.17




AMENDED NI 43-101 TECHNICAL REPORT Table 21.11 Annual Haulage Costs

Item Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Haulage by Truck

Average Cycle Time Minutes 19.00 25.75 27.51 25.13 24.99 24.70 28.94 41.72 27.60

Truck Hours Used Prod Hours 6,142 8,342 8,914 8,140 8,118 8,001 9,375 8,556 65,587

Operating Hours Op Hrs 8,774 11,917 12,734 11,628 11,597 11,429 13,393 12,223 93,696

Number of Trucks Max # 2 2 2 2 2 2 3 3 3

Truck Availability Avg % 85% 84% 83% 82% 81% 80% 80% 80% 82%

Days per Period Days 366 365 365 365 366 365 365 304 2,861


Available Truck Operating Hours AvailTrkHours 10,424 14,515 14,342 14,170 14,036 13,824 17,299 13,824 112,434


Haulage by LHD

Average Cycle Time Minutes 3.91 4.03 4.14 3.89 3.86 3.69 3.55 3.22 3.83

LHD Loading Hours Used Prod Hrs 8,212 8,474 8,705 8,158 8,152 7,767 7,457 4,275 61,199

Operating Hours Op Hrs 10,265 10,592 10,882 10,197 10,190 9,708 9,321 5,344 76,499

Number of LHD's # 2 2 2 2 2 2 2 2 2

LHD Availability % 85% 84% 83% 82% 81% 80% 80% 80% 82%


Available LHD Operating Hours AvailLhdHours 14,714 14,501 14,328 14,155 14,021 13,824 13,824 11,558 110,926


Total Haulage Consumables and Costs

Fuel Consumption (KL) KL 454 564 596 548 547 534 593 491 4,326

Fuel Cost K USD 357 447 472 434 433 423 473 402 3,441


Maintenance Costs K USD 260 305 319 295 295 286 306 242 2,308

Labour Costs K USD 186 222 233 215 215 208 226 182 1,686

Total Costs K USD 1,461 1,728 1,813 1,675 1,672 1,623 1,744 1,391 13,107

Total Haulage Costs $/t 4.07 4.80 5.04 4.65 4.63 4.51 4.84 6.10 4.75

Breakdown by Truck and LHD



Haulage to Surface $/t 1.87 2.54 2.72 2.48 2.47 2.44 2.86 4.12 2.61

LHD Haulage to Muck Bays $/t 2.19 2.26 2.32 2.17 2.17 2.07 1.99 1.99 2.14

Page 21.18




AMENDED NI 43-101 TECHNICAL REPORT Ground Support Cost

Ground support for production is estimated based on support requirements and the installation of rock

bolts and meshing. As these costs are for support related to stope production, no shotcrete costs have

been included. It is assumed that all of the support would be completed using split set bolts and mesh

only. The estimated cost for ground support is $3.32 per tonne of ore and is shown in Table 21.12.

Page 21.19




AMENDED NI 43-101 TECHNICAL REPORT Table 21.12 Annual Ground Support Costs


Bolts Required # 24,563 24,615 24,615 24,615 24,683 24,615 24,615 15,581 187,903

Mesh Required m^2 41,450 41,538 41,538 41,538 41,652 41,538 41,538 26,293 317,087

Installation - Time Required hours 9,030 9,050 9,050 9,050 9,075 9,050 9,050 5,728 69,082

Bolt costs K USD 626 627 627 627 629 627 627 397 4,786

Mesh Costs K USD 217 217 217 217 218 217 217 138 1,660


Maintenance Cost K USD 6 6 6 6 6 6 6 4 46

Labour K USD 253 253 253 253 254 253 253 160 1,934

Total Support Cost K USD 1,196 1,199 1,199 1,199 1,202 1,199 1,199 759 9,153

Total Costs $/t 3.33 3.33 3.33 3.33 3.33 3.33 3.33 3.33 3.32

Page 21.20




AMENDED NI 43-101 TECHNICAL REPORT Backfill Cost

The backfill costs have been estimated to include return haulage of backfill underground, using

20 tonne haul trucks, and placement of the material using smaller 2.5 tonne loaders. Part of the

backfill costs have been estimated based on equipment requirements for mixing of tailings with 1%

cement, to increase the stability of the tailings, so that equipment can be operated on top of the

backfill. This 1% cemented backfill represents approximately 5% of the total backfill. In some cases,

additional cement (approximately 5%) will be used where material will be mined underneath backfill.

This 5% cemented backfill represents approximately 7% of the total backfill. It is anticipated that the

balance of the backfill will not require additional cement based on the manner of placement and

blending with waste rock.

The estimated backfill cost is shown in Table 21.13 and totals $2.52 per tonne of ore. The backfill

costs take advantage of backhauls by 20-tonne haul trucks, so only the incremental haulage

component is allocated to the backfill.

Page 21.21




AMENDED NI 43-101 TECHNICAL REPORT Table 21.13 Annual Backfill Costs


Backfill Mixing K USD 75 85 85 85 85 90 87 69 661

Haulage Underground K USD 108 138 142 140 138 143 149 135 1,093

Placement K USD 447 505 517 498 487 494 464 330 3,741

Cement Costs K USD 165 186 186 187 187 197 190 151 1,450

Total Backfill Cost K USD 796 914 930 909 898 924 889 685 6,945

Total Backfill Cost $/t 2.22 2.54 2.58 2.52 2.49 2.57 2.47 3.00 2.52

Page 21.22




AMENDED NI 43-101 TECHNICAL REPORT Mine Support Cost

Mine support includes various equipment and operator costs to provide support for mine operations.

Equipment includes: service truck, lube trucks, grader, scissors truck, and a flatbed truck. The

estimated support costs are based on contractor quotations and are shown in Table 21.14. These

support costs are $1.02 per tonne of ore.

Page 21.23




AMENDED NI 43-101 TECHNICAL REPORT Table 21.14 Annual Mine Support Costs


Fuel Consumption (KL) KL 146 145 145 145 146 145 145 121 1,138

Fuel Cost K USD 160 160 160 160 160 160 160 133 1,252



Labour K USD 3 3 3 3 3 3 3 3 25

Total Equipment Cost K USD 340 339 339 339 340 339 339 283 2,660

Supplies K USD 18 18 18 18 18 18 18 15 141

Total Support Costs K USD 358 357 357 357 358 357 357 298 2,801

Total Support Costs $/t 1.00 0.99 0.99 0.99 0.99 0.99 0.99 1.31 1.02

Page 21.24




AMENDED NI 43-101 TECHNICAL REPORT Stope Delineation Cost

Stope delineation will be done prior to mining stopes. These costs are based on owner-operated and

managed diamond drilling and channel samples as shown in Table 21.15.

Drill pattern density commences at 10m x 10m in Year 1, 12.5 m x 12.5 m in Year 2, and 15 m x 15 m

for the remaining life-of-mine. The costs include drilling, labour, maintenance, and assay costs which

are estimated to average $2.33 per tonne of ore over the life-of-mine.

Page 21.25




AMENDED NI 43-101 TECHNICAL REPORT Table 21.15 Annual Stope Delineation Costs

Ore Delineation Costs Units Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Total

Delineation Drill Hole Spacing* m 10x10 12.5x12.5 15x15 15x15 15x15 15x15 15x15 15x15

Delineation Drill Holes Required # 444 284 198 198 198 198 198 76 1,793

Delineation Drill Meters Required m 48,889 31,289 21,728 21,788 21,728 21,728 21,728 8,396 197,275

Delineation Drill Samples # 8,889 5,689 3,951 3,961 3,951 3,951 3,951 1,527 35,868

Number of Channel Samples # 1,337 1,341 1,341 1,341 1,344 1,341 1,341 849 10,235

Drill Operating Hours Op Hrs 17,324 11,087 7,700 7,721 7,700 7,700 7,700 2,975 69,906

Power Consumption MWH 780 499 346 347 346 346 346 134 3,146

Power Cost K USD 101 65 45 45 45 45 45 17 409



Bits and Rods K USD 233 149 104 104 104 104 104 40 940

Drilling Labour K USD 81 81 81 81 81 81 81 47 616

Total Drilling Cost K USD 1,401 926 668 669 668 668 668 274 5,940

Drill Sample Assay Costs K USD 44 28 20 20 20 20 20 8 179

Channel Sampling Labour K USD 13 13 13 13 13 13 13 8 102

Channel Sample Assay Cost K USD 7 7 7 7 7 7 7 4 51

Total Number of Samples # 10,226 7,030 5,292 5,302 5,295 5,292 5,292 2,376 46,103

Total Equipment Cost K USD 1,319 844 586 588 586 586 586 227 5,324

Total Labour Cost K USD 95 95 95 95 95 95 95 56 718

Total Supplies K USD 18 18 18 18 18 18 18 15 141

Total Assay Costs K USD 51 35 26 27 26 26 26 12 231

Total Delineation Program Costs K USD 1,483 992 726 727 726 726 726 309 6,414

Total Delineation Program Costs $/t 4.13 2.76 2.02 2.02 2.01 2.02 2.02 1.36 2.33

*Delineation spacing is the distance between holes horizontally and vertically across the veins of the deposit.

Page 21.26




AMENDED NI 43-101 TECHNICAL REPORT Ventilation Cost

The primary cost of ventilation is for power consumption and system maintenance. The cost is

estimated to be $1.50 per tonne of ore as shown in Table 21.16.

Page 21.27




AMENDED NI 43-101 TECHNICAL REPORT Table 21.16 Annual Ventilation Costs


Ventilation Power Consumption MWh 1,060 1,064 1,067 1,063 1,066 1,063 1,059 870 8,312

Ventilation Power Cost K USD 207 213 217 213 214 212 208 157 1,641

Ventilation Lube Cost K USD 14 11 11 11 11 11 11 9 88

Ventilation Capital Recovery K USD 119 101 102 101 101 101 99 76 801

Ventilation Maintenance K USD 203 166 167 165 166 165 164 130 1,326

Ventilation Fan Costs K USD 543 491 497 490 492 489 482 372 3,856

Consumables K USD 18 18 18 18 18 18 18 15 141

Outside Services K USD 18 18 18 18 18 18 18 15 141

Total Ventilation Cost K USD 579 527 533 526 528 525 518 402 4,138

Total Ventilation Cost $/t 1.60 1.46 1.48 1.46 1.46 1.46 1.44 1.73 1.50

Page 21.28




AMENDED NI 43-101 TECHNICAL REPORT Dewatering Cost

Dewatering costs are based on maintaining up to three primary dewatering pumps for moving water

from lower elevations to the surface. The pumps are staged through the life of mine to reflect the total

depth of the mine. As the head increases due to depth, so does the cost to pump water from the mine.

The pumping rate is designed to operate between 5L/s and 25L/s for the life of the mine. In addition to

the primary pumps, three auxiliary pumps will be used, pumping 25L/s at a 50% utilization. The total

dewatering cost is shown in Table 21.17 and is estimated to be $0.26 per tonne of ore mined.

Page 21.29




AMENDED NI 43-101 TECHNICAL REPORT Table 21.17 Annual Dewatering Costs


Power Consumption MWh 223 247 246 331 331 330 318 153 2,180

Power Cost K USD 29 32 32 43 43 43 41 20 283

Maintenance K USD 21 23 23 31 31 31 30 15 207

Total Pump Cost K USD 50 56 55 74 74 74 72 34 490

Consumable Supplies K USD 18 18 18 18 18 18 18 15 141

Outside Services K USD 12 12 12 12 12 12 12 10 94

Total Dewatering Cost K USD 80 86 85 104 104 104 102 59 725

Total Dewatering Cost $/t 0.22 0.24 0.24 0.29 0.29 0.29 0.28 0.26 0.26

Page 21.30




AMENDED NI 43-101 TECHNICAL REPORT Mine General Services

Mine general services accounts for owner costs for contractor supervision, mine planning and

surveying, mine geology, and supplies. The estimated cost of $2.79 per tonne of ore is shown in

Table 21.18.

Page 21.31




AMENDED NI 43-101 TECHNICAL REPORT Table 21.18 Annual Mine General Services


Mine General Services

Supervision K USD 247 247 247 247 247 247 247 205 1,931

Hourly Personnel K USD 20 20 20 20 20 20 20 17 159

Total Mine General Personnel K USD 267 267 267 267 267 267 267 222 2,090

Engineering

Supervision K USD 189 189 189 189 189 189 189 157 1,477

Salaried Personnel K USD 152 152 152 152 152 152 152 127 1,193


Total Engineering Personnel K USD 381 381 381 381 381 381 381 318 2,987

Mine Geology

Supervision K USD 65 65 65 65 65 65 65 54 511

Salaried Personnel K USD 81 81 81 81 81 81 81 68 636


Total Mine Geology K USD 207 207 207 207 207 207 207 173 1,624

Supplies & Other

Mine General Services Supplies K USD 30 30 30 30 30 30 30 25 235

Mine Light Vehicle K USD 20 20 20 20 20 20 20 17 157

Engineering Supplies K USD 18 18 18 18 18 18 18 15 141

Engineering Light Vehicle K USD 20 20 20 20 20 20 20 17 157

Geology Supplies K USD 18 18 18 18 18 18 18 15 141

Geology Light Vehicle K USD 20 20 20 20 20 20 20 17 157

Total Supplies & Other K USD 126 126 126 126 126 126 126 105 987

Total Mine General Cost K USD 982 982 982 982 982 982 982 818 7,688

Total $/t 2.72 2.73 2.73 2.73 2.72 2.73 2.73 3.53 2.79

Page 21.32




AMENDED NI 43-101 TECHNICAL REPORT Contractor Overhead

Contractor overhead costs provide for administration and facilities that the contractor supplies in order

to manage their operations. This is a fixed monthly cost of $160,933 based on contractor quotation.

Contractor overhead adds $5.48/t to the overall mining cost.

21.3 Process Plant Capital Costs

21.3.1 Introduction


contingencies for all the facilities required for processing, support infrastructure and owner’s costs. The

execution strategy is based on an engineering, procurement, and construction management (EPCM)

implementation approach and horizontal (discipline based) packaging. This capital cost estimate has a

predicted accuracy of +/-15%.

Major plant and site infrastructure included in the capital costs consists of the following:

Process plant;

Electrical switch rooms;

44 kV power line and switch yard;

Ponds;

Ancillary buildings; and

DWMF.

21.3.2 Summary

The total estimated cost of the process plant is $59.25 million as shown in Table 21.19. This total

excludes the initial mining capital investment of $9.44 million (refer to Table 21.5).

Page 21.33




AMENDED NI 43-101 TECHNICAL REPORT Table 21.19 Process Plant Estimated Total Installed Costs ($M)

Main Area Supply Cost Installation Cost + Freight and Taxes

Total Cost






Owner’s Costs 8.50 0.00 8.50


Subtotal 53.22

Contingencies 6.03

Grand Total 59.25

21.3.3 Estimate Support Documents

Table 21.20 defines the level of development of key documents used for the estimate.

Page 21.34




AMENDED NI 43-101 TECHNICAL REPORT Table 21.20 Key Documents Level of Development

Item Feasibility

Site-based Investigations

Geographical location Defined

Topographical survey Completed

Geotechnical survey Completed

Hydrological survey Completed

Power Survey Completed

Project Team site visit Completed

Process Design

Bench-scale testwork Completed

Pilot Plant testwork Not required

PFDs Issued for Basic Design

P&IDs Not prepared

Equipment List Issued for Basic Design

Mass Balance Issued for Basic Design

Equipment datasheets Issued for major equipment

Line List Not prepared

Valve List Not prepared

Facilities Design

Site Plan Issued for Basic Design

Overall Plant Layout Issued for Basic Design

Mechanical GAs Issued for Basic Design

Structural GAs Defined & Modelled

Piping GAs Not prepared

Electrical SLDs Defined

Capital Cost Estimate

Mechanical Equipment (major) Budget Quotes

Mechanical Equipment (balance) Budget Quotes & Lycopodium Database

Concrete Foundations Take-off, budget quotes

Structural Steel Take-off, budget quotes

Piping Factored

Concrete supply & installation rates Budget quotes

Steel supply & installation rates Budget quotes

Earthworks Take-off, budget quotes

Electrical Equipment Budget Quotes

Instrumentation Equipment Budget Quotes

Steel Frame Buildings Take-off

Portable Module Buildings Budget Quotes

Dry Waste Management Facility Take-off

Page 21.35




AMENDED NI 43-101 TECHNICAL REPORT 21.3.4 Procurement

For all major equipment, technically compliant budget quotes obtained from established vendors were

used after proper adjudication. These quotes were benchmarked against pricing obtained for similar

equipment on recent Lycopodium projects. Pricing for minor equipment was obtained from budget

quotes from local vendors and information from the Lycopodium equipment pricing database. For a list

of budget quotes please see Table 21.21.

Table 21.21 List of Equipment and Materials (Budget Price/Estimation)

Item Package# Description Comment

Electrical Equipment

1 5026-7001 Main Transformers Budget Price

2 5026-7002 Medium Voltage Switchgear Budget Price

3 5026-7003 6.3KW Switchgear and MCC Budget Price

4 5026-7004 Variable Speed Drives Budget Price

Mechanical Equipment

5 5026-5001 Primary Crusher Budget Price

6 5026-5001A Regrind Ball Mill Budget Price

7 5026-5002 SAG Mill Budget Price

8 5026-5003 Apron Feeder Budget Price

9 5026-5004 Jaw Crusher Budget Price

10 5026-5005 Screens Budget Price

11 5026-5006 High Rate Thickeners Budget Price

12 5026-5007 Slurry Pumps Budget Price

13 5026-5008 Agitators Budget Price

14 5026-5009 Cyclone Cluster Budget Price

15 5026-5010 Regeneration Kiln Budget Price

16 5026-5011 Pressure Filters Budget Price

17 5026-5012 Electrowinning Cells Budget Price

18 5026-5013 Strip Solution Heater Budget Price

19 5026-5014 Smelting Furnace Budget Price

20 5026-5015 Flotation Cells Budget Price

21 5026-5016 Vertical Mills Budget Price

21.3.5 Earthworks

Quantities for bulk earthworks, access road, ponds, structural pads were derived from the 3D model

incorporating the current plant site layout and the local topography. All excavation is in rippable

material. Borrow material for engineered fill will be from mine waste available during construction.

Page 21.36




AMENDED NI 43-101 TECHNICAL REPORT The earthworks unit rates were based on competitive bids obtained from reputable local contractors.

These rates were benchmarked against Lycopodium in-house data base on actual costs from previous

projects.

21.3.6 Concrete

The vast majority of the concrete foundations, piers, equipment rafts, slabs on grade were modelled

based on preliminary sizing and available in-house data from similar projects. All foundation sizing is

based on an assumed native soil bearing capacity of 300 kPa with maximum allowable settlement of

20 mm. The concrete quantities were derived from a take-off from the 3D model plus allowances for

the portions that were not fully modelled. The final quantities were benchmarked against actual take-

offs from previous Lycopodium projects.

The unit rates are based on budget quotes from established local contractors and include the costs of

supplying the concrete and the reinforcing steel and placing and curing the concrete. The quotes were

based on detailed material take-offs identifying the concrete quantities broken down by type: slabs,

walls, piers, rafts, suspended slabs, etc. The contractors were also sent typical concrete drawings and

details together with specifications and snapshots from the 3D model showing all the concrete

structures. The unit rates obtained from the local contractors were assessed and compared with the

Lycopodium in-house rates database for similar projects.

21.3.7 Steelwork

All the major structures in the process plant were modelled based on preliminary sizing and

benchmarked against previous Lycopodium projects. This includes major load carrying members,

secondary steel, primary bracing, stairs, and grating. Tertiary steel, horizontal bracing, girts, connection

plates were not modelled and were estimated through allowances. The final material take-off plus

allowances for the members that were not modelled was compared against actual quantities from

recent projects.

Competitive bids were obtained from local contractors experienced in this type of work. These bids

were based on detailed material take-offs identifying the steel by type: light, medium, and heavy.

Typical steel drawings, specs, and snapshots from the 3D model were also sent to the contractors so

that they can visualize the types of structures part of this scope. The rates obtained from the

contractors were evaluated against similar projects and selection made based on costs and capability.

21.3.8 Plate work and Tankage

This scope includes the supply and installation of steel tanks, bins, hoppers, and chutes. The platework

and tankage quantities were estimated using vessel sizing provided in the mechanical equipment list. A

preliminary design was undertaken for each tank in order to select appropriate plate thicknesses to

develop tank tonnages. The platework quantities for major bins and hoppers were developed from

actual take-offs from projects completed by Lycopodium in the last few years. Lining materials were

quantified separately and benchmarked against existing in-house data.

Page 21.37




AMENDED NI 43-101 TECHNICAL REPORT The platework and tankage unit rates are based on budget quotations received from local contractors

with proven capabilities in similar projects. Special allowances were made to take into account the

complexity of some of the bins and chutes, which historically are underestimated by contractors. The

final unit rates adopted in the estimate are a combination between the quotes received from the local

contractors and benchmarked data from previous Lycopodium projects.

21.3.9 Mechanical Equipment

The quantities and size of the mechanical equipment were taken from the detailed mechanical

equipment list developed for this feasibility study. Budget quotations for major equipment were

obtained from local and international vendors; see Table 21.21. Technical and commercial evaluation

and final selection was made by the engineering and procurement personnel working on the project.

Costs for minor items were developed through a combination of budget quotes and information from

the Lycopodium’s in-house database.

The equipment installation hours were developed from in-house data adjusted for local conditions and

labour productivity together with information supplied by the equipment vendors. Final installation costs

include provisions for the retrieval of the equipment from storage location, handling, placing,

installation, and commissioning of the equipment.

21.3.10 Plant Pipework and Valves

Piping general arrangements were not produced for this feasibility study and the pipework was not

modelled. The supply and installation costs for in-plant piping and valves were factored as a

percentage from the mechanical equipment supply and installation costs. These factors were

developed by plant area (crushing, milling, flotation, CIL, etc.) and benchmarked against previous

projects executed by Lycopodium.

21.3.11 Overland Piping

Overland piping consists of raw water supply lines, La Veta creek diversion pipelines, and return

pipeline from the sedimentation pond to the process water tank. Detailed take-offs based on the overall

site plan were developed for the overland piping scope. Budget quotations for the supply and

installation costs were received from local contractors experienced in HDPE pipeline installation. The

final costs adopted in the estimate were benchmarked against Lycopodium in-house data.

21.3.12 Electrical and Instrumentation

Budget quotes were obtained for major electrical and instrumentation equipment from international and

local vendors. Quantities for cable, cable trays, and other bulk items were developed from the general

arrangement drawings produced for this study. Detailed material take-offs and specifications were sent

to local suppliers for pricing. The supply and installation costs for the equipment and bulks were

evaluated against the Lycopodium in-house database. Allowances were made for retrieval from

storage, handling, placing, installation, and commissioning of the equipment.

Page 21.38




AMENDED NI 43-101 TECHNICAL REPORT 21.3.13 Buildings

A building list showing the size and type of construction of all process and ancillary buildings was

developed for this feasibility study. The major steel frame buildings (plant workshop and main

warehouse, tails filtration building, gold room, reagents storage) were modelled and estimated through

material take-offs from the model. These take-offs included steelwork, concrete foundations and slabs

on grade, cladding, roofing, etc. The supply and installation costs were developed based on the unit

rates adopted for the individual commodities (steel, concrete, cladding, etc.). The overall building costs

were benchmarked against industry standard rates and data from the Lycopodium in-house database.

Budget quotes were obtained from local suppliers for all the portable module buildings (plant

administration building, laboratory, electrical switch rooms, control rooms). These quotes were

evaluated against data available from previous Lycopodium projects.

The mine workshop and warehouse and mine truckshop will be provided by the mining contractor and

are included in the mining capital estimate.

21.3.14 Labour Rates and Crew Rates

Base labour rates for different trades and classifications have been obtained from Colombian

contractors. Payroll mark-ups or burdens for social charges and uplifts have been determined from first

principles taking into consideration site conditions, work exposure, and existing legislation.

Average crew direct labour rates were developed for each discipline considering a craft mix of

foremen, skilled tradesmen, apprentice helpers, and unskilled workers. A crew may have 10 to 30

crafts people working together to install a particular commodity (steelwork, concrete, pipework, etc.).

The individual craft wage rates for each worker were then averaged to show a combined overall crew

rate that includes base pay, employer fringe benefits and government burdens per hour of direct

labour.

Equipment rental rates and contractors’ indirects were added to the direct labour rates to derive the

“All-in” crew rates per discipline. Table 21.22 provides a summary of the average direct, indirect, and

“All-in” rates per discipline.

Page 21.39




AMENDED NI 43-101 TECHNICAL REPORT Table 21.22 Crew Rates

Crew

Base + fringes rate

$ per hour

Eqpt rental rate

$ per hour

Indirects rate

$ per hour

"All-in" rate $

per hour

Earthworks 10.94 16.19 8.16 35.29

Concrete 8.87 9.85 6.50 25.22

Steelwork 10.66 11.83 8.49 30.98

Platework & Tankage 11.29 12.53 7.51 31.33

Mechanical Equipment Installation 11.49 12.75 9.36 33.60

Pipework 10.32 11.46 8.05 29.83

Electrical 11.56 12.83 7.28 31.67

Architectural - Buildings 10.32 11.46 6.77 28.55

21.3.15 Contractors’ Indirects

These Contractors’ Indirects are not included in the “All-in” crew rates and cover the costs for

mobilisation and demobilisation of labour and equipment to and from the projects site. Other items

included in these costs are the temporary site facilities and utilities for each contractor, maintenance of

temporary facilities and equipment, construction supervision support, project expenses (miscellaneous

minor licenses and permits), and contractors’ fees and overhead. The contractors’ indirects were

developed for each discipline taking into account the complexity of the scope of the work, equipment

requirements, and availability of local qualified labour. Additional allowance per discipline was included

for expatriate workforce to be present on site to ensure quality, productivity, and completion schedules

are achieved.

21.3.16 Productivity

Productivity factors were applied to direct field labour hours in order to compensate for lower labour

productivity on the job site. The following factors can contribute to poor field productivity: difficult

access to the project site, weather conditions, inexperienced work force, lack of experienced

supervision personnel. All these factors are necessary to adequately transform construction drawing

details and specifications into effective construction shift objectives, efficiently deliver material and

equipment to work areas, and effectively utilize and maintain construction equipment and power tools

on the job site. The productivity factors for this estimate were based on information obtained from local

contractors and Lycopodium’s experience in executing projects of this type and size. Table 21.23

summarizes the productivity factors adopted for this capital cost estimate. The productivity values

shown in the table are compared to the standard US Gulf Coast productivity rates, and are considered

to be conservative for the Project.

Page 21.40




AMENDED NI 43-101 TECHNICAL REPORT Table 21.23 Productivity Factor (PF)

Discipline PF

Earthworks 2.50

Concrete 2.50

Steelwork 2.00

Platework & Tankage 2.00

Mechanical 2.00

Pipework 2.00

Electrical & Instrumentation 2.50

Architectural (Buildings) 1.41

21.3.17 EPCM Services

A detailed deliverables based estimate was developed for the EPCM services required for the project

development. Engineering and drafting hours were estimated for each engineering discipline and

deliverable type. The construction management and commissioning services were estimated on a time

basis as per the execution schedule prepared for the project. The execution schedule is presented in

Section 24. Allowances for construction offices, catering, and accommodation for EPCM site based

personnel were also included in the capital costs.

21.3.18 Working Capital

Working capital was calculated based on the projected plant operating costs for the first two months of

operations.

21.3.19 Vendor Commissioning

Vendor commissioning costs were estimated for each equipment package requiring vendor

representation during construction and commissioning. The site hours needed for each package were

derived based on the complexity of equipment, information received from the respective vendors, and

in-house database from previous Lycopodium projects. An allowance for other vendor representative

expenses (catering, accommodation) were also included in the capital costs.

21.3.20 Spares

The capital cost estimate includes an allowance for commissioning and insurance spares. The

insurance spares category includes first fill warehouse spares. A minimalist approach has been

assumed with spares stocks progressively expanded during operations.

Page 21.41




AMENDED NI 43-101 TECHNICAL REPORT 21.3.21 First Fills Inventory

Quantities for first fills inventory have been calculated based on a monthly delivery schedule. This

inventory consists of critical consumable items purchased and stored on site at the onset of operations.

First fill items include sodium cyanide, hydrated lime, activated carbon, hydrochloric acid, sodium

metabisulphite, and an initial charge of grinding media. This inventory ensures adequate consumables

are available for the first stage of operation. Budget quotations were obtained for these reagents and

consumables and the costs incorporated into the estimate.

21.3.22 Exchange Rates

Quotes for equipment and construction bulks obtained in foreign currencies are expressed in the

estimate in US Dollars based on the foreign exchange rates shown in Table 21.24.

Table 21.24 Exchange Rates

United States Dollar Other Currency

1.0 USD = 1.0849 CAD (Canadian Dollars)

1.0 USD = 2.2420 BRL (Brazilian Real)

1.0 USD = 2.8455 PEN (Peruvian Nuevo Sol)

1.0 USD = 1900.0 COP (Columbian Peso)

1.0 USD = 0.5889 GBP (British Pound)

1.0 USD = 0.7369 EUR (Euro)

1.0 USD = 10.741 ZAR (South Africa RAND)

1.0 USD = 1.0641 AUD (Australian Dollar)

21.3.23 Freight

Freight costs for majority of the equipment were estimated as 12.5% of the supply costs of the

respective equipment. For mobile equipment and some construction bulks (civil) the freight costs are

included in the supply costs. Freight costs for most of the remaining bulks were estimated as 1% of the

supply costs. Freight quotes were obtained for some major equipment items (SAG mill, filter presses).

These quotes were evaluated against the available in-house data and adopted or adjusted as

applicable.

21.3.24 Owners’ Costs

The Owners’ cost budget was developed by Red Eagle Mining. These costs include:

The Owners’ team;

Recruiting and training costs;

Page 21.42




AMENDED NI 43-101 TECHNICAL REPORT The capital costs associated with the expansion of the existing exploration camp;

The costs associated with the upgrade of the access road to the project site including

any land purchase costs;

Mobile crushing / screening plant;

Communications;

Hardware and software costs;

Construction all risks insurance;

Government affairs and public relations;

Environmental insurance; and

Environmental permit and obligations during the construction year.

21.3.25 Contingency

The purpose of contingency is to make specific provision for uncertain cost items within the project

scope. Contingency does not cover scope changes, escalation, or exchange rate fluctuations. The

unforeseeable items covered by contingency are often referred to “unknown unknowns” within the

scope of the project. These can arise due to:

Labour productivity variations due to contractors not providing or not having access

to labour with the required level of skills as assumed in the various direct cost

estimates;

Labour rates or construction equipment rental rates being different from the base

assumptions adopted in the capital cost estimate; and

Equipment and bulk material cost variations from the budgetary pricing submitted for

the capital cost estimate.

Contingency is an integral part of an estimate and has been applied to all parts of the estimate, direct

costs, indirect costs, owners’ costs, etc. The contingency has been derived jointly by Red Eagle Mining

and Lycopodium on a discipline basis, taking into consideration scope definition, material supply costs,

and installation costs. The average contingency for the process plant capital costs is 11.2%.

21.3.26 Qualifications for Capital Cost Estimate

The following items have been excluded from the capital cost estimate:

Page 21.43




AMENDED NI 43-101 TECHNICAL REPORT Minor additional geotechnical investigations;

Minor additional metallurgical testwork;

Finance and interest charges during construction;

Escalation costs;

Currency exchange fluctuations;

Future exploration costs;

Owners’ management reserve;

Bulk fuel storage, as it is assumed it will be vendor supplied under a long term fuel

supply agreement;

44 kV Transmission line, to be built by EPM and amortized (covered by operating

costs);

Any environmental requirement not identified in this estimate;

Changes in Colombian tax law;

Changes in market conditions impacting equipment and bulk costs; and

Any provision for force majeure events.

The following items have been included in the financial analysis:

Sunk costs;

Value added tax and duties on imported goods; and

Reclamation costs and plant closure.

21.4 Process Plant Operating Cost Estimate

21.4.1 Introduction

The operating costs have been developed according to typical industry standards and norms

applicable to a gold processing plant producing gold doré, including all associated general and

administrative costs.

Quantities and cost data were compiled from a variety of sources including:

Page 21.44




AMENDED NI 43-101 TECHNICAL REPORT Metallurgical testwork;

Supplier quotations;

Advice from Red Eagle Mining;

Lycopodium data; and

First principle estimates.

21.4.2 Plant Design Parameters

Operating costs have been developed according to the plant design criteria: nominal ROM throughput

of 360,000 tonnes per annum; average gold feed grade of 4.57 g/t over the life-of-mine; overall gold

recovery of 95.7% and average metal production of 50,620 oz of gold per annum. This forms the basis

for the annual operating costs, which vary by year in the cash-flow model (Section 22.0) according to

the LOM process plant feed schedule and recovery described in Section 22.2.1.

21.4.3 Qualifications and Exclusions

The operating cost estimates include all direct costs associated with the Project to allow production of

gold doré. The cost estimate is presented with the following exclusions:

All mining and exploration costs, except for laboratory assays;

All taxes (included in the financial model);

Refining costs for transportation, marketing and insurance (included in the financial

model);

License fees or royalties (included in the financial model); and

Contingency allowance.

21.4.4 Operating Cost Accuracy

The expected order of accuracy for the operating cost analysis is in the range of ±15%, as required by

the estimate class, and it is deemed appropriate for the study.

21.4.5 Cost Categories

The operating cost estimate includes five major categories as defined below:

Process plant labour;

Page 21.45




AMENDED NI 43-101 TECHNICAL REPORT Operating consumables;

Power;

Maintenance; and

General and Administration.

A description of each cost category is provided in the following sections.

21.4.6 Process Plant Labour

The process plant labour is divided into the following areas: management, operations, metallurgy,

laboratory and maintenance. The process plant labour includes a combination of day and shift work.

The estimated annual process plant labour cost is $2.12 million, or $5.90/t of ore.

Wages and Salaries

Wages and salaries have been provided by Red Eagle Mining. A summary of the salary for each

position is provided in Table 21.25. These salaries were provided inclusive of overheads costs. Staff

is divided into three classes: Management (M), Salary Hourly (H) and Salary Shift (S).

Page 21.46




AMENDED NI 43-101 TECHNICAL REPORT Table 21.25 Wages and Salaries

Position Empl.

Number

of

Teams

Total

Number

Empl.

Class

Base

Annual

Salary

$

Overhead

Costs

%

Annual

Labour

Costs

$/y

Total

Annual

Labour

$/y

Total

Annual

Labour

$/t ore

Process Plant Manager 1 1 1 M 250,000 25% 312,500 312,500 0.87

Operations

Shift Supervisor 1 4 4 S 35,000 45% 50,750 203,000 0.56

Crushing Operator 1 4 4 S 18,947 45% 27,474 109,895 0.31

Milling / Flotation Operator 1 4 4 S 18,947 45% 27,474 109,895 0.31

CIL Operator 1 4 4 S 18,947 45% 27,474 109,895 0.31

Tailings Operator 1 4 4 S 18,947 45% 27,474 109,895 0.31

Goldroom Operator 2 1 2 H 18,947 45% 27,474 54,947 0.15

Loader Operator 2 4 8 S 13,241 45% 19,200 153,600 0.43

Shift General Labour 2 4 8 S 8,206 45% 11,899 95,192 0.26

Metallurgy

Senior Metallurgist 1 1 1 H 45,000 45% 65,250 65,250 0.18

Plant Metallurgist 1 1 1 H 28,000 45% 40,600 40,600 0.11

Laboratory

Lab Supervisor 1 1 1 H 21,750 45% 31,538 31,538 0.09

Assay Technician 3 1 3 H 13,241 45% 19,200 57,600 0.16

Sample Prep. Technician 2 1 2 H 8,206 45% 11,899 23,798 0.07

Maintenance

Maint. Superintendent 1 1 1 M 170,000 25% 212,500 212,500 0.59

Maint. Planner/Trainer 1 1 1 H 35,000 45% 50,750 50,750 0.14

Mechanical Supervisor 1 1 1 H 35,000 45% 50,750 50,750 0.14

Boilermakers 1.5 2 3 S 19,274 45% 27,945 83,840 0.23

Millwright 1.5 2 3 S 19,274 45% 27,947 83,842 0.23

Electricians 1.5 2 3 S 19,274 45% 27,947 83,842 0.23

Instrument Technicians 1 2 2 S 19,274 45% 27,947 55,895 0.16

Trades Assistants 1 2 2 S 8,206 45% 11,899 23,798 0.07

Process Total - - 63 - - - - 2,122,822 5.91

ROM Throughput = 360,000 t/y

Page 21.47




AMENDED NI 43-101 TECHNICAL REPORT Process Plant Operations

The daily operation of the process plant will be under the control of the process plant manager, who

will also act as the process trainer. There will be a total of four operations shift crews staffed by local

labour, to cover back-to-back 12 hour shifts.

Each shift crew will include:

One shift supervisor who will direct the day to day plant operation;

One crusher operator, who will also oversee the surge bin reclaim area;

One milling operator, who will also oversee the flotation circuit and the pre-leach

thickener;

One leach operator, who will be responsible for maintaining the CIL, carbon

regeneration kiln and reagent preparation;

One tailings operator, who will be responsible for the cyanide detoxification, filter

press and clarifier area;

Two loader operators to operate mobile equipment; and

Two general shift labourers, who will provide relief for daily activities.

In addition, the day shift will also include one gold room supervisor and one gold room operator.

Metallurgy

The daily metallurgical performance of the plant will be monitored by both a senior metallurgist and a

plant metallurgist, who will also have responsibility for metallurgical accounting. The metallurgist will

work closely with geologists and mining engineers to ensure that the plant operates at maximum

productivity.

Laboratory

Daily analytical plant support services are provided by the laboratory staff. Two sample preparation

technicians will cover all tasks in the sample preparation area such as reception, crushing, splitting and

pulverizing. Three assay technicians will cover all fusion, cupellation tasks, and atomic absorption

analysis. One laboratory supervisor will oversee daily operations and assure quality and efficiency.

Maintenance

The maintenance superintendent will control all aspects of plant, building and services maintenance.

Page 21.48




AMENDED NI 43-101 TECHNICAL REPORT The maintenance planner will be recruited early in the project to ensure early capture of all critical

equipment data and preventative maintenance requirements to Red Eagle Mining’s maintenance

planning system. The person occupying this role will also assume the duties of the maintenance

trainer.

There will be a total of two maintenance team shift crews staffed by local labour, to cover day shifts. A

total of three boilermakers, three millwrights and three electricians will be divided between two shift

rotations. In addition, each maintenance team will also include an instrumentation technician and a

trade assistant. The maintenance team will be supplemented by appropriately skilled contract labour

to undertake major tasks such as relining the crusher.

21.4.7 Consumables

The consumables category covers all the wear parts and consumable material in the process plant.

The consumables include liners for equipment such as crushers and mills, grinding media, screen

decks, and other relevant items, chemical reagents as well as fuel (diesel and natural gas).

The estimated annual operating consumables cost is $2.96 million, or $8.22/t of ore.

Consumption rates and pricing are summarized in

Table 21.26 and have been based on the following:

Crusher liner, SAG mill liner, Vertimill liner as well as steel ball consumption rates

have been based on Lycopodium calculations. The grinding steel ball consumption

rate has been calculated by using the Abrasion Index (Ai) of 0.139. Costs have been

based on vendor quotations;

Laboratory testwork results have been used, wherever possible for the establishment

of reagent consumption rates. In the absence of testwork data, reagent consumption

rates were assumed based on first principle calculations, Lycopodium experience

and generally accepted practice within the industry. To date, reagent consumption

rates have not been optimized;

The consumption rates for flotation reagents, namely PAX and MIBC have been

based on testwork data and the prices are based on vendor quotations. Copper

sulphate addition in the cyanide destruction area is based on first principle

calculations and Lycopodium experience;

Flocculent consumption is based on Lycopodium experience and pricing has been

supplied by vendors;

Page 21.49




AMENDED NI 43-101 TECHNICAL REPORT The consumption rates for hydrated lime in the CIL circuit are based on testwork.

The hydrated lime consumption rate in the cyanide destruction area is based on first

principle calculations and Lycopodium experience. Costs have been based on

vendor quotations;

The consumption rates for sodium cyanide in the CIL circuit are based on testwork.

Costs have been based on vendor quotations;

Sodium metabisulphite (SMBS) consumption is based on first principle calculation

and Lycopodium experience and pricing has been supplied by vendors;

Activated carbon consumption has been based on Lycopodium experience and the

price has been supplied by a vendor;

Elution and gold room reagent consumption rates have been based on first principles

calculation and Lycopodium experience and the price has been supplied by vendors;

Diesel fuel and natural gas consumption rates have been based on first principles

calculations and Lycopodium experience and the prices have been supplied by

vendors;

Antiscalant consumption rates have been based on Lycopodium experience and the

price has been based on vendor quotation;

Water treatment plant consumables have been based Lycopodium experience; and

Laboratory costs have been allocated on a per sample basis. These costs are

included in the General and Administration cost category.

Page 21.50





Table 21.26 Summary of Major Consumables

Item Supplier Price

US$/Unit Unit

Annual

Consumption

Total Cost

US$/y

Total Cost

US$/t ore

Jaw Crusher

Upper Cheek Plate Metso 1,834 set 1.49 set/y 2,731 0.01

Lower Cheek Plate Metso 1,132 set 1.49 set/y 1,686 0.00

Fixed Jaw Metso 6,093 set 0.74 set/y 4,537 0.01

Swing Jaw Metso 4,992 set 0.74 set/y 3,717 0.01

Mill Liners

SAG Mill Liner Outotec 295,000 set 0.70 set/y 206,795 0.57

Vertimill Liner Metso 35,128 set 1.00 set/y 35,128 0.10

Steel Balls

SAG Mill Steel Balls (105 mm) Sino 1,050 tonne 217 t/y 227,931 0.63

Vertimill Steel Balls (25 mm) Sino 999 tonne 56.3 t/y 56,224 0.16

Reagents

PAX Flottec 2,900 tonne 18.0 t/y 52,200 0.15

MIBC Flottec 3,500 tonne 16.2 t/y 56,700 0.16

Flocculent Charles Tennant 3,388 tonne 6.20 t/y 21,006 0.06

Hydrated Lime DOI 305 tonne 752 t/y 229,680 0.64

Sodium Cyanide Dupont 2,750 tonne 194 t/y 534,006 1.48

Activated Carbon Chemiqa 2,900 tonne 10.8 t/y 31,320 0.09

Sodium Hydroxide Charles Tenant 625 tonne 17.8 t/y 11,111 0.03

Hydrochloric Acid Surtiminas 763 tonne 90.5 t/y 69,051 0.19

Sodium Metabisulphite Charles Tennant 479 tonne 454 t/y 217,048 0.60

Copper Sulphate Charles Tennant 3,014 tonne 37.8 t/y 113,929 0.32

Antiscalant - 2,000 tonne 1.39 t/y 2,777 0.01

Fuel

Diesel Fuel Mining Contractor 1,110 /kL 747 kL/y 829,473 2.30

Natural Gas Indisa S.A. 0.68 /m3 174,202 m3/y 119,191 0.33

Other Consumables - - - - - 131,983 0.37

Total - - - - - 2,958,244 8.22

Page 21.51




AMENDED NI 43-101 TECHNICAL REPORT 21.4.8 Power

The plant power consumption has been determined from the installed power in the electrical load list,

excluding the standby equipment. Electrical load factors and utilisation factors were applied to the

installed power to arrive at the annual average power draw, which was then multiplied by total hours

operated per annum and the electricity price to obtain the plant power cost.

Power will be provided from an existing substation owned by Empresas Publicas de Medellin (EPM). The

power unit cost of $ 0.11 / kWh has been provided by Indisa, in addition to a $1,050,283 fixed OHL power

cost, which is amortized over the 8-year LOM at 6% interest and a monthly fee of $13,802. This equates

to an estimated annual power cost of $2.66 million, or $7.37/t of ore.

Table 21.27 Summary of Process Plant Power Costs

Description

Installed

Power

kW

Peak

Demand

kW

Average

Power

kW

Annual

Consumption

kW

Total

Cost $/y

Total

Cost $/t

ore

Process Plant Equipment 4,930 3,948 2,583 22,630,584 2,489,364 6.91

44 kV Line – OHL Fixed Power Cost - - - - 165,624 0.46

Total - - - - 2,654,988 7.37

*OHL fixed power cost amortized over 8 years

21.4.9 Maintenance

Maintenance materials’ costs have been estimated by applying factors to the ex-works mechanical

equipment cost in each area of the plant. This is done to cover the cost of all maintenance materials and

contract labour requirements, with the exception of crusher and mill wear parts, which have been included

in the consumables allowance. The factors applied are based on Lycopodium’s database and experience,

and are average costs over the life of the mine. As such, actual spares costs may be lower during the

initial years but rise later. An overall factor was calculated to be 3.5% of the mechanical supply cost ex-

works. The estimated annual maintenance cost for process plant and mobile equipment is $944,000, or

$2.62/t of ore. Maintenance costs are summarized in Table 21.28.

Mobile Equipment

The operating costs for mobile equipment have been estimated and include diesel fuel, tires and

maintenance parts. Operating allowances for mobile aggregate plant and road gravel distribution is

inclusive in the mobile equipment. The fuel costs have been included in the consumables cost category

whilst the other operating costs have been included in the overall maintenance materials cost category.

Page 21.52




AMENDED NI 43-101 TECHNICAL REPORT Table 21.28 Summary of Maintenance Costs

Area

Mechanical Supply Ex-

Works

US$

Maintenance

Factor

%

Total Maintenance

Cost

$/y

Total Maintenance

Cost

$/t ore

Plant Equipment Subtotal 11,426,360 3.5% 396,458 1.10

Mobile Equipment - - 517,570 1.44

General Maintenance

Allowance - - 30,000 0.08

Total - - 944,027 2.62

21.4.10 General and Administration Costs

This category covers the General and Administration (G&A) costs required for running the operation,

which have been supplied by Red Eagle Mining.

The estimated annual General and Administration cost is $3.49 million, or $9.68/t of ore. Costs

summarized in Table 21.29 have been based on the following:

Red Eagle Mining provided costs for the following: administrative labour; administrative

consumables; camp catering; legal fees; insurances; land use rent; training sessions;

first aid consumables; travel and personnel transport; community support; environmental

testwork and monitoring; laboratory consumables; import duties, customs fees and land

transport fees; as well as other administrative costs.

Salaries and overheads have been applied to the following administration areas:

administration; safety and security; transportation; community; and environmental. This

cost category includes mostly day work for the administration staff; with the exception of

security staff who perform shift work.

Laboratory staffing is included in the process plant labour cost category. External assay

unit costs for sample analysis of solids ($19.88/unit), solutions ($14.72/unit), carbon

($29.27/unit) and bullion ($69.30/unit) have been supplied by ALS Global regarding

sample preparation, fire assays and atomic absorption analytical testing. Internal assay

unit costs for sample analysis of solids solutions are taken at 40% of the external assay

rate; this is to cover material costs for assaying consumables.

Page 21.53




AMENDED NI 43-101 TECHNICAL REPORT Table 21.29 Summary of G&A Costs

Item Annual Costs

$/year $/t ore

Administration

Personnel Salaries plus Overhead 925,100 2.57

Camp Catering (30 personsX$16X360) 172,800 0.48

Electrical Supply (100KwX$0.13X8000hrs) 104,000 0.29

Telecommunications/IT 50,000 0.14

Site Office Supplies, Freight 15,000 0.04

Business Travel 50,000 0.14

Recruitment 15,000 0.04

Conferences 15,000 0.04

Audit 25,000 0.07

Bank Fees 50,000 0.14

Insurances 150,000 0.42

Legal Fees 50,000 0.14

Land Use Rent 100,000 0.28

Safety & Security


Safety/PPE 75,000 0.21

Training/Education 100,000 0.28

First Aid/Clinic 25,000 0.07

Army Security 109,000 0.30

Transportation


Staff & Workers' Transport 100,000 0.28

Light Vehicles Running 75,000 0.21

Community


Community Support 150,000 0.42

Road Maintenance 50,000 0.14

Environmental


Environmental Testwork 25,000 0.07

Environmental Monitoring 25,000 0.07

Laboratory

Annual Consumables 272,298 0.76

Import Duties, Taxes and Land Transport Fees

Tariff for Imported Consumables (5%) 66,338 0.18

Customs Fee for Imported Consumables (0.2%) 2,654 0.007

Land Transport Haulage Fee from Port to Site ($2,591/trip) 152,869 0.43

Subtotal - Personnel Salaries and Overhead 1,461,600 4.06

Subtotal - Other G&A Costs 2,024,959 5.62

Total 3,486,559 9.68

Page 21.54




AMENDED NI 43-101 TECHNICAL REPORT 21.4.11 Plant Operating Cost Summary

Operating costs have been developed using the plant parameters specified in the process design criteria

using a nominal throughput of 360,000 tons per annum. The operating cost estimate includes all the cost

items relevant to processing the ore by crushing and grinding, flotation, CIL, electrowinning and smelting

to produce gold doré. The operating costs listed by major category are presented in Table 21.30.

The process plant annual operating cost was determined to be $8.68 million, which equates to $24.11 per

tonne of ore processed or $171.48 per ounce of gold produced. With inclusion of G&A costs, the total

annual operating cost is $12.17 million, which equates to $33.80 per tonne of ore processed or $240.35

per ounce of gold produced.

Table 21.30 Summary of Process Operating Cost Estimate (nominal 360,000 tpa throughput)

Cost Category Total Cost Distribution

% $/y $/t ore $/oz Au

Process Plant

Process Plant Labour 2,122,822 5.90 41.94 17%

Operating Consumables 2,958,244 8.22 58.44 24%

Power 2,654,988 7.37 52.45 22%

Maintenance 944,027 2.62 18.65 8%

Subtotal - Process Plant 8,680,082 24.11 171.48 71%

General & Administration 3,486,559 9.68 68.88 29%

Total 12,166,641 33.80 240.35 100%

Figure 21.1 Overall Operating Cost Distribution (nominal 360,000 tpa throughput)

Page 21.55




AMENDED NI 43-101 TECHNICAL REPORT The major contributors to the overall operating costs are G&A (29%), consumables (24%) and power

(22%). The operating costs are presented by area in Table 21.31.

Table 21.31 Area Breakdown of Process Operating Cost Estimate (nominal 360,000 tpa

throughput)

Area Total Cost Distribution

$/year $/t ore $/oz Overall % Rel. %

Process Plant Labour 2,122,822 5.90 41.94 17% -

General and Administration

G&A Personnel Salaries and Overhead 1,461,600 4.06 28.87 12.0% 41.9%

Other G&A Costs 2,024,959 5.62 40.00 16.6% 58.1%

Sub Total - G&A 3,486,559 9.68 68.88 28.7% 100.0%

Consumables

Feed Preparation 25,958 0.07 0.51 0.2% 0.9%

Milling 434,729 1.21 8.59 3.6% 14.7%

Screening, Cyanide Detox and Tailings Disposal 415,721 1.15 8.21 3.4% 14.0%

Flotation and Regrind 200,252 0.56 3.96 1.6% 6.8%

Thickening, Leaching 712,745 1.98 14.08 5.9% 24.1%

Desorption, Carbon Regen., Refining, Electrowinning 269,005 0.75 5.31 2.2% 9.1%

Diesel Fuel (Mobile Equipment) 829,473 2.30 16.39 6.8% 28.0%

Other Consumables 70,343 0.20 1.39 0.6% 2.4%

Sub Total - Consumables 2,958,244 8.22 58.44 24.3% 100%

Power

Feed Preparation 102,045 0.28 2.02 0.8% 3.8%

Milling 1,025,367 2.85 20.26 8.4% 38.6%

Screening, Cyanide Detox and Tailings Disposal 301,414 0.84 5.95 2.4% 11.4%

Flotation and Regrind 254,487 0.71 5.03 2.1% 9.6%

Thickening, Leaching 230,300 0.64 4.55 1.9% 8.7%

Desorption, Carbon Regeneration 19,176 0.05 0.38 0.2% 0.7%

Refining, Gold Room, Electrowinning 31,317 0.09 0.62 0.3% 1.2%

Reagents 43,458 0.12 0.86 0.4% 1.6%

Water Services 76,606 0.21 1.51 0.6% 2.9%

Waste and Sewage Treatment 8,094 0.02 0.16 0.1% 0.3%

Air Services 304,112 0.84 6.01 2.5% 11.5%

Miscellaneous Facilities and Buildings 92,987 0.26 1.84 0.7% 3.5%

44 kV Line - Amortization Fee 165,624 0.46 3.27 1.4% 6.2%

Sub Total - Power 2,654,988 7.37 52.45 21.8% 100%

Maintenance 944,027 2.62 18.65 7.8% -

Total 12,166,641 33.80 240.35 100% -

ROM Throughput = 360,000 t/y; Metal Prod. = 50,620 oz Au

Page 21.56





Figure 21.2 Consumables Cost Distribution

Figure 21.3 Power Cost Distribution

Page 22.1




AMENDED NI 43-101 TECHNICAL REPORT 22.0 ECONOMIC ANALYSIS

22.1 Introduction

Lycopodium completed an economic analysis for the Santa Rosa Gold Project using the production

schedule along with capital and operating costs as described in Section 21.0. The analysis uses fully

diluted Proven and Probable reserves for the Project. The economic analysis generated a cash-flow

model, which is carried out on a pre-tax and post-tax basis. Tax implications were included based on

input from Red Eagle Mining. Net gross revenues were estimated using a $1,300 per ounce gold price

with revenue deductions for royalties, refining, transportation and insurance costs. Cash cost calculations

use the total operating cost plus royalties divided by the payable gold ounces. Total costs are calculated

using the total costs (operating and sustaining capital) plus royalties and taxes divided by the payable gold

ounces. The cash-flow model calculates the Net Present Value (NPV) based on a discounted rate of 0%


on total investment and the payback period were also calculated. A sensitivity analysis was also

conducted on parameters that are deemed to have the biggest impact on the Project financial

performance (capital cost, operating cost and gold selling price). The financial results are summarised in

Table 22.1.



NPV @ 0% K USD 171,763 131,501

NPV @ 5% K USD 136,895 103,678

NPV @ 8% K USD 119,921 90,027

IRR % 64.4% 52.6%






Page 22.2




AMENDED NI 43-101 TECHNICAL REPORT 22.2 Principal Assumptions

22.2.1 Life-of-Mine Process Plant Feed Schedule and Recovery

The economic analysis was calculated for the 8 year life-of-mine (LOM), including the ramp-up and ramp-

down years, according to the final production schedule prepared by MDA. The following provisions were

made in regards to schedule: construction and mine development start three months after the beginning of

detailed engineering; underground stope production is scheduled to commence two months prior to

commissioning; plant start-up and commissioning occurs 12 months after the commencement of

construction; and the plant ramp-up period is six weeks. Table 22.2, summarizes the annual values for

tonnes of fully diluted Proven and Probable reserves mined, ore milled, the feed grade and the metal

production of gold doré respectively.

Table 22.2 LOM Process Plant Feed Schedule and Metal Production

Year Ore Mined Ore Milled Feed Grade Metal Prod. Extraction

Tonnes Tonnes Au g/t Au oz %

-1 10,423 - - - -

1 359,230 352,118 6.21 67,675 96.2%

2 360,000 360,000 6.74 75,301 96.5%

3 360,000 360,000 4.49 49,587 95.4%

4 360,000 360,000 4.11 45,296 95.3%

5 360,986 360,986 4.67 51,786 95.6%

6 360,000 360,000 3.24 35,557 94.9%

7 360,000 360,000 2.77 30,335 94.7%

8 227,872 245,407 4.25 32,141 95.8%

LOM Total 2,758,511 2,758,511 4.57 387,678 95.7%

22.2.2 Gold Selling Price, Exchange Rate and Escalation

Economic modelling was completed using a US$ $1,300 per ounce gold price. The exchange rate of

$1,900 COP per US$ has been used for the capital and operating cost inputs to the cash-flow model. No

sensitivity analysis was conducted for fluctuations in exchange rate. Escalation is not applied in the cash-

flow model.

22.2.3 Revenue Deductions - Refining Costs and Royalties

Refining, transportation and insurance costs of $ $6.10 per ounce of gold produced are applied to the

Project cash-flow model, but are excluded in cash cost and total cost calculations ($/oz or $/t)

respectively. The refining, transportation and insurance cost estimate was provided by gold refiner budget

quotation.

Page 22.3




AMENDED NI 43-101 TECHNICAL REPORT Table 22.3 presents the “State” and “LMM” royalties applied to the cash-flow model. Producing mines in

Colombia are subject to a federal royalty of 4% of the gross value of gold and silver production at 80% of

the current London gold price so the royalty becomes in effect 3.2% (under a modification of mining law

685 of 2001). In Table 22.3, the “LMM” royalty of 3.0% refers to the royalty held by Liberty Metals and

Mining as described in Section 4.

22.2.4 Land Reclamation Cost

A land reclamation cost of $1 million is applied to the operating costs in the cash-flow model, in which one

third of this cost is covered in the second last year of production (Year 7) and the remainder is covered in

the final year of production (Year 8). This cost assumes a reasonable salvage value from the dismantling

of the process plant offsetting the reclamation cost.

22.2.5 Working Capital

A working capital allowance of $4.01 million is applied to the Project cash-flow model. This covers

2 months of operating costs (mine, process plant and G&A). This is included to fund the project from the

time production starts until the receipt of payment for the doré sales.

22.2.6 Taxes

Income taxes were estimated based on information provided to Lycopodium by Red Eagle Mining and its

financial advisors.

Corporate taxation and revenue streaming is applied to cash-flow model for the Project. Red Eagle

Mining is adopting revenue streaming between its branches in Colombia and Barbados. The cash-flow

model uses a corporate taxation rate of 33% on taxable income in Colombia and 1.75% on taxable income

in Barbados.

Red Eagle Barbados is to fund $25 million US dollars in equity, amortised over the 8-year LOM, for the

development by purchasing the right to buy all the gold doré produced from Red Eagle Mining de

Colombia for $ $1,200 per ounce. This forms the basis for the revenue streaming system of the Project.

The taxable income in Colombia included the gross free cash-flow less deductions for:

Exploration and acquisition cost amortization of $35 million, carried over 5 years;

Declining balance depreciation at 25% on capital spent, fully depreciated at the end of

LOM;

Using the revenue streaming system, any revenue above $1,200 per ounce is taxable in Barbados.

Taxes are summarized in the cash-flow presented in Table 22.3.

Page 22.4




AMENDED NI 43-101 TECHNICAL REPORT 22.2.7 Depreciation

Declining balance depreciation at a rate of 25% was applied to capital spent to reduce taxable income in

the cash-flow model once the mine is in production. The depreciation model and rate were provided by

Red Eagle Mining.

22.2.8 Cash-flow Model

A cash-flow model was created based on the production schedule, cost inputs, and economic parameters

previously discussed. Table 22.3 shows the resulting cash-flow model. The exploration phase of the

project and other project-related expenses are considered sunk costs, which are included in the cash-flow

model.

Page 22.5





Table 22.3 Santa Rosa Cash-Flow Model

Item Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Total

Revenues

Gross Gold Revenue K USD -$ 87,978$ 97,891$ 64,463$ 58,885$ 67,322$ 46,224$ 39,435$ 41,784$ -$ 503,982$

Refining, Transport, and Insurance K USD -$ 413$ 459$ 302$ 276$ 316$ 217$ 185$ 196$ -$ 2,365$

Royalty - Liberty K USD -$ 2,627$ 2,923$ 1,925$ 1,758$ 2,010$ 1,380$ 1,178$ 1,248$ -$ 15,049$

Royalty - State K USD -$ 2,802$ 3,118$ 2,053$ 1,875$ 2,144$ 1,472$ 1,256$ 1,331$ -$ 16,052$

Net Gross Revenue K USD -$ 82,136$ 91,391$ 60,183$ 54,975$ 62,852$ 43,154$ 36,817$ 39,009$ -$ 470,516$

Operating Costs

Mining K USD -$ 13,212$ 11,852$ 15,980$ 13,982$ 13,044$ 12,348$ 12,232$ 10,162$ -$ 102,811$

Processing K USD -$ 8,606$ 8,680$ 8,680$ 8,680$ 8,689$ 8,680$ 8,680$ 7,603$ -$ 68,299$

General & Administrative K USD -$ 3,487$ 3,487$ 3,487$ 3,487$ 3,487$ 3,487$ 3,487$ 3,487$ -$ 27,892$

Reclamation K USD -$ -$ -$ -$ -$ -$ -$ -$ 333$ 667$ 1,000$

Total Operating Cost K USD -$ 25,304$ 24,018$ 28,147$ 26,149$ 25,220$ 24,514$ 24,398$ 21,585$ 667$ 200,002$

Capital Cost

Mine Capital K USD 9,432$ 6,470$ 6,317$ 1,916$ 4,282$ 4,756$ 2,572$ 2,626$ 3,934$ -$ 42,305$

Process Capital K USD 45,776$ -$ -$ -$ -$ -$ -$ -$ -$ -$ 45,776$

Owners Team K USD 3,267$ -$ -$ -$ -$ -$ -$ -$ -$ -$ 3,267$

Working Capital K USD 4,014$ (4,014)$ -$ -$ -$ -$ -$ -$ -$ -$ -$

Contingency K USD 7,005$ -$ -$ -$ -$ -$ -$ -$ -$ -$ 7,005$

Import Duties K USD 399$ -$ -$ -$ -$ -$ -$ -$ -$ -$ 399$

Total Capital Cost - Excluding VAT K USD 69,892$ 2,457$ 6,317$ 1,916$ 4,282$ 4,756$ 2,572$ 2,626$ 3,934$ -$ 98,751$

VAT on Initial Capital K USD 4,309$ -$ -$ -$ -$ -$ -$ -$ -$ -$ 4,309$

Total Capital Cost - Including VAT K USD 74,201$ 2,457$ 6,317$ 1,916$ 4,282$ 4,756$ 2,572$ 2,626$ 3,934$ -$ 103,060$

Total Costs K USD 69,892$ 27,761$ 30,335$ 30,062$ 30,430$ 29,976$ 27,087$ 27,024$ 25,518$ 667$ 298,753$

Net Operating Cash Flow K USD -$ 56,832$ 67,373$ 32,036$ 28,826$ 37,632$ 18,640$ 12,418$ 17,425$ (667)$ 270,514$

Gross Free Cash Flow K USD (69,892)$ 54,375$ 61,055$ 30,120$ 24,544$ 32,876$ 16,068$ 9,793$ 13,491$ (667)$ 171,763$

Cash Cost $/oz Au -$ 454$ 399$ 648$ 658$ 567$ 770$ 885$ 752$ -$ 596.12$

Total Cost $/oz Au -$ 490$ 483$ 686$ 752$ 659$ 842$ 971$ 874$ -$ 670.56$

Depreciation

Amount Carried Forward K USD 65,878$ 65,878$ 49,409$ 37,056$ 27,792$ 20,844$ 15,633$ 11,725$ 8,794$ 6,595$

Deduction Taken K USD -$ 16,470$ 12,352$ 9,264$ 6,948$ 5,211$ 3,908$ 2,931$ 2,198$ 6,595$ 65,878$

Tax Considerations (Colombia)

Less: Exploration & Acquisition Costs Amortisation K USD -$ -$ 7,000$ 7,000$ 7,000$ 7,000$ 7,000$ -$ -$ -$ 35,000$

Less: Capital Allowance (25% declining balance) K USD -$ -$ 16,470$ 12,352$ 9,264$ 6,948$ 5,211$ 3,908$ 2,931$ 8,794$ 65,878$

Colombia Deferred Revenue From Streaming K USD -$ -$ 3,125$ 3,125$ 3,125$ 3,125$ 3,125$ 3,125$ 3,125$ 3,125$ 25,000$

Colombian Taxable Income K USD -$ -$ 23,250$ 37,298$ 12,023$ 9,192$ 18,611$ 11,729$ 3,739$ 3,941$ 119,782$

Colombia Corporate Tax K USD -$ -$ 7,672$ 12,308$ 3,967$ 3,033$ 6,142$ 3,871$ 1,234$ 1,301$ 39,528$

Less: VAT Recovery K USD -$ -$ 4,309$ -$ -$ -$ -$ -$ -$ -$ 4,309$

Net Colombia Corporate Tax K USD -$ -$ 3,363$ 12,308$ 3,967$ 3,033$ 6,142$ 3,871$ 1,234$ 1,301$ 35,219$

Tax Considerations (Barbados)

Barbados Taxable Income K USD -$ -$ 6,768$ 7,530$ 4,959$ 4,530$ 5,179$ 3,556$ 6,248$ 3,214$ 41,982$

Barbados Corporate Tax K USD -$ -$ 118$ 132$ 87$ 79$ 91$ 62$ 109$ 56$ 735$

Total Tax K USD -$ -$ 3,482$ 12,440$ 4,054$ 3,113$ 6,232$ 3,933$ 1,343$ 1,357$ 35,953$

Net After Tax Free Cash Flow K USD (74,201)$ 54,375$ 57,574$ 17,680$ 20,490$ 29,763$ 9,836$ 5,860$ 12,148$ (2,024)$ 131,501$

Page 22.6




AMENDED NI 43-101 TECHNICAL REPORT 22.2.9 Financial Results

Financial highlights for only the mining phase of the project are shown in Table 22.4. These financial

results include sunk costs.

The Project is forecasted to provide a pre-tax IRR of 64.4% (before tax but post royalty) and an NPV of

$137million, using a 5% discount rate. The payback period is 1.3 years after the start of production. Pre-

tax cash costs are $83.78 per tonne of ore milled or $596.12 per ounce of gold produced (excluding

capitalized pre-production costs and refining).

With the inclusion of duties, taxes, exploration and acquisition amortised costs, depreciation and revenue

streaming to Barbados, the Project is forecasted to provide a post-tax IRR of 52.6% and an NPV of $105

million, using a 5% discount rate. The payback period is 1.3 years after the start of production. Post-tax

total costs are 107.27 per tonne of ore milled or $763.30 per ounce of gold produced.



NPV @ 0% K USD 171,763 131,501

NPV @ 5% K USD 136,895 103,678

NPV @ 8% K USD 119,921 90,027

IRR % 64.4% 52.6%






22.2.10 Economic Sensitivity

A sensitivity analysis has been completed using the cash-flow model to investigate the sensitivity of the

undiscounted cash-flow, net present value, payback period, and internal rate of return based on changes

in operating costs, capital costs, and gold prices. The cash-flow model’s sensitivity to operating costs,

capital costs, and gold price is presented in Table 22.5 to Table 22.7. The pre-tax sensitivity of gold selling

price, net present value, and internal rate of return to changes in revenue, operating cost, and capital

costs are shown in Figure 22.1 to Post-tax sensitivity charts are shown in Figure 22.5 to Figure 22.8.

As typical with most precious-metals mining projects, the Project is most sensitive to changes in the

revenue or more specifically metal prices as shown in the slope of the revenue lines.

Page 22.7





Table 22.5 Economic Sensitivity to Operating Cost

Operating Pre-Tax Post-Tax

Cost Undisc. CF NPV 5% NPV 8% Payback IRR Undisc. CF NPV 5% NPV 8% Payback IRR

Sensitivity K USD K USD K USD Years % K USD K USD K USD Years %

80% 211,764 169,323 148,807 1.16 73.6% 158,301 125,913 110,085 1.24 60.2%

90% 191,764 153,109 134,364 1.20 69.1% 144,901 114,796 100,056 1.29 56.4%

100% 171,763 136,895 119,921 1.25 64.4% 131,501 103,678 90,027 1.34 52.6%

110% 151,763 120,680 105,478 1.31 59.6% 118,101 92,560 79,998 1.40 48.6%

120% 131,763 104,466 91,035 1.37 54.6% 104,156 81,079 69,681 1.46 44.5%

Table 22.6 Economic Sensitivity to Capital Cost

Capital Pre-Tax Post-Tax

Cost Undisc. CF NPV 5% NPV 8% Payback IRR Undisc. CF NPV 5% NPV 8% Payback IRR


80% 191,514 155,557 138,074 1.02 87.2% 144,733 117,158 103,613 1.09 71.0%

90% 181,639 146,226 128,998 1.13 74.6% 138,117 110,418 96,820 1.22 60.8%

100% 171,763 136,895 119,921 1.25 64.4% 131,501 103,678 90,027 1.34 52.6%

110% 161,888 127,563 110,845 1.38 56.0% 124,885 96,938 83,234 1.47 45.7%

120% 152,013 118,232 101,768 1.50 48.9% 118,268 90,198 76,441 1.60 39.9%

Page 22.8





Table 22.7 Economic Sensitivity to Gold Selling Price

Gold Pre-Tax Post-Tax

Price Undisc. CF NPV 5% NPV 8% Payback IRR Undisc. CF NPV 5% NPV 8% Payback IRR


1,000 62,671 45,925 37,585 1.87 29.3% 45,289 30,950 23,774 1.98 21.7%

1,050 80,853 61,087 51,308 1.72 35.8% 57,471 41,407 33,389 1.82 26.8%

1,100 99,035 76,248 65,030 1.60 41.9% 69,653 51,791 42,899 1.71 31.7%

1,150 117,217 91,410 78,753 1.50 47.8% 81,835 62,187 52,429 1.61 36.4%

1,200 135,399 106,571 92,476 1.40 53.4% 94,017 72,584 61,959 1.52 41.0%

1,250 153,581 121,733 106,198 1.32 59.0% 112,759 88,131 75,993 1.43 46.9%

1,300 171,763 136,895 119,921 1.25 64.4% 131,501 103,678 90,027 1.34 52.6%

1,350 189,946 152,056 133,644 1.19 69.8% 150,243 119,225 104,061 1.27 58.1%

1,400 208,128 167,218 147,367 1.13 75.0% 168,984 134,772 118,095 1.21 63.4%

1,450 226,310 182,379 161,089 1.08 80.2% 187,276 150,014 131,886 1.15 68.7%

1,500 244,492 197,541 174,812 1.04 85.4% 205,456 165,181 145,617 1.10 73.8%

1,550 262,674 212,703 188,535 0.99 90.5% 223,636 180,348 159,348 1.05 78.9%

1,600 280,856 227,864 202,257 0.95 95.6% 241,817 195,516 173,078 1.01 83.9%

Page 22.9




AMENDED NI 43-101 TECHNICAL REPORT Figure 22.1 Project Pre-Tax NPV @ 0% (Undiscounted C.F.) Sensitivity

Figure 22.2 Project Pre-Tax NPV @ 5% Sensitivity

$-

$50,000

$100,000

$150,000

$200,000

$250,000

$300,000

70% 80% 90% 100% 110% 120% 130%

Un

dis

cou

nte

d C

ash

-Flo

w,

K U

SD

Percent of Base Value

Capital Cost Operating Cost Gold Price

$-

$50,000

$100,000

$150,000

$200,000

$250,000

70% 80% 90% 100% 110% 120% 130%

NP

V @

5%

, K U

SD



Page 22.10




AMENDED NI 43-101 TECHNICAL REPORT Figure 22.3 Project Pre-Tax NPV @ 8% Sensitivity

Figure 22.4 Project Pre-Tax IRR Sensitivity

$-

$50,000

$100,000

$150,000

$200,000

$250,000

70% 80% 90% 100% 110% 120% 130%

NP

V @

8%

, K U

SD



-5.0%

5.0%

15.0%

25.0%

35.0%

45.0%

55.0%

65.0%

75.0%

85.0%

95.0%

70% 80% 90% 100% 110% 120% 130%

IRR

, %



Page 22.11




AMENDED NI 43-101 TECHNICAL REPORT Figure 22.5 Project Post-Tax NPV @ 0% (Undiscounted C.F.) Sensitivity

Figure 22.6 Project Post-Tax NPV @ 5% Sensitivity

$-

$50,000

$100,000

$150,000

$200,000

$250,000

$300,000

70% 80% 90% 100% 110% 120% 130%

Un

dis

cou

nte

d C

ash

-Flo

w,

K U

SD



$-

$50,000

$100,000

$150,000

$200,000

$250,000

70% 80% 90% 100% 110% 120% 130%

NP

V @

5%

, K U

SD



Page 22.12




AMENDED NI 43-101 TECHNICAL REPORT Figure 22.7 Project Post-Tax NPV @ 8% Sensitivity

Figure 22.8 Project Post-Tax IRR Sensitivity

$-

$50,000

$100,000

$150,000

$200,000

$250,000

70% 80% 90% 100% 110% 120% 130%

NP

V @

8%

, K U

SD



0.0%

10.0%

20.0%

30.0%

40.0%

50.0%

60.0%

70.0%

80.0%

90.0%

70% 80% 90% 100% 110% 120% 130%

IRR

, %



Page 23.1




AMENDED NI 43-101 TECHNICAL REPORT 23.0 ADJACENT PROPERTIES

Lycopodium has no information regarding exploration or development of properties adjacent to the Santa

Rosa Gold Project.

Page 24.1




AMENDED NI 43-101 TECHNICAL REPORT 24.0 OTHER RELEVANT DATA

24.1 Project Execution Strategy – Mining

This section describes the Project execution strategy proposed for the Santa Rosa Gold Project, for the

underground mine development and mine production.

The proposed execution strategy is based on the works being contracted to an experienced Colombian

mining contractor, managed and supervised by Red Eagle Mining.

Red Eagle Mining selected three potential contractors to pre-qualify for the mining works. Bids were

received from the three contractors, two of these were found to be technically compliant, with comparable

costs.

In June 2014 Red Eagle Mining issued tenders to the two bidders from the earlier pre-qualification

exercise. Another company expressed interest in the work, were deemed by Red Eagle Mining to have the

necessary experience and were also invited to bid. Tenders were received from the two pre-qualified

contractors in July 2014, the third bidder, after due consideration of the enquiry, declined to quote. Several

post-tender meetings were held to clarify the bids and establish the costs that have been used in the

Feasibility Study.

Technical Information Included in the Mining Contract Enquiry

Mining Technical Work Plan (Programa de Trabajo y Obras or "PTO");

Five year development plan (proposed contract for a five year term);

Proposed list of mining equipment;

Conceptual portal design;

Decline access site facilities layout;

Dimensions of the headings, drifts and raises;

Mine development and production schedule;

Services required;

Geotechnical reports; and

Hydrogeological reports.

The contractors also visited site on several occasions and inspected the drill core from the exploration

drilling programme.

Page 24.2




AMENDED NI 43-101 TECHNICAL REPORT Proposed Structure of the Mining Contract

The mining contract tender is structured into various packages of work and pricing methods, as follows:

Lump sum rates for mobilisation and site establishment;

Unit rates for development based on meterage;

Ore production on a time and materials basis;

Rates for any additional equipment proposed by the contractor;

‘Cost Plus’ work – All Inclusive Labour rates;

‘Cost Plus’ work – All Inclusive Material rates; and

Standby/Delay Rates and Charges.

In addition the contractors were requested to provide information on the following:

Statement of work experience;

Organization and supervision;

Manpower schedule and equipment requirements;

Departures and alternatives from tender documents and scope of work;

List of subcontractors (if any); and

Contractor's HSES program.

The two bids received were fully compliant in providing all of the costs and other information as requested

in the enquiry. The two bids were comparable in terms of cost.

Responsibilities of the Owner and the Mining Contractor

The following checklist shows the responsibilities of the Owner and the Contractor.

Page 24.3




AMENDED NI 43-101 TECHNICAL REPORT # Work Area Supply Install Maintain

Contr. Owner Contr. Owner Contr. Owner

1 Pre Job Preparation – Contractor’s Plant

Shipping and Packaging (seatainers, etc.) █ █ █

Packing for Transportation █ █ █

Initial Freight to Port, Cranage, etc. █ █ █

Offloading & Storage at Site █ █ █

2 Basic Site Facilities

Overall Site Security █ █ █

Local Site Security (Contractor Yard) █ █ █

Contractor Warehouse, and Stock Protection █ █ █

Owner’s Warehoused Materials – Dispensing and

Records; Backcharges as required █ █ █

3 General Site Operating

General Site Roads Maintenance █ █ █

Roads in Contractor Complex Maintenance █ █ █

Contractor Laydown Area █ █ █

Surface Fresh Water Distribution █ █ █

Mine Discharge Water Treatment █ █ █

Feed Lines to Discharge Water Ponds █ █ █

Concrete & Shotcrete Supply █ █ █

Electrical Power – To Portal █ █ █

Electrical Power – Underground Distribution █ █ █

Solid Waste Disposal (including HazMat) █ █ █

4 Compressed Air, Fuel and Lubricants

Compressed Air █ █ █

Fuel Ordering, Monitoring, Delivery █ █ █

Diesel Fuel █ █ █

Oils and Lubricants – Contractor’s Equipment █ █ █

5 Communications

Long Distance Service █ █ █

Local/Site Services █ █ █

Computer Internet Services █ █ █

6 Construction Supplies

Concrete and Cement █ █ █

Rebar, Ties, etc. █ █ █

Formwork and Accessories █ █ █

Concrete Inserts █ █ █

Anchor Bolts █ █ █

Portal Bents and Decking (for Portal) █ █ █

Page 24.4




AMENDED NI 43-101 TECHNICAL REPORT # Work Area Supply Install Maintain

Contr. Owner Contr. Owner Contr. Owner

7 General Safety Equipment and Supplies

Mine Lamp Chargers, Lamps █ █ █

Scaling Bars, etc. █ █ █

Self-Rescuers, Hearing Protection, etc. █ █ █

Miners; Belts/D Rings, Boots, etc. █ █ █

First Aid Consumables (other than those supplied

at Medical Aid Station) █ █ █

On Site Medical Aid Services █ █ █

Contractor’s Mine Rescue Equipment (2 Sets) █ █ █

8 Shop Related Items

Oxygen and Acetylene, Welding Rod, etc. █ █ █

Steel Plate and Shapes █ █ █

Shop and Hand Tools █ █ █

9 Surface Equipment

Site Crane (off-load, etc.) █ █ █

Materials/Supplies Handling █ █ █

Light Vehicles █ █ █

10 Miscellaneous Construction Items

Portable Refuge Stations(s) █ █ █

11 Mine Development

Contractor’s Maintenance Shop c/w Equipment █ █ █

Dev. Jumbos, LHDs, etc. █ █ █

U/G Haul Trucks █ █ █

Scissor Lift(s), Jeep, etc. █ █ █

Alimak Unit(s) as Required █ █ █

Raisebore or Longhole Unit if Applicable █ █ █

Temporary Fans █ █ █

Mine Dewatering Pumps █ █ █

Slimes Maintenance, etc. (U/G Sumps) █ █ █

Repair Parts for Above █ █ █

Roadbed Material for U/G █ █ █

12 Mine Development Supplies

Explosives and Accessories █ █ █

Drilling Material and Supplies █ █ █

Equipment Repairs, including Parts, Tires, etc. █ █ █

Explosives and Detonator Magazines █ █ █

Explosives and Materials Order/Delivery █ █ █

Explosives Security and Management Records █ █ █

Piping as Specified, c/w Hangers, etc. █ █ █

Face Pumps, c/w Hoses, etc. █ █ █

Ground Support Materials █ █ █

Surface Haul to Waste Dumps █ █ █

U/G Electrics, Cable, and Transformers █ █ █

Standard Rock Bolts/Dowels/Cable Bolts █ █ █

Non-Standard Rock Bolts/Dowels/Cable Bolts █ █ █

Page 24.5




AMENDED NI 43-101 TECHNICAL REPORT 13 General Contractor Support Services

Lateral Development, Raise & Construction

Surveying █ █ █

Mine Drawing Daily Updates █ █ █

Parts and Materials Expediting █ █ █

Crew Travel █ █ █

Routine First Aid █ █ █

Medivac/Equivalent █ █ █

14 Portal Construction and Conditioning

Mine Portal construction, bolts and all fasteners,

rebar, forming materials, and all other materials as

required to complete the Portal construction as

shown on the drawings.

█ █ █

15 Insurances, as shown in T&C █ █ █

16 Training

Site Orientation Leader █ █ █

Ongoing Mine Rescue Training █ █ █

Management of the Mining Contractor

Red Eagle Mining will provide all technical support, mine designs, survey control, and drawings of planned

development and production work to the contractor. The owner’s team will comprise all Colombian staff

unless otherwise noted:

Operations manager (expatriate mining engineer);

Chief mining engineer;

Mining engineers x 2;

Senior mine geologist;

Geology technicians/samplers x 3;

Core drillers x 2;

Grade control geologists x 2;

Surveyor;

Survey assistants x 2; and

Page 24.6




AMENDED NI 43-101 TECHNICAL REPORT Mine clerk.

Red Eagle Mining will be responsible for the underground delineation drilling, using their own drill rigs.

In addition to the above permanent staff, the owner has identified suitable expatriate personnel (2), with

experience in the proposed mining method (MSDF), who will be engaged for the first year of mine

production, for operational training of the contractor’s personnel. The cost of this mine training team is

included in the mine capital. This requirement is to address the lack of experience of the mining contractor

in the mining method to be adopted. Mine development is well within the capabilities of the contractors, as

this is closely allied to the tunnelling and infrastructure work that is the main source of work for this type of

civil engineering contractor in Colombia.

Contracting Strategy

To meet the Project schedule, it is necessary for the construction of the decline access portal to start at

the same time as the bulk surface earthworks. These two activities will be the first to occur on site and it is

anticipated that the contract for the bulk earthworks will be placed with the successful mining contractor.

This will reduce mobilisation costs, provide for ease of overall site management and enable optimum

utilisation of mobile equipment. Both of the bidding contractors have demonstrated their experience in civil

engineering and are both capable of executing the surface bulk earthworks contract.

24.2 Project Execution Strategy – Plant and Infrastructure

This section describes the Project execution strategy proposed for the Santa Rosa Gold Project. The

scope of work will include the process plant and all related project infrastructure.

The proposed execution strategy is based on an engineering, procurement, and construction management

(EPCM) implementation approach and horizontal discipline based packaging. Subject to final negotiations,

Lycopodium (the Engineer) will be engaged to provide EPCM services for the development of the process

plant and the associated infrastructure. Specialist consultants will be contracted to address specific

elements of the Project outside the core competency of the Engineer, mining, geotechnical,

environmental, dry waste management facility, etc.

A detailed project execution plan (PEP) will be developed by the Engineer at the beginning of the detailed

design stage. The plan will address all aspects of project execution and will establish a consistent project

delivery approach focused on logical sequencing of activities during all phases of the project: detailed

design, procurement, construction, and commissioning. The key sub documents of the PEP are the

engineering plan, health and safety management plan, contracts plan, procurement plan, construction

plan, and commissioning plan.

The key objectives during the execution phase of the project are listed below:

Page 24.7




AMENDED NI 43-101 TECHNICAL REPORT The attainment of the best HSE record possible – to accomplish this, it is required that

all contractors and involved personnel adhere to defined HSE objectives and standards

developed by Red Eagle Mining the EPCM contractor;

The timely and cost effective design, construction and commissioning of the process

plant and the associated infrastructure. Throughout all phases of project development,

the project team will remain conscious of the impact of decision making processes, be

they through design, procurement, transport or construction phases, etc.;

Leave a positive community relations legacy for on-going operations; and

Minimize environmental impact to the project site and the surrounding area.

24.3 Detailed Engineering and Design

An engineering plan will be prepared by the Engineer defining the principles and execution guidelines that

will be adopted during the design phase of the Project. The plan will identify the various engineering

deliverables required at the tender, procurement, construction, commissioning, close-out, and handover

stages.

At the beginning of the engineering phase of the project a priority will be given to finalising the process

design, freezing the plant layout, and issuing the deliverables supporting the procurement of the long lead

equipment – SAG mill, tower mill, and the filter presses.

Design reviews will be undertaken at predetermined stages of production of the technical documents and

models. The main objective of the design reviews is to verify that:

Statutory requirements, codes, and standards are complied with;

The technical documents meet the design requirements and are suitable for their

intended purpose;

Conflicts, unresolved issues, potential problems are identified, responsibility for

resolution assigned and completion dates scheduled; and

HSE requirements and lessons learned have been addressed.

Hazards and operability (HAZOP), constructability, and other reviews will be scheduled as appropriate.

Model reviews will be conducted at 30, 60, and 90% of engineering development.

Technical peer reviews for the project will be undertaken for the principal design documents and for any

significant technical risk items identified in the risk register.

The engineering discipline leads will ensure all the external documents are reviewed by all disciplines in

accordance with the procedure as set in the engineering plan.

Page 24.8




AMENDED NI 43-101 TECHNICAL REPORT 24.4 Procurement Management

The procurement scope will cover all formal tendering of packages to achieve competitive pricing and an

effective negotiating position to provide value for money to Red Eagle Mining. Tendering will be based on

a lump sum cost for equipment design and supply and on a “Schedule of Rates” basis for bulk materials

supply, where applicable. Equipment suppliers will be selected on the basis of previous experience, ability

to meet design requirements and the project schedule.

A detailed project bidders list will be developed using global sourcing prequalified vendors and the Red

Eagle Mining’s preferred suppliers list. Following the compilation of technical requirements, tender

packages will be prepared and issued to the approved bidders. Bidder’s questions will be addressed,

clarification notices will be prepared and issued as necessary, and bids formally received.

Bids will be evaluated on the basis of obtaining the best value in terms of price, delivery, and equipment

quality. Evaluation criteria will be developed for critical packages.

Technical evaluation will fundamentally be based on cost and technical compliance, but will also consider

experience, ongoing operational and maintenance support, consumable spares, reliability, HSE

performance, reputation, etc. Meetings, where required will be arranged with the bidders to clarify and

confirm technical and commercial matters and to finalize and obtain a satisfactory agreement for Red

Eagle Mining whose representative will be invited to participate in these meetings.

Participation by Colombian suppliers will be pursued to the maximum possible extent on the basis of

quality, schedule, overall cost effectiveness, previous experience, and availability to perform the work.

Direct negotiations with smaller local business groups on specific packages will be planned to encourage

local sourcing of equipment and material.

Transport, logistics, customs clearance, and expediting services will be managed by the Project’s

transport and logistics contractor. The majority of the purchase orders will be based on ex-works basis.

Purchase orders for the supply of materials and equipment will specify packaging requirements to cater for

sea freight and inland transport. The Engineer will supply the specification to which packaging of all

materials and equipment must comply.

The transport and logistics contractor will interface with the materials control personnel established by Red

Eagle Mining on site. They will allocate lay-down areas for installation contractors to carry out their

responsibilities for the receipt and storage of the “free issue” supplied equipment and material.

The installation contractors will be made responsible for off-loading of the material and equipment related

to their contracts and this will be accordingly written into their respective contracts.

Page 24.9




AMENDED NI 43-101 TECHNICAL REPORT Procurement of long lead equipment will be given the highest priority. The tender package preparation,

manufacturing, and installation of these items form the critical path for the project schedule. All the

engineering deliverables associated with these packages will be identified early in the engineering design

phase and expedited to the maximum possible extent. A preliminary list of long list items and approximate

lead times is shown on Table 24.1.

Table 24.1 Procurement Long Lead Items

Equipment Lead Time (weeks)

SAG Mill 36

Tower Mill 38

Flotation Cells 32

Thickener 32

Pressure Filter 36

Apron Feeder 28

24.5 Contracts Management

24.5.1 Strategy

The Engineer will be responsible for the preparation and implementation of the contracts plan and

organization structure of the project. The plan will be developed and updated progressively through the

detailed design phase. The following items will be covered in the contracts plan:

Development of well defined Contracts Work Packages (CWPs);

Contractor’s prequalification;

Development of individual contracts plans setting out the scope, free issue materials,

and equipment, timing and interfaces between each work package;

Incentive programs for safety, cost, or performance milestones;

Performance measurements and progress reporting;

Scheduling and package durations; and

Special considerations for project specific site conditions.

The contracts plan for project work begins with the development of well defined CWP’s based on the

overall organizational structure and allocation of risk among the contracting parties. This is followed by

tendering, award, and administration of these packages in line with Red Eagle Mining’s policy and

procedures.

Page 24.10




AMENDED NI 43-101 TECHNICAL REPORT Horizontal discipline based packaging has been selected for the construction of the process plant and the

associated plant infrastructure based on the following advantages:

Horizontal packaging provides greater flexibility to engage the expertise of medium

sized specialist contractors as required for the various construction packages for the

duration of the Project;

A horizontal contractor will be a medium sized local company, specialising in a specific

discipline and usually this will result in a lower price (given medium sized companies will

have lower overheads) and a more efficient site team;

A horizontal contract work package allows the Engineer to have direct control of the

contractor undertaking the specific scope of work;

Contractors normally maintain strength in one or two major disciplines. In a vertical

packaged scenario, the work in those disciplines for which the principal contractor does

not possess in-house expertise must be subcontracted to other groups over which the

principal contractor may not have effective control and which may incur additional costs

and schedule delays;

There is a greater pool of horizontal discipline based local contractors available than

vertical contractors; and

The horizontally packaged contractor may have a shorter duration on site which may

provide time and benefits as well as HSE benefits.

Contractors for site works will be selected on the basis of their safety record, previous experience with

similar projects, cost, schedule, availability, and capability to perform the work.

Local contractors and suppliers in the Antioquia region will be encouraged to tender for any work package

and contracts will be awarded based on their ability to meet the required conditions.

Priority will be given to fixed lump sum contracts where the volume of work is assessed by the contractors

from drawings, material take-offs, and other technical information provided during the tendering process.

These contracts are usually subjects to only minor contract variations.

Other types of contracts - Schedule of Rates (SoR) or Hourly Hire - will only be used where a fixed lump

sum contract is not practical. The Project schedule and/or practicalities dictate that earthworks, concrete

works, structural steel fabrication, platework fabrication, structural, mechanical and piping (SMP) and

electrical and instrumentation (E&I) installation packages will likely need to be undertaken under schedule

of rates contracts. With SoR type contracts the contractor will assume productivity risk and will be paid for

the actual quantities supplied and/ or installed.

The discipline and type of major contracts proposed for the Project are listed in Table 24.2.

Page 24.11




AMENDED NI 43-101 TECHNICAL REPORT Table 24.2 Proposed Major Contracts

Contract Title Work Scope Commercial Terms

Site Services Contract The EPCM in conjunction with Red Eagle Mining will prepare contracts for service

providers for:

Waste collection

Site NDT testing

Third part expediting and inspection services

Site Surveying

Geotechnical and soils testing

Lump Sum / Schedule

of Rates

Bulk Earthworks Clear and grub Plant site and DWMF areas. Mass excavation, cut to fill, cut to

waste and backfill of process plant area, portal pad area, event and monitoring

ponds and misc site earthworks.

Installation of water control ditches and site dewatering, drains and culverts.

Upgrade of main access road including drains and culverts.

Construction of the DWMF starting dyke, tailings pipeline corridor, seepage and

sediment ponds, top soil deposit dyke.

Underground services including trenching, supply and installation of service piping

and backfill to nominal construction grade. Construction of Access road from the

concession gate to the portal pad (400m)

Schedule of Rates

Contract

Concrete Supply and Installation Detailed excavation, Forming, Re-bars supply and installation and embedded

items; Placing and finishing of concrete, Supply and placing backfill against

concrete

Schedule of Rates

Contract

Transport and Logistics Services Coordination and management of sea and road transport including securing related

permits and documents to deliver all materials and equipment from port / supplier /

vendor / contractor facility to site.

Customs clearance and brokerage service for materials and equipment from

outside of Colombia.

Provision of supervision, labour and handling equipment to off-load materials and

equipment on site

Schedule of Rates

Contract

Structural Steelwork Fabrication Shop detailing, supply, fabrication, surface treatment and delivery to site of all

structural steel, grating, handrails, stairs, vertical ladders and miscellaneous items.

Schedule of Rates

Contract

Platework Fabrication Shop detailing, supply, fabrication, surface treatment and delivery to site of all

platework – including, hoppers, chutes, launders, bins, shop fabricated tanks,

underpans and miscellaneous platework items.

Supply will include all lining materials and fasteners i.e. rubber, polyethylene, liner

plates etc.

Schedule of Rates

Contract

Field Erected Tanks Shop detailing, supply, fabrication, surface treatment and delivery to site,

installation, assembly, hydro testing and final site painting of all field erected tanks

Lump Sum

Plant Conveyors Design, supply, fabrication, surface treatment and delivery to site of all plant

conveyors systems including all mechanical items, electrical items,

instrumentation, structural steelwork, walkways, handrails, grating, belting, skirts,

shedder plates, guarding, chute work and all other miscellaneous items for a

complete operating system.

Lump Sum

Structural, Mechanical and Piping

Installation Works

The scope will cover structural, platework and mechanical equipment installation

and piping supply, fabrication and installation. The scope will also include pre-

operational testing and pre-commissioning.

Part Lump Sum/Part

Schedule of Rates

Contract

Electrical and Instrumentation

Works

The scope will cover the supply of materials and installation of all electrical and

instrumentation works for the process plant and infrastructure.

Work includes all electrical testing and loop checking.

Part Lump Sum/Part

Schedule of Rates

Contract

Prefabricated Modular Buildings Design, supply, transport to site, installation and commissioning of all prefabricated

modular buildings including associated services and fit out.

Control rooms, Electrical switch rooms, Plant administration building, and Guard

house.

Mostly Lump Sum but

part schedule of rates

Assay/Metallurgical Laboratory Design, supply, transport to site, installation and commissioning of the assay and

metallurgical laboratory including floor coverings, curtains and all other fit out,

ventilation systems, laboratory equipment and

Furnishings.

Lump Sum

Pre-Engineered Buildings Design, supply, transport to site and installation of:

Plant workshop and main warehouse building. Includes the workshop and

warehouse fit out.

Lump Sum

44 kV Overhead Power Line Design, supply and installation of 44kV overhead power line (8.9 km) from the EPM

substation to the project site switchyard including:

Supply of all materials.

Transport to site, handle and install civil works electromechanical works; and

Pre-commissioning and energise.

Lump Sum

Site Security Hardware Design, supply, installation and commissioning of the site security system for

offices and gold room

Lump Sum

Security Fencing Supply and installation of site security fencing and gates. Lump Sum

Page 24.12




AMENDED NI 43-101 TECHNICAL REPORT 24.5.2 Bulk Earthworks

The scope of this contract will include all the clearing and grubbing; mass excavation and backfill in the

plant site and the DWMF area; all access and haul roads at the site including the 400 m long main access

road from the concession gate to the main portal; all ditches, dewatering drains and culverts together with

the underground trenching including pipe supply and backfill. The DWMF starter dyke, the top soil deposit

dyke, the event, monitoring, seepage, and sedimentation ponds, and the tailings pipeline corridor will also

be part of this contract.

The bulk earthworks contractor will be the first contractor to mobilize to site. Due to the nature of the

contract and the lack of final engineering deliverables during the tendering phase and award for this

package this will be a “Schedule of Rates” contract. The work will start at the plant site and will progress

east towards the DWMF, the decline portal, and top soil deposit areas.

Every effort will be made to secure the services of a reputable local contractor experienced in this type of

work. It will be beneficial from cost and schedule perspective if the mining contractor can execute the bulk

earthworks contract. This will be investigated in the next phase of the project.

At this point the scope of work related to the upgrades required at the main access road from Highway 25

to the concession gate (approximately 8.2 km) is not finalized. These upgrades will be required in order to

facilitate shipping large equipment to the Project site during the construction period. Once the scope is

finalized a decision will be made if the bulk earthworks contractor can also execute this work package.

24.5.3 Concrete Supply and Installation

Based on the estimated volume of concrete required to be poured on site (2,400 m3) and a five month

schedule a single contractor will be engaged to complete this work. The contractor will be responsible for

detailed excavation, forming, re-bars supply and installation, all embedded items, and placing and

finishing concrete. The placing of granular fill against the finished concrete surface will also be included in

this scope. The contractor will be responsible for preparing all bending schedules and holding down

bolting schedule. Taking into account that the largest pour on site will be the 120 m3 raft foundation for the

SAG mill it is assumed that a medium size mixing plant operated by the contractor will be sufficient to

supply the concrete mix. Several local contractors have been identified capable to handle the concrete

supply and installation scope. Based on the current execution schedule the project may not be in a

position to award the contract on a lump sum basis, a “Schedule of Rates” contract with an option to be

converted to a lump sum (when all the final quantities are available) will be the likely contract strategy for

this construction work package.

Right from the start the focus will be on the ring beam foundations in the CIL tanks area since the field

erected tankage contractor will start their scope in this area. The crushing area concrete works and the

SAG mill raft foundation are also on the critical path due to the large volumes of concrete involved and the

lack of float in the SMP installation in these areas.

Page 24.13




AMENDED NI 43-101 TECHNICAL REPORT 24.5.4 Field Erected Tankage

This contract scope will cover the CIL and some other large tanks at the plant site. Smaller tanks can be

fabricated off-site. This construction package will include the shop detailing, fabrication, transport to site,

and installation of the field erected tanks. The scope of the contract will also cover hydro testing and final

site painting of all the field erected tanks. This contractor will mobilize to site as soon as the concrete

contractor can provide access to the CIL ring beams. Since the tankage plate quantities are not expected

to change significantly the plan is to let this contract on a lump sum basis.

24.5.5 Structural Steel Supply and Fabrication

This construction package will be medium scale (around 380 t of steel) and will include the shop detailing,

supply and fabrication, painting, transportation to site of all structural steel, grating, stairs, and vertical

ladders. The supply of building cladding and roofing will also be part of this scope. This will be a “Schedule

of rates” contract with an option of converting portions of it to a lump sum contract when the final

quantities are available. Several local / regional contractors were identified to have the capabilities and

appropriate experience to undertake this scope.

24.5.6 Structural, Mechanical and Piping (SMP) Installation Works

The SMP package will include the erection of structural steelwork, platework and mechanical equipment

installation, and piping supply, fabrication and installation. This is one of the most important contracts on

site and the planning and monitoring of its progress will be given the highest priority. The engineering

deliverables related to this scope will not have sufficient level of detail in order to award a lump sum

contract. The SMP package will be awarded as a part lump sum contract and part “Schedule of Rates”

contract. The activities related to this scope are on the critical path of the project. It will be imperative to

mobilize the contractor as early as possible to build sufficient float in the schedule. A detailed description

of the SMP contract package scope is given below:

Offloading, handling and installation of all free issue structural works, including

steelwork, grating, floor plate, handrails, stair treads, purlins, girts, sheeting, security

mesh, guarding, fasteners, and miscellaneous items including levelling, aligning and

grouting;

Offloading, handling and installation of all free issue mechanical equipment and

platework items, including levelling, aligning and grouting, including first fill lubricants

and pre-operational testing and cleaning, rotation and testing of equipment;

Offloading, handling and installation of all free issue valves, instrumentation, fire

hydrants and extinguishers and miscellaneous in-line items;

Preparation of isometric drawings, materials take-offs, supply of all piping, fittings,

gaskets and fasteners, shop fabrication, NDT testing, surface treatment and transport to

site of pipe spools and stock length piping and fittings;

Page 24.14




AMENDED NI 43-101 TECHNICAL REPORT Preparation of shop detailed drawings, supply of all materials, shop fabrication, surface

treatment and transport to site of pipe supports;

Supply and installation of all insulation materials and pipe and valve identification

materials for pipe systems;

Installation of all piping systems including in-line items, free issue control and manual

valves, pipe supports, pipe guides for a complete pipe system;

Flushing and service leak testing of all piping systems;

Flushing and hydro testing of all buried service piping systems; and

Flushing, pickling, hydro testing of all hydraulic, fuel and acid piping systems.

24.5.7 Electrical and Instrumentation (E&I) Installation Works

The E&I contract package will include the supply and installation of all electrical and instrumentation works

for the process plant and the associated infrastructure. The contract will be tendered and awarded on the

basis of preliminary engineering deliverables and as such will be mostly “Schedule of Rates” contract with

a small portion contracted as a lump sum. This will be the last contractor to mobilize to the site. Parts of

the plant that are on the critical path in terms of E&I installation are the milling and regrind areas due to

the late arrival on site of the SAG mill and the tower mill. It is estimated that this contract will be completed

in 6 to 7 months. The scope of work for the E&I construction package will include:

Offloading, handling and installation of all free issue electrical equipment and

instruments items, including levelling, aligning and grouting;

Supply, transport to site and installation of all supports, cable tray materials, conduit,

earthing cable, power cables, control cables, small power and lighting, fasteners and all

miscellaneous items;

Installation of required electrical components from LV side of the unit substation

transformers to plant distribution and termination;

Installation and termination of owner supplied instruments; and

Electrical testing and loop checking, pre-commissioning and testing of all E&I works.

24.6 Construction Management

The construction methodology proposed for the Santa Rosa Gold Project has the following aims:

To provide a safe working environment;

Page 24.15




AMENDED NI 43-101 TECHNICAL REPORT To achieve cost and schedule targets;

To adopt a cost effective and fit for purpose construction methodology in contracting and

site management based on tried and proven philosophies;

To allow optimisation in constructability;

To provide a management plan that complies with the requirements of both Red Eagle

Mining and the Engineer’s safety and environmental policies; and

To achieve maximum possible utilization of local resources.

A construction management execution plan will be prepared to provide a project specific statement and

work plan on how the Engineer will organise, perform, and execute the construction management

responsibilities for the Project. The plan will define the interfaces with engineering, procurement, and

commissioning, to ensure construction is executed in a timely, cost effective manner in accordance with all

project objectives. The construction plan will document the intended construction approach to the Project

starting with early site capture through commissioning and start-up. The intent of this document is to

clearly lay out the planned approach to construction for understanding by all Project stakeholders.

The Engineer’s construction management team will manage and coordinate all construction activities

within the scope of the Project to ensure control over cost, schedule, and quality. The construction

manager will be located in the engineering office during the first few months of the project and then

transferred to the site when construction commences. He will define the specific duties of key construction

management personnel to suit the construction requirements of the Project. He will be responsible for the

overall construction planning, cost, and scheduling. He will also ensure that all aspects of the work are

properly set up with the necessary project controls for items such as: planning and scheduling; cost

control, document control, accounting, project risk analysis, forecasting, trending, and change control.

The field engineering team will be responsible for providing engineering design support on any technical

matters arising during construction.

24.7 Commissioning

24.7.1 General

The main objective of commissioning is to safely introduce production material to the process plant on the

earliest possible date and turn over to Red Eagle Mining an integrated plant capable of continuous and

reliable performance.

Page 24.16




AMENDED NI 43-101 TECHNICAL REPORT A commissioning manager in conjunction with Red Eagle Mining’s process plant manager and his team

will plan, coordinate, and execute all pre-commissioning and commissioning activities. The pre-

commissioning, commissioning, turnover and acceptance methodology to be used for the facilities will be

a systems based approach. The total scope of facilities for the Project will be divided into sub-systems

based on operational function. A systems index will be developed and maintained as a key control

document. The sub-systems will be grouped into operable systems that will be identified on the systems

index and on the Project schedule. Turnover packages including pre-commissioning and commissioning

documentation will be managed by operable systems with detailed tasks being performed at the sub-

system level.

Upon completion of work (system completion) for the facilities, critical pre-commissioning and

commissioning activities shall be performed in order to confirm all parts of the facilities are in good working

order and meet the minimum acceptable performance requirements.

24.7.2 Pre-commissioning and Testing

After verification that the plant has been constructed in accordance with the design (conformance with

P&IDs and drawings), construction and installation testing would typically include hydrostatic pressure

tests, flushing of lines, alignment checks, electrical point to point checks, and component identification

checks. Dry commissioning includes motor direction tests, all drives run, conveyors run and tracked,

instruments checked, control system verified and facility sequence testing. By the conclusion of pre-

commissioning all equipment and systems must be cleaned out.

The construction manager, via the construction supervisors, is responsible for managing all pre-

commissioning activities, along with recording and approval of results. The testing will be conducted by

the appropriate contractor. The commissioning manager will assist with coordinating the dry

commissioning phase of pre-commissioning.

24.7.3 Mechanical Completion

Mechanical completion of a section of the plant is achieved when pre-commissioning is complete and it

meets all requirements with respect to design, safety, physical operability and specifications and the

relevant module is ready for extended operation and/or the introduction of ore / process fluids.

24.7.4 Wet Commissioning

Wet commissioning consists of successfully testing and operating the equipment grouped together into

systems or modules, but without ore or reagents or other process material. At the successful conclusion of

wet commissioning, ore / process fluids are introduced into the circuit and process commissioning

commences.

Page 24.17




AMENDED NI 43-101 TECHNICAL REPORT 24.7.5 Process Commissioning

Process commissioning follows the successful completion of wet commissioning. During this time, the

initial introduction of ore and reagents to the process will occur. The circuit will be operated to achieve

nominal throughput and metallurgical performance. Process commissioning will be managed by the

commissioning manager using Red Eagle Mining’s operating personnel. The commissioning will be

performed at a multi-system level, incorporating systems defined in the process functional specifications.

The system start-up sequence will follow the order defined in the process flow sheets for the start-up of

the process plant under normal operation. Completion of this phase is achieved once the milestone

production rates for each system within the process plant have been achieved.

24.8 Project Close-out and Handover

Project close-out involves finalising all outstanding issues and work items when the work is complete and

the Engineer’s responsibilities end. At the completion of all construction and commissioning activities, the

Engineer will provide the following close-out information to Red Eagle Mining:

As-built drawings;

Piping and instrumentation diagrams (P&IDs);

Electrical as-built drawings;

Commissioning data and records;

Quality records;

Project close-out report; and

Operating manuals and recommended spares lists.

The Engineer will create and issue for Red Eagle Mining’s sign off a handover certificate reflecting the fact

that the plant is complete and operational, has been commissioned, that all performance warranties have

been achieved and is fully functional.

24.9 Project Execution Schedule

A Project execution schedule has been prepared as part of this Feasibility Study. The schedule is

provided in Fig 24.1. The estimated project duration is 16 months. The engineering activities will take

approximately 9 months, the site construction activities will be completed in 12 months followed by

commissioning. This schedule is based on Lycopodium’s understanding of the project scope, current lead

times for the delivery of critical equipment (SAG mill and tower mill), and typical duration of engineering

and site activities based on similar size projects executed by Lycopodium. The major project milestones

are summarised in Table 24.3 below.

Page 24.18




AMENDED NI 43-101 TECHNICAL REPORT Table 24.3 Major Project Milestones

Major Milestone Month

Start of Detailed Engineering Month 1

Award SAG mill package Month 2

Award Tower mill package Month 2

Frozen PDC and Flow sheets Month 3

Award Bulk Earthworks Contract Month 3

Start Construction (Bulk Earthworks) Month 4

Start Construction of mine portal Month 4

Start Concrete Works Month 6

Start Field Erected Tankage Month 8

Complete Bulk Earthworks Month 9

Start SMP Installation Month 9

Detailed Engineering Complete Month 9

Start E&I Installation Month 10

Concrete Works Complete Month 11

Field Erected Tankage Complete Month 12

SAG Mill and Tower mill arrive on site Month 12

SMP Installation Complete Month 15

E&I Installation Complete Month 16

Start Commissioning Month 15

Commissioning Complete Month 16

The critical path for the Project has been identified as follows:

Tender and Place SAG mill and mower mill orders;

SAG mill and tower mill manufacture and transport to site;

Milling area SMP installation;

Milling area E&I installation;

Plant services area E&I installation;

Pre-commissioning; and

Commissioning.

Page 24.19





Figure 24.1 Project execution schedule

Page 25.1




AMENDED NI 43-101 TECHNICAL REPORT 25.0 INTERPRETATION AND CONCLUSIONS

25.1 Geology and Mineral Resource

The San Ramon deposit is a mesothermal, shear-zone-hosted gold deposit within the Cretaceous

Antioquia Batholith in northern Colombia. The age of mineralization is not specifically known, but

mineralization could have taken place at any time from the late Cretaceous to Tertiary. The shear zone

and mineralization strike east-west and dip predominantly 70° to 85° to the north; the dip shallows to 50°

to 60° at depth. The known deposit, roughly 2,000 m in strike length, is well-defined at the west end but is

open-ended down-plunge to the east.

The geological model completed in 2013 and used in this Feasibility Study imparts a high level of

confidence in the resource and in the understanding of the project. It has laid the groundwork for clear

and effective exploration, as well as resource estimation. The geologic models of the shear zone, country

rock, redox, and saprolite are based on core photos; detailed drill-hole logging of lithology, alteration,

weathering, oxidation and structure; and lithology, redox, and weathering summaries normalized for

consistent interpretations. Red Eagle Mining geologists provided continuous interpretive input and

oversight. QA/QC programs have been in place to monitor the quality of San Ramon assays, and the

results of the programs are generally positive.

A high level of confidence in the estimated global gold content at San Ramon deposit is demonstrated by

the amount of the Measured and Indicated material (75%) in the resource; the remainder is Inferred. Most

of the Inferred material is due to wide-spaced drill density (~100 m) below 200 m to 250 m depth.

25.2 Mining and Mineable Reserves

Mineable reserves were developed using the resource modelled high-grade domains along with undiluted

grade estimates. The high-grade domains were used as a basis for stope designs. Measured and

Indicated resources above and below the economic cut-off grades (oxide 1.96g Au/t, transition 2.14g Au/t,

and sulfide 2.00g Au/t) were determined for the stope solids. Dilution was deemed necessary due to the

narrowness of the stope solids (averaging 3m), and 0.25m selvedge was added to each side of the stope

outlines. Dilution with Measured and Indicated material was added at modelled grades; waste and Inferred

dilution material was added at zero grade.

A mining method was developed that was considered suitable to the geometry of the stope solids, and the

geotechnical characteristics of the shear zone containing the high grade domains. The method developed

was Mechanized Shrinkage with Delayed Fill (MSDF). This method has the advantages of:

dilution and ore loss during mucking are greatly reduced;

the powder factor required is reduced in comparison to the similar method of cut-and-fill

mining;

leaving mined ore in the stope maintains wall stability until the ore is mucked out;

Page 25.2




AMENDED NI 43-101 TECHNICAL REPORT underground stockpiles of ore will be available that can be delivered as required to the

mill; and

easy to convert to cut-and-fill techniques where ground conditions become weak.

Life of mine production schedules were created based on required development and the designed stopes.

The schedules were done on a monthly basis and include productivity from contractor quotations. These

schedules were used for first principle cost evaluation and economic analysis.

Reserves have been classified in order to increase confidence into Proven and Probable categories to be

in compliance with the “CIM Definition Standards – for Mineral Resource and Mineral Reserves” (2014)

and therefore Canadian National Instrument 43-101.

25.3 Metallurgical Testing and Recovery Methods

Metallurgical testwork has been performed on numerous sample composites from the San Ramon

deposit. The earliest program was intended to characterize the major lithology, redox, and weathering

types at a wide range of grades. Variability within the unoxidized shear zone material was addressed in

the 2013 program. However, only two of six of the composites were of grades above the potential

underground-mining cut-off. The most recent metallurgical program focused on testing composites that

are representative of underground-mineable material. These composites were of grades exceeding 2.0 g

Au/t and are comprised of unoxidized and unweathered material spatially distributed within the shear

zone, which represents the bulk of the economic mineralization.

Metallurgical laboratory testwork achieved the desired quality and demonstrated that by using the

designed process flowsheet it is possible to economically recover gold from the San Ramon deposit.




process prior to the tailings being filtered. Part of the filtered tailings will be stacked in a dry waste

management facility; the balance will be used as backfill in the mine. Filtrate will be recycled back to the

process plant to minimise the raw water requirement.





Site infrastructure facilities in support of the mining and processing of the San Ramon deposit have been

developed to take into consideration the local topographic features, water courses and access. The plant

site footprint has been optimised to take into account local topography and minimise visibility and noise

impact to the communities nearby. The level of the detail and planning is commensurate with that normally

associated with Feasibility Study level.

Page 25.3




AMENDED NI 43-101 TECHNICAL REPORT There is adequate space for the intended facilities and there are no known adverse conditions that could

affect the design and construction of the required project infrastructure. The layouts of equipment and

building sizes including auxiliary buildings, maintenance facilities, and power supply infrastructure will be

compatible with other similar size projects.

The current level of infrastructure design incorporates all necessary facilities to protect the environment

including water treatment, sewage disposal, surface water catchment, retention ponds and recirculation

systems.

25.5 Capital Cost Estimate


contingencies for all the facilities required to bring the Santa Rosa Gold Project into production, as defined

by this Feasibility Study. All equipment and material are assumed to be new. The labour rate build up is

based on the statutory laws governing benefits to workers in effect in Colombia at the time of the estimate.

Colombian import tariffs have been applied.

The mining capital cost estimate has been costed from the mine development design using firm contractor

mining tendered prices.

The estimate does not include any allowances for escalation, exchange rate fluctuations or project risks.

The execution strategy for the process plant and infrastructure is based on an EPCM implementation

approach and horizontal (discipline based) packaging, in conjunction with the Red Eagle Mining

construction and management team. The capital cost estimate has a predicted accuracy of +/- 15%.

The total estimated cost of the overall project (mine, process plant, and infrastructure) is 69.90 million.

This amount excludes recoverable value added taxes.

25.6 Operating Cost Estimate

Mine operating costs have been developed from detailed stope production scheduling and equipment

cycle times. Firm contractor mining man and equipment hire rates were then applied to the results of these

exercises to obtain costs per tonne of ore mined.

The mine annual operating cost was determined to average 37.36 per tonne of ore mined. In addition with

expensed development capital averaging approximately 4.0 M dollars per year or 11.40 per tonne of ore

mined. Process plant operating costs have been developed using the plant parameters specified in the

process design criteria. The operating cost estimate includes all the cost items relevant to processing the

ore by crushing and grinding, flotation, CIL, electrowinning and smelting to produce gold doré.

The process plant annual operating cost was determined to be 8.87 million US, which equates to 24.64

per tonne of ore processed or 175.23 per oz of gold produced. With inclusion of G&A costs, the total

annual operating cost is 12.38 million US, which equates to 34.38 per tonne of ore processed or 244.53

per oz of gold produced.

Page 25.4




AMENDED NI 43-101 TECHNICAL REPORT 25.7 Economic Analysis

In the economic analysis, the compilation of the operating, sustaining, capital, taxes, royalties and other

costs reflects the mine development and operating scenario.

The analysis uses fully diluted Proven and Probable reserves for the Project. The generated cash-flow

model was carried out on a pre-tax and post-tax basis. Net gross revenues were estimated using a 1,300

per ounce gold price with revenue deductions for royalties, refining, transportation and insurance costs.

Cash cost calculations use the total operating cost plus royalties divided by the payable gold ounces.

Total costs are calculated using the total costs (operating and capital) plus royalties and taxes divided by

the payable gold ounces.

The cash-flow model calculates the Net Present Value (NPV) based on a discounted rate of 0%


on total investment and the payback period were also calculated. A sensitivity analysis was also

conducted on parameters that are deemed to have the biggest impact on the Project financial

performance (capital cost, operating cost and gold selling price).

The economic analysis demonstrates robust economics and confirms the overall viability of the project.

25.8 Environmental Studies, Permitting, and Social and Community Impact

All the environmental impacts have been identified and mitigation measures developed. An Environmental

Impact Assessment (EIA) was prepared for the Project and submitted to the state environmental

regulatory authority.

Red Eagle Mining has a number of sustainable programs planned that will positively develop the social

environment. This will be achieved by the existing commitment to maximize the recruitment of local

personnel during both construction and operations. It is estimated that at least 150 people will benefit from

direct employment and at least another 500 will benefit indirectly. Local suppliers of goods and support

services to the mine will be able to develop their businesses through increased opportunities. Education

and training programs will raise the skills level of the workers in the area. Culturally the local communities

are willing to accept change and have a high potential and desire to improve and benefit themselves.

The project footprint lies within an area owned by a single landowner and therefore there is limited effect

on land or property tenure.

There are no impacts related to displacement of the local population and any impacts on their way of life

are minimal and management plans are in place to mitigate these. The remote location of the project in

relation to the local communities and population assists in minimizing any potential impacts.

Page 25.5




AMENDED NI 43-101 TECHNICAL REPORT 25.9 Risks and Opportunities

The following overview provides details of the risk management process that took place during the Santa

Rosa Gold Project Feasibility Study, in order to identify risks which may be encountered during the Project

execution phase.

In general, the risk management process helps ensure that a project is designed, constructed,

commissioned, and operated with an acceptable level of risk in relation to its financial viability, to the

environment, to the community, and to the health and safety of personnel and plant integrity.

The risk management workshop completed during this Feasibility Study addressed the risks associated

with the various future project activities, including:

Engineering and studies;

Permitting;

Procurement; and

Construction and Commissioning.

The following list will discuss some of the risks identified during the risk management workshops as

potentially severe or high after mitigation.

Concentrate regrind mill may be slightly undersized. Work is currently underway to

optimize both the regrind size and the mass pull in flotation in order to specify the

regrind more accurately. Although the regrind mill is designed to produce a P80 of 25

microns, indications from both the testwork and mineralogical studies suggest that only

a “light” regrind is required. Furthermore, the flotation mass pull is high compared to the

total sulfide content of the ore and optimizing this stage is expected to significantly

reduce the tonnage of concentrate requiring regrinding;

Delays in construction of the 44 kV power line from the EPM substation to the

switchyard at the project site. The impact would be delayed commissioning of the

process plant resulting in loss of revenue in the first year of production. This would be a

design, build contract with EPM. To mitigate this risk the Project should incorporate as

much float as possible into the power line construction schedule so that it would not be

on the critical path of the project construction and commissioning. Another mitigation

measure would be to secure the supply of rental diesel generators to provide power to

the plant in case the power line is not complete. Several local vendors have been

identified capable of supplying the required size rental generators to satisfy the power

needs of the project;

Page 25.6




AMENDED NI 43-101 TECHNICAL REPORT There is a minor risk that the rock quality encountered in the initial decline development

may be worse than anticipated, and if a failure occurs, it will cause excessive delays and

costs to remedy. This has been mitigated by drilling geotechnical holes into the saprolite

and analysing the rock quality. It is planned to install weep holes into the saprolite to

relieve any presence of water saturation. Additionally, steel arch sets, rockbolts and

reinforced shotcrete have been allowed for in the upper section of the decline;

The narrow widths of the ore zones have the potential to attract additional dilution if not

mined carefully. This has been mitigated by the allowance of wall dilution, or selvedge of

0.25 m on either side of the stope outlines. This increased the minimum mining width to

2.5 m. The mining method (MSDF) also has the advantage of managing wall overbreak

closely. However, there is still risk of wall sloughing during and after mucking out and

backfill placement;

MSDF is a relatively uncommon mining method, and will be undertaken with a

Colombian contractor who is inexperienced in ore production mining techniques. This

will be mitigated by the introduction of an expatriate specialised mining training team for

the first year of production. Additionally, and most importantly, the mining method can be

easily converted to conventional cut-and-fill techniques, which minimises large wall

exposures;

There is potential for a higher degree of sulphide oxidation resulting in possible acid

generation in the DWMF underdrainage system. This is mitigated by the blending of

mine development waste rock (granodiorite), which has relatively strong acid consuming

characteristics. The mine waste rock will be added to the filtered tailings in ratio of

approximately 1 to 2;

There is always the risk of security in Colombia, although the situation is constantly

improving over time. The Santa Rosa Gold Project has never had an incident of security

breach. This is mitigated by the fact that Red Eagle Mining has introduced the services

of the Colombian national army under contract, as a precaution, to patrol the project site

and surroundings around the clock. This service has been in operation for 2 years to

date and will continue for the life of the mine; and

Mineralization at the 2 g Au/t, the economic cut-off grade, is continuous over distances

of hundreds of meters within the San Ramon deposit. While MDA has demonstrated that

the overall quantity of gold in the ore reserves is as stated, there is a lack of

demonstrated continuity in the very high-grade sub-domain (>5 g Au/t) at the current

50m spacing. Mitigation will require detailed delineation drilling to be carried out in

advance of design and development of the final stoping blocks in readiness for mining.

This discontinuity of the highest-grade zones may also be a problem in maintaining

steady feed grades to the process plant. This has been mitigated by the MSDF mining

method, whereby draw of ore from a number of stopes will provide grade control

flexibility.

Page 25.7




AMENDED NI 43-101 TECHNICAL REPORT The following opportunities have been identified:

Current testwork shows significantly lower cyanide consumption compared to that

established from earlier testwork. If these results are confirmed by further testwork it will

result in a reduction in operating costs;

Testwork results indicate that a coarser grind size (150 microns) will not impact gold

recoveries. If confirmed with additional testwork then a coarser grind size will reduce the

overall grinding power requirement thus reducing operating costs. The mill size could

also be reduced resulting in capital cost reduction. This will also assist in the tailings

filtration process;

The inclusion of a tails thickener will allow for recovery of cyanide thus lowering the

overall cyanide consumption. This will reduce the operating cost but its impact on the

capital costs needs to be evaluated;

Testwork results indicate that a reduction in the leach time will not impact gold

extraction. If confirmed with additional testwork then a shorter leach time will reduce the

size of the leach train thus resulting in capital and operating cost reductions;

The San Ramon deposit appears to be plunging to the east into a tenement recently

acquired from AngloGold Ashanti. This area has never been drill tested. Red Eagle

Mining now owns property extending for over 2 km further to the east of the present

extent of the deposit. The known deposit is also open-ended down-dip;

If conditions underground indicate continuous ore zones of constant widths and dip, then

the method of sub-level retreat long haul stoping could be adopted in these areas. This

method would reduce costs and increase productivity significantly;

Because the continuity of mineralization exceeding 5 g Au/t is less than the current 50 m

drill spacing, there is potential to encounter additional very high-grade quartz / sulfide

zones during infill drilling from underground and during mining that were not located by

drilling from the surface. However, this potential cannot be quantified and has not been

included in the current resource; and

Sample integrity studies show that gold grade decreases significantly at core recoveries

below 70%, which may be due to washing out of gold from softer rocks during drilling.

Evaluation of core recoveries for the various lithologic, weathering, and oxide types

indicates that recovery in saprolite / oxidized rock is much lower than in unweathered

and unoxidized rocks. Additionally, free gold or gold associated with sulfide minerals in

soft, clayey gouge in the shear zone could be washed out, resulting in poor recovery

and preferential loss of gold in the sample. Assays of samples of such rock may,

therefore, be understated where core recoveries are below 70%.

Page 26.1




AMENDED NI 43-101 TECHNICAL REPORT 26.0 RECOMMENDATIONS

26.1 Geology and Mineral Resources

This Feasibility Study demonstrates that the San Ramon deposit is a project of merit that warrants

continued exploration to expand the resource. Once the shear zone mineralization is accessed by the

exploration decline, testing can be conducted on the mineral deposit in situ. The decline will also be

available for infill-drilling recommended to further delineate the deposit, and for exploration for new

mineralization down-dip and to the east of the current known resource.

26.2 Mining and Mineral Reserves

Additional study for mining should be undertaken and include:

The Feasibility Study assumes mining contractors will be used to offset mining capital.

Proposals were received as part of the mining study, and several discussions have been

underway with multiple mining contractors. These negotiations need to be completed

with the goal of developing a working partnership with the contractor that will promote

productivity and safety while maintaining costs at a minimum;

Mining contractors in the area are very experienced in underground construction and

development, but are less experienced in mine production. Red Eagle Mining has

decided to bring in a team of experienced mining experts to help train the contract

miners. To make this strategy successful these mining experts need to be sourced;

Management of dilution and ore loss will be the key to the success of the underground

operations. This will require planning and implementation of a delineation program and

the procedures to be used. Red Eagle Mining should determine the procedures and

work flow required to ensure a successful program of delineation for stopes. This

includes how drill planning is done, how drilling results are tracked, establishment of

quality control protocols and procedures, modelling of results, inclusion into detailed

stope designs, and reconciliation of planned and actual mining; and

Explosives are supplied by the Colombian Government and are long lead orders.

Negotiations with contractors need to be completed and orders placed to ensure that

explosive products are available when required.

Page 26.2




AMENDED NI 43-101 TECHNICAL REPORT 26.3 Metallurgical Testing and Recovery Methods

To compliment the results of the Feasibility Study, additional metallurgical testwork is recommended as

follows:

Bulk flotation testwork to generate 10 to 20 kg of flotation concentrate using 100 to 200

kg sample. It is recommended that the sample is extracted from the existing Hilo Azul

underground cross cut, which intersects the full width of the shear zone in sulphides;

Regrind testwork should be undertaken by the selected equipment vendor using the

flotation concentrate. This testwork will confirm the size of the regrind equipment, refine

the operating cost estimate and will facilitate process guarantees upon purchase of the

equipment;

The bulk flotation tail and reground product should be collected after testing. The

samples should be sent for vendor thickening and filtration testwork to facilitate process

guarantees upon purchase of the equipment; and

Given the lower cyanide consumption and lower final cyanide concentration in the coast

down tests, further testing is recommended to optimize the cyanide concentration and

dosage schedule during the leach step.

In addition to metallurgical testwork the following modelling and value engineering assessments are

recommended:

Value engineering assessment of the inclusion of a tail thickener to recover cyanide

should be undertaken. The current design minimises capital rather than operating cost;

and

Given the closed circuit nature of the flowsheet it is recommended that a more detailed

mass and chemical modelling of the circuit be undertaken including both the plant and

site water system to better define the chemistry of circulating loads and the final effluent.


Due to the reduction in the project site footprint and process plant relocation late in the Feasibility Study,

an extension to the completed geotechnical program needs to be undertaken for the equipment and

building foundations in the current location of the plant site to supplement the drilling done as part of this

Feasibility Study. The program must be completed prior to the start of detailed engineering.

Detailed placement plans will be required to match the quantities of tailings and mine rock as per the mine

plan with the stability, water management and erosion requirements of the DWMF.

Page 26.3




AMENDED NI 43-101 TECHNICAL REPORT 26.5 Environmental Studies, Permitting, and Social and Community Impact

Red Eagle Mining needs to obtain the necessary Environmental License in order to start construction

activities at the Project site. The social program of information and workshops that have been running for

over two years should continue through the duration of operations. The corporate policy of Red Eagle

Mining is to remain transparent to all communities and stakeholders in the region.

Red Eagle Mining must involve residents of the local communities in its ongoing environmental monitoring

programs, and also in all operational activities.

Red Eagle Mining is committed to continue the existing and planned future community education and

training programs.

26.6 Main Recommendation

The Feasibility Study for the Santa Rosa Gold Project has been completed in sufficient detail to refine the

economics to a +/-15% level of accuracy and outline the issues facing the project going forward. The

Project economics are sufficiently robust to warrant moving to the next phase of detailed engineering and

construction. The estimated cost for the entire mine development and process plant and infrastructure

construction up to, and including start-up and commissioning is 69.90 million (excluding VAT to be paid,

but reclaimed in the first year of production).

Table 26.1 Recommended Additional Work Cost Estimate

Additional Work Cost Estimate

Mining Contract Negotiations $15,000

Bulk Flotation Testwork $35,000

Regrind and Thickening Testwork No Cost (By Vendors)

Reagent Consumption Testwork (Cyanide) $20,000

Metallurgical Consultant for Testwork $25,000

Value Engineering $40,000

Geotechnical Drilling and Testwork $35,000

Mass and Chemical Circuit Modelling $35,000

TOTAL: $205,000

Page 27.1




AMENDED NI 43-101 TECHNICAL REPORT 27.0 REFERENCES

Alvarez, O. A. G., (undated), Portafolio de Presentacion – Proyecto Yaruma San Ramon. Santa Rosa de

Osos (Antioquia), Republica de Colombia, Sur America.

Aspden, J.A. and Litherland, M., (1992), The geology and Mesozoic collisional history of the Cordillera

Real, Ecuador: Tectonophysics, v. 205, p. 187-204.

Cediel, F., and Caceres, C., 2000, Geological map of Colombia, 1: 2,000,000: Geotec Ltd., Bogotá, Third

Edition, digital format with legend and tectono-stratigraphic chart.

Cediel, F., Shaw, R.P. and Caceres, C., 2003, Tectonic assembly of the Northern Andean Block, in

Bartolini, C., R.T. Buffler, and J. Blickwede, eds., The Circum Gulf of Mexico and Caribbean: Hydrocarbon

habitats, basin formation, and plate tectonics: AAPG Memoir 79, p. 815 - 848.

Cobos, D., Espinosa, R., and Eldridge, T., May 2014, Memorandum: Rock mass characterization for

underground mine development.

Conestoga-Rovers Colombia SAS (CRA), 2013, Estudio de Impacto Ambiental para el Proyecto Minero

Santa Rosa.

Dyer, T. L., Lindholm, M. S., Schlitt, W. J., and Defilippi, C., 2014 (amended March 31; original October

10, 2013), Amended technical report and preliminary economic assessment, San Ramon deposit, Santa

Rosa project, Colombia: Report prepared for Red Eagle Mining Corporation by Mine Development

Associates, 207 p.

Feininger, T., and Botero-Arango, G., 1982, The Antioquian batholith, Colombia: Publicaciones Especiales

del INGEOMINAS, no. 12, Bogota, 50 p.

GeoLogic Associates, 2014, Lab Report Project #3848, September.

Golder Associates, 2013, Red Eagle Mining Proyecto Santa Rosa – Balance Hídrico Preliminar,

December.

Golder Associates, 2014a, Field report.

Golder Associates, 2014b, Santa Rosa Project Tailings Management Facility Design Report.

Gorham, J., 2007 (October 22), Summary report on the Gramalote property, Department of Antioquia,

Colombia: NI43-101 Technical Report prepared for B2Gold Corp.

Jemielita, R. A., 2011a (March 31), Canadian NI43-101 Technical Report; an assessment of the Santa

Rosa gold project, Department of Antioquia, Colombia: Report prepared for Red Eagle Mining

Corporation, 100 p.

Jemielita, R. A., 2011b (June 12), Canadian NI43-101 Technical Report (amended); an assessment of the

Santa Rosa gold project, Department of Antioquia, Colombia: Report prepared for Red Eagle Mining

Corporation, 148 p.

Kwok, D., and Choi, T., 2012 (September 7), Metallurgical tests for Red Eagle Mining, Santa Rosa project

in Colombia, Project No. 12015: Report prepared for Red Eagle Mining Corporation by AcmeMet, 28 p.

plus appendices.

Page 27.2





Lindholm, M. S., and Schlitt, W. J., 2013a (January 22), Technical report on the San Ramon deposit,

Santa Rosa project, Colombia: Report prepared for Red Eagle Mining Corporation by Mine Development

Associates, 86 p.

Lindholm, M. S., and Schlitt, W. J., 2013b (September 10), Updated technical report on the San Ramon

deposit, Santa Rosa project, Colombia: Report prepared for Red Eagle Mining Corporation by Mine

Development Associates, 120 p.

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mineros del Ecuador, v. 1: British Geological Survey and Direccion Nacional de Geologia, Proyecto de

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SGS Canada, July 2013, Modified ABA Certificate of Analysis, Project ES 29857-1.

Simard, J., 2011 (August), Report on an induced polarization survey performed on the Santa Rosa project,

Antioquia Department, Colombia: Report prepared for Red Eagle Mining Corporation by Geofisica TMC

SA de CV, 18 p.

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Report prepared for Red Eagle Mining Corporation by Paterson, Grant & Watson Limited, 12 p.

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101 Technical Report: Prepared for Medoro Resources, Gran Colombia Gold, S.A., Tapestry Resources

Corp.

CERTIFICATE OF QUALIFICATION

I, Stefan Gueorguiev, P.Eng. do hereby certify that:

1. I am currently employed as Manager of Projects/ Senior Project Manager in the consulting firm

Lycopodium Minerals Canada Ltd. located at 5060 Spectrum Way, Mississauga, Ontario,

Canada, L4W 5N5.

2. This certificate accompanies the report, dated October 27, 2014, and titled “Amended NI 43 101

Technical Report Feasibility Study of the Santa Rosa Gold Project”.

3. I graduated with a Bachelor of Science degree in Civil/ Structural Engineering from the

University of Architecture and Civil Engineering, Sofia, Bulgaria in 1991 and a Master of

Project Management degree from the Pennsylvania State University, Behrend College, Erie

Pennsylvania in 2010.

4. I am in good standing as a member of the Professional Engineers of Ontario (#90417163), the

Association of Professional Engineers and Geoscientists of Alberta (#92555), and the

Association of Professional Engineers and Geoscientists of Saskatchewan (#31311).

5. I have practiced my profession continuously since my graduation. My relevant experience

includes consulting and managing mining projects in various stages of the execution cycle from

scoping, pre-feasibility and feasibility studies through detailed design, construction, and

commissioning. I have been involved in gold, iron ore, diamond, and tar sands mining projects in

North America, South America, Europe, and West Africa.

6. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-

101”) and certify that by reason of my education, affiliation with a professional association (as

defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a

“qualified person” for the purposes of NI 43-101.

7. I am independent of the Issuer and related companies applying all of the tests in section 1.5 of

National Instrument 43-101.

8. I am one of the authors of this Technical Report titled “Amended NI 43-101 Technical Report

Feasibility Study of the Santa Rosa Gold Project” prepared for Red Eagle Mining Corporation,

effective as of October 6, 2014. I am responsible for Sections 2, 3, 4, 5, 17, 18, 19, 20, 22, 23,

24, portions of Section 21 in relation to process plant capital and operating costs, and relevant

portions of Sections 1, 25, and 26 relating to processing.

9. I visited the Santa Rosa Project site on February 19, 2014.

10. As of the date of the certificate, to the best of my knowledge, information and belief, the

Technical Report contains the necessary scientific and technical information to make the

Technical Report not misleading.

11. I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been

prepared in compliance with that instrument and form.

Dated October 27

th, 2014.

“Stefan Gueorguiev"

Stefan Gueorguiev, P.Eng.

Hydrometal, Inc.

CERTIFICATE of QUALIFIED PERSON – William Joseph Schlitt

I, W. Joseph Schlitt, do hereby certify that:

I am President of Hydrometal, Inc., P.O. Box 309, Knightsen, CA, 94548, U.S.A.

I graduated with a Bachelor of Science degree in Metallurgical Engineering (with highest

honors) from Carnegie Institute of Technology, Pittsburgh, PA in 1964 and with a Doctor of

Philosophy degree in Metallurgy (with honors) from The Pennsylvania State University,

University Park, PA in 1968.

I am a Qualified Professional member of the Mining and Metallurgical Society of America

No. 01003QP with specialty in metallurgy and a Registered Professional Metallurgical

Engineer (Texas No. 53603).

I have practiced continuously as a Metallurgist for more than 45 years since my graduation

from university. The bulk of this work has focused on the extraction of base metals and

gold/silver from ores and concentrates. Some of the extraction studies involved milling and

flotation of sulfidic ores, where the sulfides were the carriers for such metals as copper, gold

and silver. The initial concentrates were either leach directly or further upgraded for sale to

a third party. Other extraction studies involved direct leaching of the mined ores or, in some

cases, sub-mill grade mine wastes.

I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-

101”) and certify that by reason of my education, affiliation with a professional association

(as defined in NI 43-101), and past relevant work experience, I fulfill the requirements to be

a “qualified person” for the purposes of NI 43-101.

I am responsible for the preparation of Section 13.0 Mineral Processing and Metallurgical

Testing, plus subsection 1.7 Mineral Processing and Metallurgical Testing in the Technical

Report entitled “Amended NI 43-101 Technical Report Feasibility Study of the Santa Rosa

Gold Project” and dated October 27, 2014 (the “Technical Report”). I visited the Santa

Rosa property on February 19-22, 2013. On several occasions I have visited the major

metallurgical laboratories where testwork has been performed on the various samples.

I was co-author of the previous Technical Report on the San Ramon deposit dated January

22, 2013 entitled “Technical Report on the San Ramon Deposit, Santa Rosa Project,

Colombia.” I was also the co-author of the previous Technical Report dated September 10,

2013 entitled “Updated Technical Report on the San Ramon Deposit, Santa Rosa Project,

Colombia” and the previous Technical Report dated March 31, 2014 entitled “Amended

Technical Report and Preliminary Economic Assessment, San Ramon Deposit, Santa Rosa

Project, Colombia”. Prior to the work done for the previous Technical Reports, I have not

been involved with Red Eagle Mining for the property that is the subject of this Technical

Report.

As of the effective date of this Technical Report, to the best of my knowledge, information,

and belief, those parts of the Technical Report for which I am responsible contain all

scientific and technical information that is required to be disclosed to make the Technical

Report not misleading.

I am independent of the issuer applying all of the tests in section 1.5 of National Instrument

43-101.

Hydrometal, Inc.

I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has

been prepared in compliance with that instrument and form.

Dated this 27th Day of October 2014

/s/ W. Joseph Schlitt___

(Signed and sealed)

W. Joseph Schlitt, President

Hydrometal, Inc.

CERTIFICATE OF QUALIFIED PERSON

I, Terry Eldridge, P.Eng. of Golder Associates, 500-4260 Still Creek Drive, Burnaby, British Columbia, hereby certify that:

1. I am a consulting geotechnical engineer, employed full time by Golder Associates, at the

address set out above.

2. I am a graduate of the University of British Columbia in civil engineering, with the degree

B.A.Sc granted in 1980. I also hold the degree of M.A.Sc in geotechnical engineering granted

by the University of British Columbia in 1983.

3. I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia, Registration Number 14894.

4. I have worked as a Professional Engineer in the field of mine development since 1985, in many parts of the world, including Canada, United States, Colombia, Peru, Chile, Argentina, Brazil, Spain, Russia, Indonesia, Uruguay, Kyrgyzstan, Kazakhstan and Australia. I have worked on dry stack tailings projects since 1994 and in the past 3 years have been involved in the design of four dry stack tailings facilities for gold projects in South America. I am the South American Mine Waste Management Practice Lead for Golder Associates.

5. I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI 43-

101") and certify that by reason of my education, affiliation with a professional association (as

defined in NI 43101) and past relevant work experience, I fulfill the requirements to be a

"qualified person" for the purposes of NI 43-101.

6. I am responsible for the preparation of Sections 16.4 and 18.12 of the technical report entitled

"Amended NI 43-101 Technical Report Feasibility Study of the Santa Rosa Gold Project”,

dated October 27, 2014 relating to the Santa Rosa property. I visited the Santa Rosa property

February 20 to 22, 2013.

7. Prior to 2013 I had no involvement with the property that is the subject of the Technical Report.

8. As of the date hereof, to the best of my knowledge, information and belief, the Sections of the

Technical Report set out above contain all scientific and technical information that is required

to be disclosed to make the Technical Report not misleading.

9. I am independent of the issuer as described in Section 1.5 of National Instrument 43-101. I may inadvertently be the beneficial owner of an interest in any publicly traded company through participation in mutual funds over whose portfolios I have no control.

10. I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has


11. I consent to the filing of the Technical Report with any stock exchange and other regulatory

authority and any publication by them for regulatory purposes, including electronic publication

in the public company files on their websites accessible by the public, of the Technical Report.

Dated this 27th day of October 2014.

Terry L. Eldridge Terry L. Eldridge, P.Eng.

Principal, Golder Associates

CERTIFICATES OF QUALIFIED PERSONS

I, Thomas L. Dyer, P.E., do hereby certify that I am currently employed as Senior Enginer by Mine

Development Associates, Inc., 210 South Rock Blvd., Reno, Nevada 89502 and:

1. I am responsible for Sections 1.5, 1.6, 1.10.1, 1.11.1, 1.13.2, 15, 16, 21.1, 21.2, 25.2, and 26.2,

and jointly responsible for section 25.9 of this technical report, titled Amended NI 43-101

Technical Report - Feasibility Study of the Santa Rosa Gold Project, prepared by Lycopodium

Minerals Canada Ltd., for Red Eagle Mining Corporation, effective as of October 6 and dated

October 27, 2014 (“Technical Report”).

2. I graduated with a Bachelor of Science degree in Mine Engineering from South Dakota School

of Mines and Technology in 1996. I have worked as a mining engineer for a total of 18 years

since my graduation. During my Engineering career I have held various positions of

increasing responsibility at operating mines performing life of mine planning and cost

estimates. During the last 7 years I have been engaged in consulting on various lead, zinc,

gold, silver, copper, and limestone deposits both for underground and open pit operations.

This consulting work has primarily consisted of providing production schedules, mine cost

estimates, and cash-flow analysis. I am a Registered Professional Engineer in the state of

Nevada (#15729) and a Registered Member (#4029995RM) of the Society of Mining,

Metallurgy and Exploration.



defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a


4. I visited the property on September 25 through September 29, 2012 and again from February

19 to 22, 2013. I am independent of Red Eagle Mining Corporation and its subsidiaries,

applying all of the tests in section 1.5 of National Instrument 43-101. I have had prior

involvement with the Santa Rosa gold project as a co-author and Qualified Person responsible

for portions of the Technical Report and Preliminary Economic Assessment, San Ramon

Deposit, Santa Rosa Project, Colombia for Red Eagle Mining Corporation effective as of

October 1, 2013, and dated March 31, 2014 (“Technical Report”).



6. As of the effective date of this Technical Report, to the best of my knowledge, information,

and belief, those parts of the Technical Report for which I am responsible contain all scientific

and technical information that is required to be disclosed to make the Technical Report not

misleading.

Dated this 27th day of October, 2014.

“Thomas L. Dyer”

Thomas L. Dyer, P.E.

Print Name of Qualified Person

CERTIFICATES OF QUALIFIED PERSONS

I, Michael S. Lindholm, C.P.G., do hereby certify that I am currently employed as Project Geologist by

Mine Development Associates, Inc., 210 South Rock Blvd., Reno, Nevada 89502, and:

1. I am responsible for Sections 1.2 through 1.4, Sections 6 through 12, 14, 25.1, and 26.1, and

jointly responsible for Section 25.9 of this Technical Report, titled Amended NI 43-101

Technical Report - Feasibility Study of the Santa Rosa Gold Project, prepared by Lycopodium

Minerals Canada Ltd., for Red Eagle Mining Corporation, effective as of October 6 and dated

October 27, 2014 (“Technical Report”).

2. I graduated with a Bachelor of Science degree in Geology from Stephen F. Austin State

University in 1984 and with a Master of Science degree in Geology from Northern Arizona

University in 1989. I am a Professional Geologist in the state of California (#8152) and a

Certified Professional Geologist (#11477) with the American Institute of Professional

Geologists. I have worked as a geologist for a total of 29 years since graduation from

undergraduate university. During my career as a project geologist in the mining industry, I

have held various positions of increasing responsibility in the exploration, development and

mining of precious and base metal deposits. During the last 5 years I have been engaged in

consulting on a range of deposit types in varied geologic environments, primarily in North and

South America, including shear-hosted gold deposits similar to the Santa Rosa Gold project

The consulting work has consisted of independent verification of project data, geologic and

mineral domain modeling, and resource estimation.



defined in NI 43-101) and past relevant work experience, I fulfill the requirements of


4. I visited the Santa Rosa property most recently on May 14 through 17, 2013. I am

independent of Red Eagle Mining Corporation and all their subsidiaries as defined in Section

1.5 of NI 43-101 and in Section 1.5 of the Companion Policy to NI 43-101. I had no

involvement with the project prior to preparing and serving as a co-author and Qualified

Person responsible for previous Technical Reports dated January 2013, September 2013 and

March 2014 on the Santa Rosa project.



6. As of the effective date of this Technical Report, to the best of my knowledge, information,

and belief, those parts of the Technical Report for which I am responsible contain all scientific

and technical information that is required to be disclosed to make the Technical Report not

misleading.

Dated this 27th day of October, 2014.

“Michael S. Lindholm” Signature of Qualified Person

Michael S. Lindholm, C.P.G.

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