Technical Report: Blue River Tantalum and Niobium Project (Preliminary Economic Assessment)

Blue River Tantalum–Niobium ProjectBritish Columbia, CanadaNI 43-101 Technical Report on Preliminary Economic Assessment

Prepared for:Commerce Resources Corporation

Prepared by:Albert Chong, P.Geo.Tomasz Postolski, P.Eng.Ramon Mendoza Reyes, P.Eng.Tony Lipiec, P.Eng.Behrang Omidvar, P.Eng.Effective Date: 29 September 2011

Project No. 162230

CERTIFICATE OF QUALIFIED PERSON

Albert Chong, P.Geo. AMEC Americas Limited

111 Dunsmuir Street, Suite 400 Vancouver, B.C. V6B 5W3

Phone: (604) 664-4116 E-mail: [email protected]

I, Albert Chong, P.Geo., am employed as a Principal Geologist with AMEC Americas Limited.

This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of 29 September, 2011 (the “Technical Report”).

I am a Professional Geoscientist in the Province of British Columbia (P.Geo. #23773). I graduated from McMaster University, Hamilton, Ontario with a B.Sc. degree in Geology, and from the University of Tasmania with a M.Sc. degree in Exploration Geoscience.

I have practiced my profession for 26 years since graduation. I have been directly involved in green fields and brown fields exploration, mining operations, consulting, and resource estimation of base metal, precious metal and rare metal deposits.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I visited the Blue River property from July 11 to 16, 2010, June 27 to 30, 2011, and September 6 to 14, 2011.

I am responsible for Sections 1 to 12, 20, and 23 to 27 of the Technical Report.

I am independent of Commerce Resources Corporation as independence is described by Section 1.5 of NI 43–101.

I have been involved as an independent consultant on the Blue River Ta-Nb Project since 2010.

I have read NI 43–101 and this report has been prepared in compliance with that Instrument.

As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

“signed and sealed”

Albert Chong, P.Geo.

Dated: 31 October, 2011


Tomasz Postolski, P.Eng. AMEC Americas Limited



I, Tomasz Postolski, P.Eng., am employed as a Senior Geostatistician with AMEC Americas Limited.

This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of 29 September, 2011 (the “Technical Report”).

I am a Professional Engineer in the Province of British Columbia (P.Eng. #34784). I have graduated from The University of Mining and Metallurgy, Krakow, Poland with a Magister Inzynier degree in Geological Engineering, and from the University of British Columbia with a Master of Applied Science degree also in Geological Engineering. I have completed the Citation Program in Applied Geostatistics at the Centre for Computational Geostatistics at the University of Alberta.

I have 17 years of consulting, mine operations, and academic experience specializing in geostatistical mineral resource estimation and geological evaluation of gold, copper, rare earth metals and other mineral deposits in Canada and abroad.


I visited the Blue River property June 27 to 30, 2011.

I am responsible for Section 14 and those portions of the Summary, Interpretation and Conclusions, and Recommendations that pertain to this Section of the Technical Report.


I have been involved with mineral resource estimation on the Blue River Ta-Nb Project since 2010.




Tomasz Postolski, P.Eng.

Dated: 31 October 2011


Ramon Mendoza Reyes (P.Eng.) Suite 400, 111 Dunsmuir Street

Vancouver, B.C., Canada Tel: +1 (604) 664-3075; Fax: +1 (6040 664-3057

email: [email protected]

I, Ramon Mendoza Reyes (P.Eng.) am employed as a Principal Mining Engineer with AMEC Americas Limited. This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of 29 September, 2011 (the “Technical Report”).

I am a Professional Engineer in the province of British Columbia. I graduated in 1989 from the National Autonomous University of Mexico with a bachelor’s degree in Mining Engineering, and in 2003 completed a M.Sc. Degree in Mining & Earth Systems Engineering from the Colorado School of Mines in Golden, Colorado, USA. I have practiced my profession for 22 years, and have previously been involved with mine designs, mine planning and mine operations for base metal, disseminated sulphide and industrial mineral projects in North America and South America.


I visited the Blue River property in British Columbia between the 12 and the 14 of July, 2010.

I am responsible for Sections 15, 16, and 18 and those portions of the Summary, Cost Estimates Interpretation and Conclusions, and Recommendations that pertain to the mining sections of the Technical Report.


I have been involved with the mining aspects of the Blue River Tantalum–Niobium Project since January 2010.




Ramon Mendoza Reyes, P.Eng.



Ignacy (Tony) Lipiec (P.Eng.) Suite 400, 111 Dunsmuir St

Vancouver, BC., Canada Tel: 604-664-3130; Fax: 604-664-3057

Email: [email protected]

I, Ignacy (Tony) Lipiec (P.Eng.) am employed as a Principal Metallurgical Engineer with AMEC Americas Limited. This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date of 29 September, 2011 (the “Technical Report”).

I am a Professional Engineer in the province of British Columbia. I graduated from the University of British Columbia with a B.A.Sc. degree in Mining & Mineral Process Engineering, in 1985. I have practiced my profession for 25 years, and have previously been involved with metallurgical design and process engineering for base metal and disseminated sulphide projects in North America and South America.


I did not visit the Blue River property.

I am responsible for Sections 13, 17, 18, 21 and those portions of the Summary, Interpretation and Conclusions and Recommendations that pertain to those sections of the Technical Report.


I have been involved as an independent consultant with the Blue River Ta-Nb Project since 2010.




Tony Lipiec, P.Eng.



Behrang Omidvar, P.Eng. AMEC Americas Limited



I, Behrang Omidvar, P.Eng., am employed as a Financial Analyst with AMEC Americas Limited.

This certificate applies to the Technical Report titled “Blue River Tantalum–Niobium Project, British Columbia, Canada, NI 43-101 Technical Report on Preliminary Economic Assessment” with an effective date 29 September, 2011 (the “Technical Report”).

I am a Professional Engineer in the Province of British Columbia (P.Eng. #35500). I have graduated from The University of British Columbia with a Mechanical Engineering degree.

I have seven years of experience in engineering, project management and financial analysis for mining and other projects. I have prepared cash-flow models and conducted financial and throughput analyses of numerous mines and development properties in Canada and internationally.


I have not visited the Blue River property.

I am responsible for Sections 19, 21, 22 and those portions of the Summary, Interpretation and Conclusions and Recommendations that pertain to those Sections of the Technical Report.


I have been involved as an independent consultant with the Blue River Ta-Nb Project since 2010.




Behrang Omidvar, P.Eng.


IMPORTANT NOTICE

This report was prepared as a National Instrument 43-101 Technical Report for Commerce

Resources Corporation (Commerce) by AMEC Americas Limited (AMEC). The quality of

information, conclusions, and estimates contained herein is consistent with the level of

effort involved in AMEC’s services, based on: i) information available at the time of

preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and

qualifications set forth in this report. This report is intended for use by Commerce subject

to the terms and conditions of its contract with AMEC. Except for the purposes legislated

under Canadian provincial securities law, any other uses of this report by any third party is

at that party’s sole risk.

Commerce Resources Corp.

Blue River Tantalum–Niobium Project

British Columbia, Canada

NI 43-101 Technical Report on Preliminary Economic Assessment

Project No.: 162230

29 September 2011 TOC i

C O N T E N T S

1.0 SUMMARY ................................................................................................................... 1-1

1.1 Key Outcomes .......................................................................................................... 1-1

1.2 Location and Access ................................................................................................. 1-1

1.3 Mineral Tenure, Surface Rights, Royalties, and Agreements .................................... 1-1

1.4 Environment, Permitting and Socio-Economics ......................................................... 1-2

1.5 Geology and Mineralization ....................................................................................... 1-3

1.6 Exploration ................................................................................................................ 1-3

1.7 Exploration Potential ................................................................................................. 1-4

1.8 Drilling ....................................................................................................................... 1-4

1.9 Sample Analysis and Security ................................................................................... 1-5

1.10 Data Verification ....................................................................................................... 1-6

1.11 Metallurgical Testwork .............................................................................................. 1-6

1.12 Mineral Resource Estimate ....................................................................................... 1-7

1.13 Proposed Mine Plan .................................................................................................. 1-9

1.13.1 Geotechnical Considerations ........................................................................... 1-10

1.13.2 Dilution Considerations .................................................................................... 1-10

1.13.3 Drilling and Blasting ......................................................................................... 1-10

1.13.4 Mine Development ........................................................................................... 1-10

1.13.5 Mineralized Material and Waste Haulage ......................................................... 1-11

1.13.6 Mine Services .................................................................................................. 1-11

1.14 Mine Production Forecasts ..................................................................................... 1-12

1.15 Process Design ....................................................................................................... 1-12

1.16 Tailings and Waste Management ............................................................................ 1-15

1.17 Planned Project Infrastructure ................................................................................. 1-15

1.18 Markets ................................................................................................................... 1-16

1.19 Capital Costs .......................................................................................................... 1-17

1.20 Operating Costs ...................................................................................................... 1-17

1.21 Financial Analysis ................................................................................................... 1-18

1.22 Cash Costs ............................................................................................................. 1-19

1.23 Sensitivity Analysis ................................................................................................. 1-20

1.24 Interpretations and Conclusions .............................................................................. 1-21

1.25 Recommendations .................................................................................................. 1-23

2.0 INTRODUCTION .......................................................................................................... 2-1

2.1 Terms of Reference .................................................................................................. 2-1

2.2 Qualified Persons ..................................................................................................... 2-1

2.3 Site Visits and Scope of Personal Inspection ............................................................ 2-1

2.4 Effective Dates .......................................................................................................... 2-3

2.5 Information Sources and References ........................................................................ 2-3

2.6 Previous Technical Reports ...................................................................................... 2-4





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29 September 2011 TOC ii

3.0 RELIANCE ON OTHER EXPERTS .............................................................................. 3-1

3.1 Mineral Tenure .......................................................................................................... 3-1

3.2 Surface Rights .......................................................................................................... 3-1

3.3 Royalties and Agreements ........................................................................................ 3-1

3.4 Environmental, Permitting and Liability Issues .......................................................... 3-2

3.5 Markets ..................................................................................................................... 3-2

4.0 PROPERTY DESCRIPTION AND LOCATION ............................................................. 4-1

4.1 Project Ownership..................................................................................................... 4-1

4.2 Mineral Tenure .......................................................................................................... 4-1

4.3 Surface Rights .......................................................................................................... 4-4

4.4 Royalties and Agreements ........................................................................................ 4-4

4.5 Permits ..................................................................................................................... 4-4

4.6 Environment.............................................................................................................. 4-5

4.7 Social and Community Impact .................................................................................. 4-5

4.8 Comment on Section 4 ............................................................................................. 4-5

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ................................................................................................................... 5-1

5.1 Accessibility .............................................................................................................. 5-1

5.2 Climate ..................................................................................................................... 5-1

5.3 Local Resources and Infrastructure .......................................................................... 5-2

5.4 Physiography ............................................................................................................ 5-2


6.0 HISTORY ..................................................................................................................... 6-1

6.1 Commerce Exploration ............................................................................................. 6-1

6.2 Commerce Mineral Resource Estimates ................................................................... 6-1

7.0 GEOLOGICAL SETTING AND MINERALIZATION ...................................................... 7-1

7.1 Regional Geology ..................................................................................................... 7-1

7.2 Project Geology ........................................................................................................ 7-1

7.2.1 Metasedimentary Rocks .................................................................................... 7-3

7.2.2 Gneisses and Schists ........................................................................................ 7-3

7.2.3 Amphibolites ...................................................................................................... 7-3

7.2.4 Intrusive Rocks .................................................................................................. 7-5

7.2.5 Pegmatite Dykes .............................................................................................. 7-10

7.3 Structural Geology and Metamorphism ................................................................... 7-10

7.4 Geochronology ....................................................................................................... 7-11

7.5 Carbonatites ........................................................................................................... 7-12

7.5.1 Fir Carbonatite ................................................................................................. 7-12

7.5.2 Verity Carbonatite ............................................................................................ 7-13

7.5.3 Exploration Targets .......................................................................................... 7-13

7.6 Mineralogy .............................................................................................................. 7-14

7.6.1 Ferrocolumbite ................................................................................................. 7-16

7.6.2 Pyrochlore ....................................................................................................... 7-16

7.6.3 Fersmite ........................................................................................................... 7-16

7.6.4 Fenite Mineralization ........................................................................................ 7-17

7.6.5 Mineral Zoning ................................................................................................. 7-17

7.7 Comments on Section 7 .......................................................................................... 7-17





Project No.: 162230

29 September 2011 TOC iii

8.0 DEPOSIT TYPES ......................................................................................................... 8-1


9.0 EXPLORATION ............................................................................................................ 9-1

9.1 Grids and Surveys .................................................................................................... 9-1

9.2 Geological Mapping .................................................................................................. 9-1

9.3 Geochemical Sampling ............................................................................................. 9-1

9.3.1 Stream Sediment Sampling ............................................................................... 9-1

9.3.2 Soil Sampling ..................................................................................................... 9-2

9.3.3 Rock Chip, Grab, and Channel Sampling ........................................................... 9-2

9.4 Bulk Sampling ........................................................................................................... 9-3

9.5 Research Programs .................................................................................................. 9-4


10.0 DRILLING ................................................................................................................... 10-1

10.1 Core Drilling Strategy .............................................................................................. 10-2

10.1.1 Core Sizes ....................................................................................................... 10-2

10.1.2 Collar Surveys ................................................................................................. 10-2

10.1.3 Down Hole Surveys ......................................................................................... 10-4

10.1.4 Oriented Drill Core ........................................................................................... 10-4

10.1.5 Core Handling .................................................................................................. 10-4

10.1.6 Core Recovery ................................................................................................. 10-4

10.2 Drill Intercepts ......................................................................................................... 10-5

10.3 Comment on Section 10.......................................................................................... 10-5

11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY........................................... 11-1

11.1 Sampling Methods .................................................................................................. 11-1

11.2 Metallurgical Sampling ............................................................................................ 11-2

11.3 Density Determinations ........................................................................................... 11-2

11.4 Analytical and Test Laboratories ............................................................................. 11-3

11.5 Sample Preparation and Analysis ........................................................................... 11-4

11.6 Quality Assurance and Quality Control .................................................................... 11-4

11.6.1 Assessment of Accuracy with SRM Control Samples ....................................... 11-5

11.6.2 Assessment of Accuracy with Secondary Laboratory Pulp Checks .................. 11-5

11.6.3 Assessment of Precision with Duplicates ......................................................... 11-6

11.6.4 Assessment of Contamination Using Blanks .................................................... 11-7

11.7 Databases .............................................................................................................. 11-7

11.8 Security ................................................................................................................... 11-7


12.0 DATA VERIFICATION ................................................................................................ 12-1

12.1 Comments on Section 12 ........................................................................................ 12-1

13.0 MINERAL PROCESSING AND METALLURGICAL TESTING .................................... 13-1

13.1 Head Samples for Initial Testing ............................................................................. 13-2

13.2 Phase I Testing ....................................................................................................... 13-2

13.2.1 Grinding Size ................................................................................................... 13-2

13.2.2 Roughing and Cleaning Gravity Concentration ................................................ 13-3

13.3 Phase II Testing ...................................................................................................... 13-5

13.3.1 Flotation Tests ................................................................................................. 13-5

13.4 Review of Concentrate Treatment Options ............................................................. 13-8





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29 September 2011 TOC iv

13.5 Accuracy of Assaying .............................................................................................. 13-8


14.0 Mineral Resource Estimates ....................................................................................... 14-1

14.1 Geological Models .................................................................................................. 14-1

14.2 Exploratory Data Analysis ....................................................................................... 14-2

14.3 Density Assignment ................................................................................................ 14-2

14.4 Grade Capping/Outlier Restrictions ......................................................................... 14-3

14.5 Composites ............................................................................................................. 14-3

14.6 Variography ............................................................................................................ 14-3

14.7 Estimation/Interpolation Methods ............................................................................ 14-3

14.8 Block Model Validation ............................................................................................ 14-4

14.9 Classification of Mineral Resources ........................................................................ 14-4

14.10 Reasonable Prospects of Economic Extraction ................................................... 14-5

14.10.1 Tantalum Pricing .......................................................................................... 14-5

14.10.2 Niobium Pricing ............................................................................................ 14-6

14.10.3 Physical Assumptions .................................................................................. 14-6

14.10.4 Operational Considerations .......................................................................... 14-7

14.10.5 Economic Assumptions ................................................................................ 14-7

14.10.6 Economic Cut-off .......................................................................................... 14-7

14.11 Mineral Resource Statement ............................................................................... 14-8

14.12 Factors That May Affect the Mineral Resource Estimate ................................... 14-11

14.13 Comment on Section 14 .................................................................................... 14-11

15.0 Mineral Reserve Estimates ......................................................................................... 15-1

16.0 MINING METHODS.................................................................................................... 16-1

16.1 Optimization ............................................................................................................ 16-1

16.1.1 Assumptions .................................................................................................... 16-1

16.1.2 Mining Method ................................................................................................. 16-2

16.1.3 Production Rate ............................................................................................... 16-3

16.1.4 Backfill Option .................................................................................................. 16-3

16.2 Geotechnical Conditions ......................................................................................... 16-3

16.3 Conceptual Mining Method ..................................................................................... 16-4

16.4 Stoping Design ....................................................................................................... 16-5

16.4.1 Stability Analysis and Ground Support ............................................................. 16-5

16.4.2 Stope Geometry ............................................................................................... 16-7

16.4.3 Mining Sequence ............................................................................................. 16-7

16.4.4 Conceptual Mine Design .................................................................................. 16-7

16.4.5 Mining Dilution and Recovery .......................................................................... 16-7

16.5 Drilling and Blasting ................................................................................................ 16-8

16.6 Mine Development .................................................................................................. 16-8

16.7 Mineralized Material and Waste Rock Haulage ....................................................... 16-9

16.8 Mine Services ....................................................................................................... 16-12

16.9 Mine Development and Production Forecasts ....................................................... 16-12

16.10 Mine Equipment Requirements ......................................................................... 16-14

16.11 Mine Infrastructure ............................................................................................ 16-14

16.12 Mining Personnel ............................................................................................... 16-14






Project No.: 162230

29 September 2011 TOC v

17.0 RECOVERY METHODS ............................................................................................ 17-1

17.1 Plant Design ........................................................................................................... 17-1

17.2 Comminution (Crushing, Storage and Grinding) ...................................................... 17-1

17.3 De-sliming and Flotation ......................................................................................... 17-2

17.4 Filtration .................................................................................................................. 17-3

17.5 Concentrate Pre-Treatment .................................................................................... 17-4

17.6 Chlorination and Distillation ..................................................................................... 17-4

17.7 Product/Materials Handling ..................................................................................... 17-4

17.8 Energy, Water and Process Materials Requirements .............................................. 17-4


18.0 PROJECT INFRASTRUCTURE ................................................................................. 18-1

18.1 Site Layout .............................................................................................................. 18-1

18.2 Buildings ................................................................................................................. 18-1

18.2.1 Mine Service Building ...................................................................................... 18-1

18.2.2 Truck shop ....................................................................................................... 18-1

18.2.3 Warehouse ...................................................................................................... 18-1

18.2.4 Process Building .............................................................................................. 18-3

18.2.5 Crushing and Screening Circuit ....................................................................... 18-3

18.2.6 Portal Infrastructure ......................................................................................... 18-3

18.2.7 Explosives Storage .......................................................................................... 18-3

18.2.8 Aggregate Crushing and Concrete Batch Plants .............................................. 18-4

18.3 Roads and Logistics ................................................................................................ 18-4

18.3.1 Access Road.................................................................................................... 18-4

18.3.2 Haul Road ........................................................................................................ 18-4

18.4 Co-Disposal Storage Facilities ................................................................................ 18-5

18.4.1 Drystack Considerations .................................................................................. 18-5

18.4.2 Evaluation of Potential Sites ............................................................................ 18-6

18.4.3 Site Selection ................................................................................................... 18-7

18.4.4 Facility Design ................................................................................................. 18-7

18.4.5 Co-Disposal Facility Geohazards Considerations ............................................ 18-8

18.4.6 Co-Disposal Facility Stability Considerations ................................................... 18-8

18.4.7 Co-Disposal Facility Surface Water Run-off Considerations ............................. 18-9

18.4.8 Co-Disposal Facility Closure Considerations.................................................. 18-10

18.5 Avalanche Hazard................................................................................................. 18-11

18.6 Water Supply, Distribution, and Treatment Systems ............................................. 18-11

18.7 Waste Considerations ........................................................................................... 18-11

18.8 Accommodation .................................................................................................... 18-12

18.9 Power and Electrical ............................................................................................. 18-12

18.10 Fuel ................................................................................................................... 18-12


19.0 MARKET STUDIES AND CONTRACTS ..................................................................... 19-1


20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT 20-1

20.1 Environmental Assessment for Mining Projects ...................................................... 20-1

20.2 Project Studies ........................................................................................................ 20-2





Project No.: 162230

29 September 2011 TOC vi

20.3 Environmental Setting and Review of Environmental Baseline ............................... 20-3

20.4 Closure Considerations ........................................................................................... 20-9

20.5 Current Environmental Liabilities ........................................................................... 20-10

20.6 Closure Plan ......................................................................................................... 20-10

20.7 Permitting ............................................................................................................. 20-10

20.8 Considerations of Social and Community Impacts ................................................ 20-10

20.8.1 First Nations .................................................................................................. 20-12

20.8.2 Local Communities ........................................................................................ 20-13

20.9 Comment on Section 20........................................................................................ 20-14

21.0 CAPITAL AND OPERATING COSTS ......................................................................... 21-1

21.1 Capital Cost Estimates ............................................................................................ 21-1

21.1.1 Basis of Estimate ............................................................................................. 21-1

21.1.2 Infrastructure.................................................................................................... 21-1

21.1.3 Material Handling ............................................................................................. 21-2

21.1.4 Process Plant................................................................................................... 21-2

21.1.5 Mining .............................................................................................................. 21-2

21.1.6 Contingency Costs ........................................................................................... 21-3

21.1.7 Indirect Costs ................................................................................................... 21-3

21.1.8 Sustaining Capital ............................................................................................ 21-4

21.1.9 Mine Closure.................................................................................................... 21-4

21.1.10 Capital Cost Estimate Summary ................................................................... 21-4

21.2 Operating Cost Estimates ....................................................................................... 21-5

21.2.1 Basis of Estimate ............................................................................................. 21-6

21.2.2 Mine Operating Costs ...................................................................................... 21-6

21.2.3 Process Operating Costs ................................................................................. 21-7

21.2.4 Infrastructure Operating Costs ......................................................................... 21-7

21.2.5 General and Administrative Operating Costs ................................................... 21-7

21.2.6 Operating Cost Summary................................................................................. 21-7

21.3 Comment on Section 21........................................................................................ 21-10

22.0 ECONOMIC ANALYSIS ............................................................................................. 22-1

22.1 Valuation Methodology ........................................................................................... 22-1

22.2 Financial Model Parameters ................................................................................... 22-2

22.2.1 Mineral Resources and Mine Life ..................................................................... 22-2

22.2.2 Metallurgical Process ....................................................................................... 22-2

22.2.3 Commodity Prices and Foreign Exchange ....................................................... 22-2

22.2.4 Taxes ............................................................................................................... 22-3

22.2.5 Depreciation/Salvage Value ............................................................................. 22-3

22.2.6 Financing ......................................................................................................... 22-4

22.2.7 Capital Costs ................................................................................................... 22-4

22.2.8 Operating Costs ............................................................................................... 22-4

22.2.9 Working Capital ............................................................................................... 22-4

22.2.10 Inflation ........................................................................................................ 22-4

22.2.11 Royalty ......................................................................................................... 22-4

22.3 Financial Results .................................................................................................... 22-4

22.4 Cash Costs ............................................................................................................. 22-7

22.5 Sensitivity Analysis ................................................................................................. 22-8





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29 September 2011 TOC vii


23.0 ADJACENT PROPERTIES ......................................................................................... 23-1

24.0 OTHER RELEVANT DATA AND INFORMATION ...................................................... 24-1

25.0 INTERPRETATION AND CONCLUSIONS ................................................................. 25-1

26.0 RECOMMENDATIONS .............................................................................................. 26-1

26.1 Phase 1 .................................................................................................................. 26-1

26.2 Phase 2 .................................................................................................................. 26-1

27.0 REFERENCES ........................................................................................................... 27-1

T A B L E S

Table 1-1: Blue River Project Estimated Mineral Resources; Effective Date 29 September 2011,

Tomasz Postolski, P. Eng., Qualified Person ......................................................................... 1-8

Table 1-2: Conceptual Production Schedule ........................................................................................... 1-14

Table 1-3: Summary of Estimated Capital Costs ..................................................................................... 1-17

Table 1-4: Average Life-of-Mine Operating Cost Summary ..................................................................... 1-18

Table 1-5: Summary Financial Analysis at Various Discount Rates ........................................................ 1-19

Table 1-6: Life of Mine Cash Cost Summary ........................................................................................... 1-20

Table 1-7: Sensitivity Summary, 8% Discount Rate ................................................................................ 1-20

Table 2-1: Site Visit and Areas of Report Responsibilities ........................................................................ 2-2

Table 6-1: Blue River Exploration History Summary ................................................................................. 6-2

Table 10-1: Drill Campaign Summary ...................................................................................................... 10-3

Table 10-2: Upper Fir Deposit Trench and Bulk Samples ....................................................................... 10-3

Table 10-3: Example Drill Hole Intercept Summary Table ....................................................................... 10-7

Table 11-1: Specific Gravity Measurements for Blue River Rock Types ................................................. 11-4

Table 13-1: Head Assay Grades, Bulk Samples BS-2F and BS–2G ...................................................... 13-3

Table 13-2: Results from F81 .................................................................................................................. 13-7

Table 13-3: Results of a Sequential Hydrochloric Acid Leach of Flotation “Middling” ............................. 13-8

Table 14-1: Block Model Dimensions ...................................................................................................... 14-2

Table 14-2: Blue River Project Estimated Mineral Resources; Effective Date 29 September 2011,

Tomasz Postolski, P. Eng. .................................................................................................. 14-10

Table 14-3: Blue River Project Sensitivity of Estimated Mineral Resources to Tantalum Price;

Effective Date 29 September, 2011, Tomasz Postolski, P.Eng, (Base Case is bolded)... 14-10

Table 16-1: Minimum Stope Dimensions for Constraining the Subset of Mineral Resources within

Designed Stopes .................................................................................................................. 16-3

Table 16-2: Rock Mass Characteristics by Rock Group .......................................................................... 16-4

Table 16-3: Major Joint Sets .................................................................................................................... 16-5

Table 16-4: Stope Faces and Hydraulic Radius ...................................................................................... 16-7

Table 16-5: Mine Development and Production Forecasts .................................................................... 16-13

Table 16-6: Mining and Tailings Facility Equipment Requirements ....................................................... 16-15

Table 16-7: Mining Personnel Requirements ......................................................................................... 16-16

Table 20-1: Provincial Permits, Approvals, Licences and Authorizations .............................................. 20-11

Table 20-2: Federal Permits, Approvals, Licences and Authorizations ................................................. 20-11

Table 21-1: Summary of Estimated Capital Costs ................................................................................... 21-5

Table 21-2: Average Life-of-Mine Operating Cost Summary ................................................................... 21-8





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29 September 2011 TOC viii

Table 21-3: Operating Costs by Year ...................................................................................................... 21-9

Table 21-4: Operating Costs Per Milled Tonne........................................................................................ 21-9

Table 22-1: Project Cashflow Model ........................................................................................................ 22-5

Table 22-2: Summary Pre-Tax Financial Analysis ................................................................................... 22-7

Table 22-3: Life of Mine Cash Cost Summary ......................................................................................... 22-8

Table 22-4: Sensitivity Summary, Project Base Case NPV at 8% ........................................................... 22-9

F I G U R E S

Figure 1-1: Sensitivity Summary, 8% Discount Rate ............................................................................... 1-21

Figure 4-1: Project Location Map .............................................................................................................. 4-2

Figure 4-2: Blue River Mineral Tenure Map .............................................................................................. 4-3

Figure 7-1: Local Geology Map ................................................................................................................. 7-2

Figure 7-2: Deposit Area Surface Geology Map ....................................................................................... 7-4

Figure 7-3: Drill Collar and Vertical Section Locations ............................................................................. 7-6

Figure 7-4: Lower Road Longitudinal Section 352800 E .......................................................................... 7-7

Figure 7-5: Geology Section 5796737 N ................................................................................................... 7-8

Figure 7-6: Geology Section 5796425 N ................................................................................................... 7-9

Figure 7-7: Exploration Target Location Plan ......................................................................................... 7-15

Figure 10-1: Drill Collar Plan .................................................................................................................... 10-6

Figure 13-1: Sample BS-2F – Gravity Separation (Different Grinds) ....................................................... 13-4

Figure 13-2: Sample BS-2G – Gravity Separation (Different Grinds) ....................................................... 13-4

Figure 13-3: Rougher and Cleaners by Centrifugal Gravity Concentration .............................................. 13-5

Figure 13-4: Upgrading by Wilfley and Mozley Units ................................................................................ 13-6

Figure 16-1: Conceptual Mine Layout Plan (plan view projection) ........................................................ 16-10

Figure 16-2: Aerial View of the Mining Area from Upper Portal ............................................................. 16-11

Figure 17-1: Concentration and Refining of Blue River Mineralization ..................................................... 17-2

Figure 18-1: Proposed Site Layout Plan .................................................................................................. 18-2

Figure 21-1: Distribution of Capital Costs ................................................................................................ 21-5

Figure 21-2: LOM Operating Cost Summary ........................................................................................... 21-8

Figure 22-1: Sensitivity Summary NPV at 8% ......................................................................................... 22-9





Project No.: 162230 Page 1-1 29 September 2011

1.0 SUMMARY

AMEC Americas Limited (AMEC) was commissioned by Commerce Resources

Corporation (Commerce) to prepare a Preliminary Economic Assessment (PEA) on the

wholly-owned Blue River tantalum–niobium Project (the Project), located within the

North Thompson River valley of east–central British Columbia (BC), Canada. This

technical report supports the findings of the PEA.

Commodity prices are quoted in US dollars. All other costs, unless otherwise indicated,

are in Canadian dollars.

1.1 Key Outcomes

The key findings of the PEA are summarized below:

• Internal rate of return 9.1% (before tax)

• Net present value $18.5 million at 8% discount rate (before

tax)

• Payback 6.3 years

• Average grade in mine plan 185 ppm of Ta2O5 and 1,591 ppm of Nb2O5

• Operating cost $38.44/t milled

• Capital cost $379 million

• Proposed product Technical-grade tantalum and niobium

oxides.

1.2 Location and Access

The Blue River Project is situated 250 km north of the city of Kamloops, approximately

90 km south of the town of Valemount and 23 km north of the community of Blue

River, in the North Thompson River valley of east–central British Columbia. The

property is accessed from BC Highway 5 (Yellowhead Highway) via a 4 km well-

groomed gravel road. Within the Project area, access is by forestry service and

logging roads or by helicopter.

1.3 Mineral Tenure, Surface Rights, Royalties, and Agreements

The Project comprises 249 two-post claim, four-post claim, and mineral cell title

submission (MCX) claims in good standing that encompass just over 1,000 km2






(105,195 ha) within the Kamloops Mining Division. These claims are wholly-owned by

Commerce. There are no known royalties, back-in rights, agreements, or

encumbrances attributed to the claims.

1.4 Environment, Permitting and Socio-Economics

The Blue River Project will require approval under the federal and provincial

environmental assessment (EA) processes prior to receiving the necessary permits

and authorizations for construction and mine operation.

Overall the environmental review of the project is a process that will take up to

18 months to complete. The process will include the development of several important

documents by Commerce, including the Project Description, Assessment Information

Requirements and an Environmental Impact Assessment application, followed by the

review of these documents by the public, interested stakeholders, First Nations and

regulators.

Environmental monitoring, baseline studies and site investigations have been ongoing

at the Blue River Project site since the summer season of 2006, and will continue for

the duration of the Project. Field studies completed by specialist consultants

independent of Commerce include: site hydrology, snow course depths, fisheries and

aquatics, soils, flora and fauna and terrestrial ecosystem mapping, wildlife studies and

habitat suitability mapping, geochemistry, mineralized material and waste rock

characterization, surface water and sediment quality, groundwater, and terrain stability

assessment. No significant concerns have been identified.

A preliminary list of the Federal and Provincial permits required for operation of a mine

has been developed. The permit requirements will be reviewed and updated as the

Project advances.

Socioeconomic and cultural heritage studies have not yet been initiated for the Blue

River Project. Basic community profiling has been completed of the individual nearby

communities and relevant regional government and planning organizations.

The Blue River Project lies on lands which comprise part of the traditional territory of

the Simpcw First Nation. First Nations engagement, with respect to exploration

activities, began in 2007, and will continue for the duration of the Project. Engagement

activities have included presentations and discussions with Chief and Council, one-on-

one meetings and a site visit by elders. Members of the Simpcw First Nation

completed Archaeological Overview Assessments (AOA) over all areas of proposed

disturbance related to Commerce’s exploration activities, as well as over key areas of

potential project infrastructure. No concerns were noted by the archaeology field






technicians and exploration activities were approved to proceed by the Simpcw

archaeologist with no further recommendations for work necessary in the areas

surveyed.

Traditional Knowledge/Traditional Use studies, as well as a detailed archaeological

impact assessment, will need to be undertaken and will also involve Simpcw First

Nation participation.

On 25 October 2010, Simpcw First Nation and Commerce signed an Exploration

Agreement with respect to exploration activities on the Blue River project. Amongst

other aspects, this agreement formalized a process for ongoing discussion regarding

all exploration activities, recognizes the traditional cultural, heritage, and environmental

interest of the Simpcw, and ensures that benefits from the project are realized by

Simpcw First Nation. Commerce has also committed to involve the Simpcw in

environmental plans to gain from their knowledge of the region, as well as to keep

them informed of project goals.

Public consultation to date has included meetings with local councils (e.g., Valemount,

Barriere) and informal discussions with local land-owners.

1.5 Geology and Mineralization

The Blue River deposit is hosted within a carbonatite sill swarm with an average

thickness of 30 m and a strike length of 1,000 m. The carbonatite swarm is part of

Late Proterozoic supracrustal rocks which lie on the north-eastern margin of the

Shuswap Metamorphic Complex within the Omineca terrane. The Blue River

carbonatites are hosted in the Mica Creek assemblage of the Horsethief Creek Group.

Two units of the Mica Creek assemblage underlie much of the Project area and

comprise the lower pelite unit, and the stratigraphically overlying semipelite–

amphibolite unit. Both dolomitic carbonatites and calcitic carbonatite occur at Blue

River. Contacts between carbonatite and the host metasediments are typically sharp

and mantled by zones of metasomatized host rock, known as fenite.

Mineralization comprises niobium- and tantalum-bearing minerals that have

crystallized in carbonatite by primary magmatic concentration and in fenite. Primary

economic minerals, with their generic end-member formulae, are ferrocolumbite

((Fe,Mn,Mg)(Nb,Ta)2O6,) and pyrochlore ((Ca,Na,U)2(Nb,Ti,Ta)2O6(OH,F)).

1.6 Exploration

The Blue River area has been the subject of intermittent exploration since the

discovery of vermiculite-bearing carbonatite rock in 1949. Work by parties other than






Commerce has included claim staking, magnetometer and scintillometer geophysical

surveys, trenching, sampling, and 3,954.2 m of NQ core drilling at the Verity, Mill, Fir

and Bone Creek carbonatite occurrences.

Since Project acquisition in 2000, Commerce has completed surface mapping,

trenching, soil, rock chip, grab and channel sampling, core drilling, metallurgical

testing, bulk sampling, and Mineral Resource estimation.

1.7 Exploration Potential

The Upper Fir carbonatite has exploration potential directly northward of known

deposit extents based. Additional resource definition drilling is warranted.

The Bone Creek and Fir carbonatites have additional exploration potential along, and

across, strike. Additional in-fill soil sampling is warranted prior to diamond drilling to

assess for potential connections with the Upper Fir carbonatite.

In addition, Commerce has identified numerous tantalum-in-soil anomalies from

geochemical programs that require follow up.

1.8 Drilling

Core drilling is the most extensively used exploration tool at Blue River. As of the

2010 Mineral Resource estimate there were 215 core drill holes within the Upper Fir,

Bone Creek and Fir (Lower) carbonatites comprising 41,115 m of HQ and NQ diameter

drill holes.

Commerce completed 54 holes in 2010 and 34 holes in 2011. These holes were not

used for the 2010 Mineral Resource estimate and have not been used in the 2011

PEA. Drill core and logged lithology information from the 2010 and 2011 holes was

available for review by AMEC. The carbonatite intercepts from these holes were

compared on screen and paper cross sections against the 3D carbonatite model used

in the resource estimation and generally support the geologic interpretation. Some

discrepancies were observed which warrant local re-interpretation. Some holes

expand carbonatite volume and some holes reduce it where the carbonatite/wall rock

contact is off by 5 m to 10 m.

In 2010, McElhanney Associates of Vancouver undertook a differential GPS survey of

all 2010 drill collar locations and several pre-2010 collar locations. Some drill collar

locations have been lost due to subsequent construction activities.






Drill-hole orientations have typical azimuths of vertical, 090°, or 270° and inclinations

that range from vertical to -60°. Drill hole depths range from a minimum of 32 m and a

maximum of 388 m, averaging about 200 m. Vertical holes were generally not

surveyed down hole. The dip and azimuth of inclined drill holes were typically tested

at three points in each hole using Flexit Single Shot down-hole orientation tools.

Core recovery is very good within the waste and carbonatite rocks (typically >95%).

Drill-core samples were collected from an area approximately 1,600 m north–south by

1,000 m east-west. Average spacing between drill-hole intercepts in the resource area

is 50 m. Sample spacing increases with depth

Core sampling methods have been consistent through the 2005 to 2011 drill programs.

Core logging involved collection of both geotechnical and geological information. All

drill core was photographed prior to splitting.

Commerce regularly collected density measurements using a water displacement

method.

1.9 Sample Analysis and Security

Sampling was on average 1 m length half core, logged and sawn at a facility in the

community of Blue River. Samples were shipped to Acme Analytical Laboratories or

PRA/Inspectorate Laboratories for preparation. Analyses were completed at Acme

Analytical Laboratories. Between 2005 and 2008, Ta and Nb were analysed by ICP-

MS following a lithium metaborate/tetraborate fusion and nitric acid digestion. Analysis

in 2009 was by X-Ray fluorescence methods following a lithium metaborate fusion.

Analysis in 2010 and 2011 includes both X-Ray fluorescence and ICP methods

Quality control procedures used by Commerce to monitor laboratory results have

evolved over the life of the project. Between 2005 and 2007 there was minimal

insertion of blank, duplicate, or standard reference material control samples. During

this period, analysis of several pulp check samples was completed at six different

laboratories. In 2008 the control sample insertion rate was increased to an average of

3% for each of blanks, quarter core field duplicates, and standard reference material

control samples. In 2009 control sample insertion rates were increased to an average

of 5% per control sample type and included pulp duplicates.

Laboratory analysis of the 2010 drill-core samples has been completed and quality

control results are being reviewed. Laboratory analysis of the 2011 drill core samples

is in progress. Ta and Nb results for the 2010 and 2011 samples are expected to be

similar to the pre 2010 results.






1.10 Data Verification

AMEC completed a database verification check and concluded the collar coordinates,

down-hole surveys, lithologies, and assay databases were sufficiently free of error and

that the data are suitable to support Mineral Resource estimation.

1.11 Metallurgical Testwork

Testwork began in 2009 and continued into 2010 to develop a process flowsheet for

the Blue River Project. The testwork was based on material produced from two bulk

samples, BS-2F and BS–2G. Mineralogical analysis was performed to obtain

knowledge regarding the occurrence of the tantalum and niobium within the material.

The testwork primarily took place in two phases:

• Phase I – concentrated on the recovery of the tantalum–niobium minerals by

gravity although grinding and mineralogy investigations were also performed.

• Phase II – concentrated on the recovery and upgrading of the tantalum–niobium

minerals by flotation.

A large amount of work was performed in Phase I that showed gravity could

concentrate the material to a low-grade product, but that upgrading increasingly gave

lower levels of benefit as grade was sought. Mineralogical work completed before and

during this phase of work showed that the tantalum was not present as tantalite but

rather as the minerals ferrocolumbite and pyrochlore, which limits recovery by the

gravity route due to the low differential specific gravity between pyrochlore and gangue

minerals.

Work in Phase II saw the use of flotation concentration technology similar to that being

used for niobium-bearing carbonatites at Iamgold’s Niobec Mine in Quebec, Canada.

There was immediate success in the first phases of the work. Although there are

several stages to the concentration, the overall level of equipment, risk, and complexity

to produce a saleable or treatable concentrate is lower than the gravity route.

In both work phases, the emphasis of concentration techniques was to create a

material which would be easily upgraded by hydrometallurgical methods,

pyrometallurgical methods, or a combination of both. These processes would permit

the separation of tantalum from niobium, allowing payment for both products.

There is confidence that the concentrate could be reduced to metal by the

aluminothermic process. Subsequently there would be chlorination of the granulated

metal alloy product and distillation of the anhydrous metal chloride products to produce






high purity niobium and tantalum chlorides. Tantalum chloride is the precursor to

capacitor grade tantalum powder and can be marketed as such. Both tantalum and

niobium chlorides can be hydrolyzed and calcined to generate high-purity technical

grade oxide products for other applications.

The metallurgical testwork has shown that it is possible to concentrate the tantalum

and niobium minerals into a concentrate suitable for extraction of the metals into

saleable products. The first step of the process uses typical grinding followed by

flotation. The secondary treatment or metal extraction of the material is possible by an

existing method such as aluminothermic reduction followed by chlorine refining.

Tantalum and niobium occur as ferrocolumbite and pyrochlore, which are amenable to

conventional flotation and proven refining processes with estimated recoveries of 65%

to 70%. For the purposes of the financial analysis, it was assumed that the process

plant will have a 65% recovery for tantalum and 69% recovery for niobium in the

flotation stage. The refining process will have a 97% recovery for both tantalum and

niobium.

1.12 Mineral Resource Estimate

The PEA conducted was based on a mineral resource estimate announced in

February 2011 (“Blue River Ta-Nb Project NI 43-101 Technical Report, Blue River,

British Columbia” by AMEC with effective date 31 January 2011). AMEC used the drill

results up to the end of 2009, which includes 183 drill holes comprising 37,446 metres

of HQ drill core and 8,218 sawn core samples to develop the mineral resource

estimate.

Results from the 54 holes drilled in 2010 were compared to the existing resource

model and were found to be reasonably consistent with the predicted geology

predicted by the model. As well, the 2011 preliminary drill results were inspected at

site on vertical cross-sections and were also found to be reasonably consistent with

the predicted geology model.

Ta2O5 and Nb2O5 were estimated using an inverse distance to the power of 3 method

for the carbonatite domains. Capped drill core assays were composited down the hole

to a fixed length of 2.5 m honouring geology boundaries. A four pass interpolation

approach was used with each successive pass having greater search distances. A

hard boundary was used, meaning that composites from outside the carbonatite were

ignored in the interpolation process.






The model was validated by comparing composites to block grades on screen,

declustered global statistics checks, local bias checks using swath plots, and finally

model selectivity checks.

The Mineral Resources were classified in accordance with the definitions in the

Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Definition Standards for

Mineral Resources and Mineral Reserves (2010), incorporated by reference into NI 43-

101. Table 1-1 shows the estimated mineral resources for the Project. AMEC

cautions that Mineral Resources are not Mineral Reserves as they do not have

demonstrated economic viability.

Table 1-1: Blue River Project Estimated Mineral Resources; Effective Date 29 September 2011, Tomasz Postolski, P. Eng., Qualified Person

Ta Price Confidence Mass Ta2O5 Nb2O5 Ta2O5 Nb2O5

[US$/kg] Category [tonnes] [ppm] [ppm] [1000s of kg] [1000s of kg]

317 Indicated 36,350,000 195 1,700 7,090 61,650

Inferred 6,400,000 199 1,890 1,300 12,100

Notes:

1. Assumptions include US$317/kg Ta, US$46/kg Nb, 65.4% Ta2O5 recovery, 68.2% Nb2O5 recovery,

US$32/tonne mining cost, US$17/tonne process and refining cost. Mining losses = 0% and dilution = 0%.

2. Mineral resources are amenable to underground mining methods and have been constrained using a “Stope

Analyzer”

3. An economic cut-off was based on the Ta and Nb values per block which is variable based on the location of

blocks used in the mineral resource estimate. A block unit value cut-off ranged from $52 to $59.

4. Discrepancies in contained oxide values are due to rounding.

5. In situ contained oxides reported.

The mineral resource estimate is supported by base case price assumptions for Ta

and Nb which are higher than historic average prices. A review of publicly available

market analysts’ opinion shows a general agreement that current political and market

conditions support the probability of sustained higher prices.

The base case price for tantalum is reasonable for constraining mineral resources

based on current market conditions, but is higher than historical prices. There is a risk

that using current price assumptions may not reflect the long term price of Ta and Nb

over the life of the Project, particularly in the present volatile market conditions.

To support the 2010 Mineral Resource estimate underground mining methods were

envisioned (room and pillar or variants), with mining recovery assumed to vary from 65

to 85% depending on the success in which pillars could be mined on retreat and/or fill

is utilized.






The Mineral Resource classification at Blue River was restricted to Indicated or

Inferred, based on the following:

• Confidence limits resulting from drill hole spacing studies

• Concerns of analytical precision and accuracy for the sample dataset

• Local discrepancies in the model identified with the 2010 drilling

• Lack of metallurgical testwork on the final stage of the proposed metallurgical

process.

The assumptions used for assessment of reasonable prospects of economic extraction

used to support the 2010 Mineral Resource estimate were modified for the 2011 PEA

mine plan or in the financial analysis based on that mine plan.

1.13 Proposed Mine Plan

The Blue River deposit is defined by the following physical conditions:

• Tantalum–niobium mineralization is hosted in carbonatite

• Continuous mineralization is found in moderately flat and wide carbonatite bodies

with moderate dips

• Mineralized areas whose thickness can range from 20 m to 80 m in height are

expected in several zones

• Steep topography provides access to the mineralized areas in the form of hillside

adits

• Fair to good rock conditions are expected in the majority of the deposit

• At least two faulted zones have been identified.

As mineralization is close to surface, extraction of the mineralized material can occur

by open pit or underground methods, or a combination of both. Initial assessment

identified technical challenges and increased costs for tailings and waste rock

disposition related to local topography, stream courses, and prevailing climate

conditions. For these reasons, a decision was made to consider an underground

mining scenario for PEA purposes. The assumed mining method is sub level open

stoping with no backfill and no pillar recovery.

A processing rate of 7,500 t/d was assumed for the conceptual design of an

underground mine for the Blue River Project.






Price assumptions used in mine planning were US$317/kg tantalum metal and

US$46/kg niobium metal contained in oxide product.

1.13.1 Geotechnical Considerations

Rock types of the Blue River Project have been grouped into two main geotechnical

domains: Intrusive and Layered Rocks. The Intrusive group encompass carbonatite

and fenite rocks while the Layered Rocks group encompass gneiss and amphibolites.

Generally the rock mass ratings (RMR) indicated rock that can be considered to be

“good” in RMR terms.

1.13.2 Dilution Considerations

Material deemed to be mined by bulk mining methods represents 84% of the

resources. Within the mineable shapes there was internal dilution of 2% waste rock. It

was assumed that during mining 2% of waste material would be added as external

dilution and 2% of the broken material would not be recovered from the stopes due to

operational conditions.

The geotechnical investigation indicates that an extraction ratio of 67.5% is

reasonable. Applying this to the subset of the Mineral Resources considered in the

mine plan results in an overall mining extraction of 58% and provides 25.0 Mt of

material as run-of-mine (ROM) production to be processed. Applying internal and

external mining dilution, the overall subset Mineral Resource grades were diluted to

185 ppm of Ta2O5 and 1,591 ppm of Nb2O5 for mine planning purposes.

1.13.3 Drilling and Blasting

AMEC considered the implementation of conventional drilling methods. Due to the

high precipitation in the region and water continuity along fractures and rock layers,

wet conditions were assumed for development and stoping areas. The use of bulk-

blasting systems based on emulsion type explosives was assumed.

1.13.4 Mine Development

The deposit will be accessed through two main portals, the Upper and Lower Portals;

these portals are located in places where the deposit crops out on the hillside. The

Upper Portal will be used as the main entry and will have most of the mine services;

the Lower Portal will be used for haulage trucks access. Access to the portals will be

by a road upgraded from existing exploration roads.






The mine will be accessed by adits driven in pairs from the portals. For this study, all

entries, ramps, drifts and crosscuts are considered to be 5 m wide by 5 m high with

semi-arched backs. Ramps will be driven at grades to a maximum of 15% to provide

access to the production areas. Two ramps or adits will be driven to each area to

provide single-way traffic of haulage trucks and to facilitate the implementation of

ventilation circuits.

Top access crosscuts are driven from the main ramps to each level on vertical

intervals between 20 to 30 m. Stope access crosscuts are driven at level from west to

east. Bottom access crosscuts are driven to function as mucking drifts. The total

underground development was estimated as 92,500 m for the life-of-mine.

1.13.5 Mineralized Material and Waste Haulage

Radio remote-controlled load-haul-dump units will be used to extract the mineralized

material out in the stope beyond the safety of the brow. The mineralized material from

stopes will be loaded directly to the haulage trucks that will be spotted at the end of the

crosscut. Underground trucks will haul the mined material through the access drifts

and ramps, unloading into the primary crusher surface stockpile near the Lower Portal.

The mine plan envisages that tailings material will be dry-stacked, and waste material

will be stored in the same general area. For convenience, the combined tails and

waste rock storage area is referred to as the “co-disposal facility”.

Waste from development will be initially utilized for construction of a structural shell for

the co-disposal site on surface, which will be located in an area east of the processing

plant site.

A conveyor will be used to transport this material from the mine to a stockpile by the

plant site. A surface road developed at +10% grade will connect the plant with the co-

disposal site.

Trucks will haul waste from the plant to the co-disposal site when required.

1.13.6 Mine Services

Underground mine services such as ventilation and air heating, compressed air, water

for drilling and power supply will be provided to the mine via adits from the portals.

The sub-station for the main power distribution system and the air compressors will be

installed in facilities located adjacent to the Upper Portal.






Other mine services will include all the systems and supplies needed for the mining

operations, including: explosives storage, communications, monitoring and control

systems, road maintenance and mine equipment maintenance.

Portable self-contained refuge stations will be provided for the mine and will be located

at convenient locations.

1.14 Mine Production Forecasts

Production was estimated at 2.7 Mt/a of mineralized material to be extracted over

10 years. The first year was considered as preproduction, leaving nine years of full-

scale production.

At this preliminary level of study the stope mining sequence was not defined and

therefore average grades were used for each year in the mine plan.

There is opportunity to increase the net present value of the project by mining higher-

grade zones early in the mine life providing that the sequence and overall recovery of

the stopes is not negatively affected.

The proposed production schedule is presented in Table 1-2.

Mine equipment and personnel requirements were defined and are appropriate to the

proposed production plan.

1.15 Process Design

The design for the process facilities considered a nominal processing capacity of

7,500 t/d. Where data were not available at the time of flowsheet development, AMEC

developed criteria for sizing and selection of equipment based on comparable industry

applications, benchmarking, and the use of modern modelling and simulation

techniques.

The mineral processing and the refining are based on conventional technology and

industry-proven equipment. A mineral processing method using a standard grind-

flotation procedure to make a concentrate of ferrocolumbite-pyrochlore is assumed for

Blue River material.

Metallurgical testing indicates a mineral concentrate assaying about 30% combined

Nb–Ta pentoxide within the recovery range of 65% to 70% is possible.






Run of mine mineralized material will be crushed to minus 5/8” and fed into a

comminution circuit comprised of a rod and ball mill using cycloning for classification.

After grinding, the flotation feed will be first de-slimed using high frequency fine

screens and cyclones. The coarse product resulting from de-sliming will be sent to

four concentration steps that will include pyrrhotite flotation, carbonate flotation, and

magnetite separation with all three concentrates from these processes reporting to

tailings. The fourth step, pyrochlore flotation, will recover a concentrate which is

reground and cleaned in five stages of cleaning. The mass of material will be reduced

substantially, to less than 1% of the feed into the plant.

This concentrate would be further processed to produce marketable separate Ta and

Nb products. The proposed processes are mature, are already used industrially, and

consist of reducing the concentrate to metals as ferroalloys in a standard

aluminothermic furnace followed by chlorinating the alloys and distilling the product to

separate high purity metal chlorides, TaCl5 and NbCl5. Recoveries from concentrate to

pure chlorides are expected to be 97%. Both Ta and Nb chloride products are then

readily converted and marketed as high purity oxides Ta2O5 and Nb2O5 respectively.






Table 1-2: Conceptual Production Schedule

Production Schedule Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5

Nominal Production Rate tpd 7,500 7,500 7,500 7,500 7,500 7,500

Scheduled working days days/year 360 360 360 360 360 360

Production t/year 2,700,000 2,700,000 2,700,000 2,700,000 2,700,000

Ta2O5 Grade ppm 185 185 185 185 185

Nb2O5 Grade ppm 1,591 1,591 1,591 1,591 1,591

Lateral Development m/year 12,000 15,000 12,834 10,980 9,394 8,038

Capital Development m/year 12,000 5,000 1,216 1,040 890 761

Operational Development m/year 0 10,000 11,618 9,940 8,505 7,276

Production Schedule Units Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Total

Nominal Production Rate tpd 7,500 7,500 7,500 7,500 7,500

Scheduled working days days/year 360 360 360 360 360

Production t/year 2,700,000 2,700,000 2,700,000 2,700,000 700,000 25,000,000

Ta2O5 Grade ppm 185 185 185 185 185 185

Nb2O5 Grade ppm 1,591 1,591 1,591 1,591 1,591 1,591

Lateral Development m/year 6,877 5,884 5,034 4,307 2,153 92,500

Capital Development m/year 651 557 477 408 0 23,000

Operational Development m/year 6,225 5,326 4,557 3,899 2,153 69,500






1.16 Tailings and Waste Management

A series of tailings storage location screening assessments were carried out during

2008, 2009, and 2010 focusing mainly on conventional tailings storage. Some

evaluation of potential waste rock storage sites and a tailings drystack facility was

undertaken during studies completed in 2009.

Site WSF3, about 8 km from the proposed plant site, and identified in 2009, was

selected as the preferred location for a tailings drystack. WSF3 has flatter slopes than

other site alternatives and therefore has less stability concerns. It is also located in

closer proximity to the proposed plant site, although uphill haulage will be required.

The facility was designed to hold a total storage volume of approximately 20.9 Mm3.

Surface water management systems will be required for the tailings drystack to divert

non-contact (clean) water around the facility, and to collect run-off water which had

been in contact with the tailings drystack area.

1.17 Planned Project Infrastructure

The planned Upper and Lower Portals will be located about 4 km from the plant site.

At the front of the Upper Portal sufficient space will be provided to accommodate the

required facilities for operation.

The plant service building will be a multi-purpose complex in a two-story building to be

located east of the process building. A 24 m by 36 m truck shop on the southwest end

of the site will be operated by a qualified contractor. The warehouse will be a 24 m by

50 m Coverall-type fabric building.

The potable water system and layout is designed to service buildings and a

process/administration workforce of 120 persons. Potable water for the mining area

will be constructed as part of the portal/underground works.

A modular sewage treatment system will be installed as part of the initial construction

infrastructure.

Waste-water treatment sludge will be trucked away to a nearby municipal facility or

approved landfill.

Waste lubrication and hydraulic oils from vehicle maintenance will be stored in

dedicated tanks and sent to a recycling facility offsite.






Contractors and employees will commute from nearby towns such as Blue River and

Valemount during construction. No on-site permanent accommodation will be

provided for personnel. It is assumed that the workforce, including management staff,

will reside in the nearby communities and will commute, via buses on a daily basis.

The road access design includes a short new road with a 7.2 m wide gravel surface

from the existing road to plant site about 80 m in length and a 1.5 km new service road

from the existing road to the Upper Portal and upgrades to the current access road.

There are two main haul roads, one from the mine portal to site of about 4 km length,

and the other from the site to the co-disposal facility, with a length of about 8 km.

An existing 80 m-long bridge crossing over the Thompson River has a limited load

capacity and might not qualify for crossing heavy loads during construction or long-

term use during the life of the mine. Therefore a new bridge has been included in

capital cost estimate. Using the existing railway for shipment should be investigated in

the next phase of study.

Ammonium nitrate, blended emulsion, and explosives will be delivered to site on

demand by contractors. Fuel will be delivered to the mine site using tanker trucks. A

crushing and stockpiling facility will be required during construction to provide crushed

product for roads and surfacing. AMEC considers that there is no need to establish a

concrete batch plant, and that concrete supply from nearby towns will be more

economic.

1.18 Markets

Commerce has prepared assessments of the tantalum and niobium markets which

outline the supply and demand for tantalum and niobium. The tantalum assessment

was prepared by a tantalum market expert, although he is not independent of

Commerce. His analysis reflects the general consensus of other analysts regarding

the tantalum market expressed in publicly available information.

The niobium assessment was prepared by an independent niobium expert and also

reflects the general consensus of analysts in publicly-available information for the

niobium market.

As the Project is still at an early evaluation stage, Commerce has not initiated requests

from potential buyers for expression of interests from potential buyers of the proposed

Blue River products and has not negotiated any purchase or off-take agreements.






1.19 Capital Costs

All costs are expressed in first quarter (Q1) 2011 Canadian dollars. No allowance has

been included for escalation, interest or financing fees, taxes or duties, or working

capital during construction. The level of accuracy for the estimate is +40/-20% of

estimated final costs, as per the Association of Advanced Cost Estimators (AACE)

Class 5 (scoping level) definition.

Contractor-mining is not envisaged for pre-production development. The estimate

covers the direct field costs of executing the project, plus the Owner’s indirect costs

associated with design, construction, and commissioning. The preproduction costs are

capitalized and include all the expenditures before Year 1 of production.

The total estimated capital cost to design and build the Blue River tantalum project at

7,500 t/d capacity is $379 million. The estimate is summarized in Table 1-3.

Table 1-3: Summary of Estimated Capital Costs

Calendar year Total

(x $1,000) 2011

(x $1,000) 2012

(x $1,000)

Project year 1 2

Production year -2 -1

Capital expenditure

Initial Capital Infrastructure 29,500 10,300 19,200

Process Initial Capital 116,200 40,700 75,600

Mining Initial Capital 89,400 89,400

Material Handling 8,000 8,000

Contingency 43,600 12,800 30,900

Indirect/Owner Costs 92,300 29,600 62,600

Total 379,000 93,400 285,600 Note: Summation discrepancy due to rounding.

1.20 Operating Costs

The operating costs for the Blue River project are based on an Owner-operated mining

fleet and process facility and have been prepared in first quarter 2011 Canadian

dollars.

Operating costs over the life-of-mine are estimated at $38.44/t milled.

Operating costs include the three key areas of mining, process, and overall general

and general and administrative costs. The estimates are based on the staffing level,

consumables, and expenditures detailed as part of the underground mine plan and

process design.






Average operating costs are illustrated in Table 1-4.

Table 1-4: Average Life-of-Mine Operating Cost Summary

Summary of Average Production Costs LOM Total (x $1,000)

Cost per Tonne Milled ($/t)

Mining 528,900 21.16

Process 338,500 13.54

Material Handling 18,500 0.74

G&A 75,000 3.00

Sub-total 960,900 38.44

1.21 Financial Analysis

The financial analysis is partly based on 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 Assessment based on these Mineral Resources will

be realized. Approximately 15% of the Mineral Resources that support the financial

model have been classified as Inferred Mineral Resources.

The results of the economic analyses discussed in this section represent forward-

looking information as defined under Canadian securities law. The results depend on

inputs that are subject to a number of known and unknown risks, uncertainties and

other factors that may cause actual results to differ materially from those presented

here.

Information that is forward-looking includes:

• Mineral Resource estimates

• Assumed commodity prices and exchange rates

• The proposed mine production plan

• Projected recovery rates

• Infrastructure construction costs and schedules

• Assumptions that an EA will be approved by Provincial and Federal authorities.

The project has been evaluated using a discounted cash flow (DCF) analysis. Cash

inflows consist of annual revenue projections for the mine and two years of






preproduction. Cash outflows such as capital and operating costs are subtracted from

the inflows to arrive at the annual cash flow projections.

The resulting net annual cash flows are discounted back to the date of valuation end-

of-year 2010 dollars and totalled to determine NPVs at the selected discount rates.

The IRR is calculated as the discount rate that yields a zero NPV. The payback period

is calculated as the time needed to recover the initial capital spent.

The Project Base Case (8% discount rate) returns an NPV of $18.5 M and an IRR of

9.1% before tax. Table 1-5 summarizes the NPV for the Project at a range of discount

rates, with the base case highlighted.

Table 1-5: Summary Financial Analysis at Various Discount Rates

Summary of Cash Flow Pre-tax

Cumulative net cash flow

Undiscounted CAD$000 236,631

Net present value

Discounted at 5% CAD$000 80,349



Discounted at 8% (Project Base Case) CAD$000 18,487


Discounted at 10% CAD$000 (13,514)

Internal rate of return % 9.1

Payback period Years 6.3 Note: base case is highlighted

1.22 Cash Costs

The cash cost value represents the cost incurred to produce 1 kg of primary product

after deducting the revenue from sales of secondary products. Since the price

analysis for the report is performed around Ta price variation, Ta is chosen as the

main product and Nb is treated as the secondary product for the assessment of cash

cost.

The tantalum cash cost is calculated to be approximately $24.91/kg contained in oxide

product (after credit for niobium contribution) as shown in Table 1-6.






Table 1-6: Life of Mine Cash Cost Summary

Section LOM Total (x $1,000)


Cost per Kg Ta Payable ($/kg)

Cash costs

Mining 528,900 21.16 220.13

Process 338,500 13.54 140.87

G&A 75,000 3.00 31.21

Material Handling 18,500 0.74 7.71

Sub-total 960,900 38.44 399.92

Credits

Nb (901,100) (36.04) (375.01)

Sub-total (901,100) (36.04) (375.01)

Adjusted cash costs

Total 59,800 2.40 24.91 Note: The figures in this table do not include considerations of working capital or royalty payments

1.23 Sensitivity Analysis

The sensitivity analysis shows that the project is more sensitive to changes in

operating expenditures than capital expenditures. The project is most sensitive, in

order, to changes in exchange rate, operating expenditure, niobium price, tantalum

price and capital expenditure. Since the sales currency is US dollars and operational

costs are in Canadian dollars, a rising US dollar value versus Canadian dollar value

improves the mine profitability.

The project IRR increases to 14.4% and the NPV increases to $125 M at an 8%

discount rate if a Ta price of US$380/kg (20% increase) is assumed.

Sensitivities are summarized in Table 1-7 and Figure 1-1 for the 8% discount base

case rate.

Table 1-7: Sensitivity Summary, 8% Discount Rate

SENSITIVITY OF NPV @ 8% Change in Factor

-30% -20% -10% 0% 10% 20% 30%

Fa

cto

r

Exchange rate 448.7 269.5 130.0 18.5 (72.8) (148.8) (213.2)

Capital expenditure 117.9 84.8 51.6 18.5 (14.6) (47.8) (80.9)

Operating expenditure 190.5 133.1 75.8 18.5 (38.8) (96.2) (153.5)

Nb price (140.9) (87.8) (34.6) 18.5 71.6 124.7 177.9

Ta price (123.3) (76.1) (28.8) 18.5 65.8 113.0 160.3






Figure 1-1: Sensitivity Summary, 8% Discount Rate

1.24 Interpretations and Conclusions

The Blue River project has a 10 year mine life and capital costs, at $379 M, could be

paid back in 6 years. The project is most sensitive to commodity price. Available

quotes on tantalum price through 2011 were significantly higher than the tantalum

price used to support the 2010 Mineral Resource estimate and the 2011 PEA. A

higher tantalum price could not only improve profitability but also increase the mine

life. Additional exploration potential could also provide additional mine life. A two or

more times capital payback is possible.


at the Blue River Project site since the summer season of 2006. First Nations

engagement began in 2007. Public consultation to date has included meetings with

local councils and informal discussions with local land-owners.

Opportunities to improve the project NPV include:

• Optimization of the mine plan to mine higher-grade zones early in the mine life

providing that a practical mining sequence can be implemented and the overall

recovery of the Mineral Resources is not negatively affected






• Optimization of the mine layout to minimize development costs

• Advanced geotechnical studies to identify and understand ground conditions which

could allow an increase in the size of stopes and production drifts

• Optimization of the supply and pricing of reagents for the refining.

The risk factors are:

• The base case mineral resource estimate is supported by current tantalum and

niobium prices which are significantly higher than historic average prices and may

not reflect long term prices. Market analysts are in general agreement that current

political and market conditions support the probability of sustained higher prices,

but this may not occur.

• Commerce has not initiated requests from potential buyers for expression of

interests from potential buyers of the proposed Blue River products and has not

negotiated any purchase or off-take agreements.

• The proposed refining methods have been used in commercial applications but

have not been demonstrated in test work of Blue River material.

• Testwork to date has not considered factors such as water recycling. A water

treatment plant may be required and may result in increased capital costs.

• The financial analysis is partly based on 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 based on

these Mineral Resources will be realized

• The Blue River Project will require approval under the federal and provincial


and authorizations for construction and mine operation. Overall the environmental

review of a project is a process that can take up to 18 months to complete.

• Traditional Knowledge/Traditional Use (TK/TU) studies, as well as a detailed

archaeological impact assessment will need to be undertaken.

The projects warrant additional work to examine the opportunities and mitigate the

risks. On completion of this work Commerce may consider proceeding with a pre-

feasibility study.






1.25 Recommendations

AMEC has recommended a two-stage work program of $3 M. The first stage

comprises a re-estimation of Mineral Resources and incorporates results of the 2010

and 2011 drill programs. The second stage comprises completion of mining and

related studies that can be used to support an additional mineral resource update, and

conversion of Mineral Resources to Mineral Reserves through a pre-feasibility study.

Phase one consists of an update to the Mineral Resource estimate, and is budgeted at

$350,000 to $400,000. Phase 2 consists of completion of mining and related studies

so as to position Commerce to make a decision on completion of a pre-feasibility

study. This phase is estimated at $2.6 M.





Project No.: 162230

29 September 2011 Page 2-1

2.0 INTRODUCTION

2.1 Terms of Reference

AMEC Americas Limited (AMEC) was commissioned by Commerce Resources

Corporation (Commerce) to prepare a Preliminary Economic Assessment (PEA) on the

wholly-owned Blue River tantalum–niobium Project (the Project), located within the

North Thompson River valley of east–central British Columbia (BC), Canada. This

technical report supports the findings of the PEA.

Commodity prices are quoted in US dollars. All other costs are in Canadian dollars

unless otherwise indicated. Volumes, weights, and distances are metric unless

otherwise indicated.

2.2 Qualified Persons

The Qualified Persons (QPs) for the Report are AMEC employees, based out of

AMEC’s Vancouver office, as follows:

• Mr. Albert Chong, P.Geo., Principal Geologist

• Mr. Tomasz Postolski, P.Eng., Senior Geostatistician

• Mr. Ramon Mendoza Reyes, P.Eng., Principal Engineer

• Mr. Tony Lipiec, P.Eng., Principal Metallurgical Engineer

• Mr. Behrang Omidvar, P.Eng., Financial Analyst

2.3 Site Visits and Scope of Personal Inspection

Mr. Chong visited the property July 11 to 16, 2010, June 27 to 30, 2011, and

September 6 to 14, 2011. During the site visits Mr. Chong inspected outcrops, drill-

hole collars locations, and geologic interpretation. Mr. Chong reviewed procedures for

diamond drilling, logging, sampling, core storage, and sample shipment. In 2011 Mr.

Chong also completed an aerial inspection of proposed future infrastructure locations.

Mr. Mendoza Reyes conducted a site visit July 12 to 16, 2010. During the site visit Mr.

Mendoza inspected sites amenable for locating potential infrastructure.

Mr. Postolski conducted a site visit June 27 to 30, 2011. During the site visit Mr.

Postolski was unable to access the property due to a road wash-out but was able to





Project No.: 162230


complete an aerial inspection of proposed future infrastructure locations. The purpose

of the trip was to review the geology interpretation and resource model 3D solid which

was completed at the Commerce office in Blue River. During this visit Mr. Postolski

was able to visit the core logging facilities and observe core logging procedures.

Mr. Lipiec did not complete a site visit.

Mr. Omidvar did not complete a site visit.

A summary of QP site visits is shown in Table 2-1.

Table 2-1: Site Visit and Areas of Report Responsibilities

Qualified Person Site Visit Sections of Report Responsibility

Albert Chong, P.Geo. July 11 to 16, 2010

June 27 to 30, 2011

September 6 to 14, 2011

Sections 1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12,

20, 23, 24, 25, 26, and 27

Tomasz Postolski, P.Eng. June 27 to 30, 2011

Section 14 and those portions of the

Summary, Interpretation and Conclusions,

and Recommendations that pertain to that

Section

Ramon Mendoza Reyes,

P.Eng

July 12 to 14, 2010 Sections 15, 16, 18, and those portions of

the Summary, Capital and Operating

Costs, Interpretation and Conclusions, and

Recommendations that pertain to those

Sections

Tony Lipiec, P.Eng No site visit Sections 13 17, 18, 21, and those portions

of the Summary, Interpretation and

Conclusions, and Recommendations that

pertain to those Sections

Behrang Omidvar, P.Eng No site visit Sections 19, 21, 22, and those portions of

the Summary, Interpretation and

Conclusions, and Recommendations that

pertain to those Sections





Project No.: 162230


2.4 Effective Dates

The Report has a number of effective dates, as follows:

• Effective date of the Mineral Resources: 29 September 2011

o The database closure includes all drill information up to the end of

2009.

o Results from the 54 holes drilled in 2010 were compared by AMEC to

the existing resource model and the results were found to be

reasonably consistent.

o The 2011 preliminary drill results were inspected at the Project site on

vertical cross-sections by AMEC during September 2011, and the

results were found to be consistent with the geological interpretation.

• Effective date of the Report: 29 September 2011.

The overall effective date of the Report is based on the date of the completion of the

Preliminary Economic Assessment and the financial analysis and review of the latest

available drilling information. There has been no material change to the Project

between the effective date of the Report, and the signature date.

2.5 Information Sources and References

The primary data source for this Report is the PEA, entitled:

AMEC, 2011: Commerce Resources Corporation, Blue River Tantalum–Niobium

Project, Blue River, British Columbia, Preliminary Economic Assessment: unpublished

report prepared by AMEC Americas Ltd. for Commerce Resources Corporation, dated

August 2011.

The PEA was prepared by AMEC. Information for the Report was obtained from work

completed by AMEC, and materials provided by, and discussions with, Commerce

personnel and third-party contractors retained by Commerce.

Additional information was provided by Commerce personnel in the areas of

environmental studies and permitting. This information was based on reports prepared

by third-party consultants retained by Commerce.





Project No.: 162230


Marketing studies were provided as confidential reports to AMEC. AMEC has

reviewed the information provided and considers that the studies support the

commodity prices used to constrain the Mineral Resource estimates and financial

evaluation.

Except where noted, figures were generated by AMEC.

Reports and documents listed in the Section 3, Reliance on Other Experts and Section

27, References sections of this Report were also used to support preparation of the

Report. Additional information was sought from Commerce personnel where required

to support preparation of this Report.

2.6 Previous Technical Reports

Commerce has previously submitted the following technical reports on the Project:

Chong, A., and Postolski, T., 2011: Commerce Resources Corporation, Blue River Ta-

Nb Project, Blue River, British Columbia, NI 43-101 Technical Report: unpublished

technical report prepared by AMEC Americas Ltd. for Commerce Resources

Corporation, effective date 31 January 2011

Stone, M., and Selway, J., 2010: Independent Technical Report, Blue River Property,

Blue River, British Columbia, Canada: unpublished technical report prepared by

Caracle Creek International Consulting Inc. for Commerce Resources Corporation,

effective date 24 August, 2009

Gorham, J., 2007: Technical Report on the Upper Fir Ta-Nb Bearing Carbonatite:

unpublished technical report prepared for Commerce Resources Corporation, effective

date 20 June 2007.





Project No.: 162230


3.0 RELIANCE ON OTHER EXPERTS

The QPs state that they are qualified persons for those areas as identified in the

appropriate QP “Certificate of Qualified Person” attached to this Report. The authors

have relied upon and disclaim responsibility for information derived from the following

reports pertaining to mineral tenure, surface rights, royalties, environment and

permitting, and marketing.

3.1 Mineral Tenure

The AMEC QPs have not reviewed the mineral tenure, nor independently verified the

legal status, ownership of the Project area or underlying property agreements. AMEC

has fully relied upon, and disclaims responsibility for, information derived from legal

experts for this information through the following document:

Letter from Clark Wilson LLP titled Commerce Resources Corp. – Mineral Claim Title

Opinion to Mr. Greg Kulla, dated 29 October 2010

Information from this letter has been used in Section 4.2 of this Report.

3.2 Surface Rights

The AMEC QPs have not reviewed the status of the Project surface rights, nor

independently verified the surface rights status of the Project area. AMEC has fully

relied upon, and disclaims responsibility for information derived from experts for this

information through the following document:

Letter from David Hodge, President Commerce Resources Corp. to Mr Greg Kulla,

entitled “Commerce Resources Corp – Blue River Property Encumbrances and

Surface Rights”, dated 14 October 2011

This information has been used in Section 4.3 of this Report.

3.3 Royalties and Agreements

The AMEC QPs have not reviewed the royalties and agreements for the Project, nor

independently verified the royalties and agreements status of the Project area. AMEC





Project No.: 162230


has fully relied upon, and disclaims responsibility for information derived from experts

for this information through the following document:

Letter from David Hodge, President Commerce Resources Corp. to Mr Greg Kulla,

entitled “Commerce Resources Corp – Blue River Property Encumbrances and

Surface Rights”, dated 14 October 2011

This information has been used in Section 4.4 of this Report.

3.4 Environmental, Permitting and Liability Issues

The AMEC QPs have not reviewed the permitting requirements, nor independently

verified the permitting status of the Project area. AMEC has fully relied upon, and

disclaims responsibility for information derived from experts for this information through

the following document:

Letter from Sage Resource Consultants Ltd. titled Commerce Resources Corp. Upper

Fir Deposit Preliminary Economic Assessment – Independent Professional Opinion on

Environmental Permitting and Liability Issues to Mr. Greg Kulla and dated 29

September 2010.

This information has been used in Section 20 of this Report.

3.5 Markets

The AMEC QPs have relied on tantalum and niobium market analyses derived from

experts for this information through the following documents:

Confidential memo from Dr. Axel Hoppe titled “Ap#6 Introduction to Tantalum

Markets_Finalpdf_2June09.pdf” received 18 October 2010

Confidential memo from Michel Robert titled “Niobium_v3jh.doc” received 18 October

2010

Confidential memo from Michel Robert titled “Niobium_v3jh.doc” received 19 October

2010

Dr. Hoppe is an internationally acknowledged leader in the tantalum field and is

Chairman of the Board of Directors for Commerce Resources.





Project No.: 162230


Mr. Robert has extensive experience in niobium markets and is independent of the

company.

This information has been used in Section 19 of this Report, and to assess reasonable

prospects of economic extraction in Section 14.10.





Project No.: 162230


4.0 PROPERTY DESCRIPTION AND LOCATION

The Project is located within the North Thompson River valley of east-central British

Columbia 25 km to 60 km north and northeast of the community of Blue River, British

Columbia (Figure 4-1).

The NTS sheets which cover the Project are: 83D.004-.006; 83D.014-.016; 83D.024-

.027; 83D.034-.037; 83D.045-.047.

The Project is centered at approximately 52° 19' N latitude and 119° 10' W longitude.

4.1 Project Ownership

The Project is wholly-owned by Commerce and held in the name of Commerce

Resources Corp.

4.2 Mineral Tenure

The Project comprises 249 two-post claim, four-post claim, and mineral cell title

submission (MCX) claims in good standing that encompass just over 1,000 km2

(105,195 ha) within the Kamloops Mining Division. The claim boundaries are shown in

Figure 4-2.

A table listing claim details is included in Appendix A.

Currently, all but five of the mineral claims were valid until 31 March 2019. The

exceptions are the four “Wasted” claims tenure numbers 588427-30) which were to

expire on 31 July 2011 and the Join (tenure number 798362) which were to expire on

25 June 2011. Commerce filed a 2010 work program by the end of day on 25 June

2011 to group the two blocks of claims with the larger claims area. About $977,000

worth of eligible 2010 work will be filed for assessment credit at a filing cost of

approximately $42,000. When approved, this will bring all Blue River claims to a

common expiry date of 31 March 2021.

Property boundaries are established in accordance with the Mineral Tenure Act of

British Columbia. Commerce has staked the claims by a combination of ground and

on-line staking.

Two-post and four-post claims were established through a legacy system of ground

staking which involved physically establishing claim posts on the ground.





Project No.: 162230


Figure 4-1: Project Location Map

Figure courtesy Commerce Resources Corp.





Project No.: 162230


Figure 4-2: Blue River Mineral Tenure Map

Note: Figure courtesy Dahrouge Consulting Ltd. Grid is in metres for UTM NAD83 Zone 11.





Project No.: 162230


MCX claims are established using the Government of British Columbia’s Mineral Titles

Online (MTO) staking system. MTO is an Internet-based mineral titles administration

system that allows the mineral exploration industry to acquire and maintain mineral

titles by selecting the area on a seamless digital GIS map of British Columbia. The

electronic Internet map allows selection of single or multiple adjoining grid cells. Cells

range in size from approximately 21 ha (457 m x 463 m) in the south to approximately

16 ha at the north of the province. All boundaries are oriented north–south and east–

west.

4.3 Surface Rights

Commerce holds no surface rights on the property. Legal access to the property is

provided through the British Columbia Mineral Tenure Act. The Act provides for a

recorded claim holder to use, enter and occupy the surface of a claim or lease for the

exploration and development or production of minerals or placer minerals, including

the treatment of ore and concentrates, and all operations related to the exploration and

development or production of minerals or placer minerals and the business of mining.

Access to surface rights held by third parties typically requires compensation.

There are surveyed parcels with surface rights held by other parties which overlap the

property mineral tenure claims. These parcels occur along the western edge of the

property and most are tree farm licences. Commerce is not aware of any material

issues that would prevent negotiation for access or surface rights of these surveyed

parcels should they be required in the future. The claims do not host Mineral

Resources, and currently no carbonatites are known within the claims.

4.4 Royalties and Agreements

There are no royalties, back-in rights, payments, or other agreements or

encumbrances to which the Blue River property is subject to other than the annual

claim maintenance fees due to the government as set out by the British Columbia

Mineral Tenure Act and Regulations.

4.5 Permits

Permits required to support Project development are discussed in Section 20.





Project No.: 162230


4.6 Environment

Environmental studies are discussed in Section 20.

4.7 Social and Community Impact

The social and community impact assessments of the Project are discussed in Section

20.

4.8 Comment on Section 4

In the opinion of the AMEC QPs, the following conclusions are appropriate:

• The mineral tenure has been acquired in accordance with relevant BC regulations

• The mining tenure is valid and is sufficient to support declaration of Mineral

Resources

• Work filed in 2011 with relevant regulatory authority will bring all Blue River claims

to a common expiry date of 31 March 2021

• There are no known royalties, back-in rights, agreements, or encumbrances

attributed to the claims

• Exploration activities have been conducted within the regulatory framework

required by the BC Government

• Additional permits will be required for Project development.





Project No.: 162230


5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Accessibility

The Project is located 23 km north of the community of Blue River, British Columbia,

approximately 250 km north of the city Kamloops and approximately 90 km south of

the town of Valemount. The property is accessed from B.C. Highway 5 (Yellowhead

Highway) via a 4 km well groomed gravel road.

The Upper Fir and Bone Creek deposits can be reached from the Bone and Gum

Creek forestry service road which branches from Highway 5 approximately 23 km

north of Blue River. The east side of the property can be reached by forest service

roads along the west side of Kinbasket Lake and up Howard Creek. Logging roads on

Serpentine, Bone, Hellroar and Mud Creeks allow four-wheel drive and quad bike

access to most of the property. Access to remaining portions of the property is by

helicopter.

5.2 Climate

The local climate is typical of the interior of British Columbia. The area is part of a “wet

belt” that occupies part of eastern British Columbia. Heavy snow falls over almost

every winter, in which temperatures stay close to the freezing point when maritime air

dominates. Rain is frequent in other seasons. Summer days are typically warm or

occasionally hot, with thunderstorms often spawning over the nearby mountains.

In July, the average daily temperature is 16.4°C and the average rainfall accumulation

is 97.5 mm for Blue River (Environment Canada Climate Normals 1971–2000 web site:

http://climate.weatheroffice.gc.ca/climate_normals/index_e.html). In January, the

average daily temperature is -9°C and the average snowfall accumulation is 109 cm

for Blue River. The average snow depth is 83 cm in February. Local rainfall and

snowfall accumulations on parts of the property may be much higher due to elevation

and orographic effects.

Drilling is feasible from mid-May through early to mid-October. Snowfall can exceed

10 m making winter drilling very expensive and difficult, but not impracticable. It is

expected that any future mining activity could be conducted year-round.





Project No.: 162230


5.3 Local Resources and Infrastructure

There is no existing Project infrastructure. Exploration activities are currently supplied

out of Kamloops, Clearwater and Valemount.

The city of Kamloops currently supports mining operations at the New Afton and

Highland Valley mines, and mineral exploration for the surrounding area. Services for

mining operations are reasonably available at Prince George, Vancouver, or

Edmonton.

Power transmission lines, rail, paved, and gravel roads are all adjacent to the Project

near the Yellowhead Highway. The Yellowhead Highway runs sub-parallel to the

North Thompson River. The community of Blue River has a municipal airport for light

aircraft and helicopter support.

The main line of the Canadian National Railway passes through the western part of the

property. Sidings currently exist at Lempriere (16.5 km north) and Blue River (23.7 km

south of the Upper Fir deposit). The flat area immediately north of Bone Creek may be

suitable for a siding and is 4.9 km from the Upper Fir deposit.

The BC Hydro 136,000 volt supply line for the North Thompson valley passes through

the west side of the property adjacent to the rail line. The 20 megawatt Bone Creek

run-of-river hydroelectricity project owned by Transalta Corp was commissioned in

June 2011, and is adjacent to the Project area.

Infrastructure requirements as detailed within the PEA for Project development are

discussed in Section 18 of this Report.

5.4 Physiography

The Project topography ranges from 700 m to 3,100 m elevation above sea level and

is located largely along the steep, west-facing slopes of the Monashee Mountains, to

the east of the North Thompson River.

The highest peak, Mt. Lempriere, is 3,183 m. Ice fields and glaciers dominate the

higher elevations on the property. Significant major tributaries feeding into the North

Thompson River in the area include Serpentine Creek, Pyramid Creek, Gum Creek,

Bone Creek, Hellroar Creek and Mud Creek.

Mountain slopes are typically covered by thick undergrowth consisting of grasses,

buck brush, devil’s club, and shrubs of willow, alder, rhododendron, huckleberry,





Project No.: 162230


currants, gooseberry, thimbleberry, and raspberry. White spruce is common in

replanted logging areas. Former trails and flat wet areas are typically overgrown by

dense alder and willow. Areas not subjected to recent logging are covered by dense

stands of hemlock, cedar, fir and white pine. Within the area, the tree line is at

approximately 2,000 m elevation. Except for the Paradise Lake, Felix, Howard Creek

and Gum Creek localities, all other carbonatites are below the tree line, and outcrop

exposure is generally poor.


The existing and planned infrastructure, availability of staff, the existing power, water,

and communications facilities, the methods whereby goods could be transported to

any proposed mine, and any planned modifications or supporting studies are

reasonably well-established. The requirements to establish such infrastructure are

reasonably well understood by Commerce. In the opinion of the QPs, the access,

physiography, local services, plus existing and planned infrastructure can support the

declaration of Mineral Resources.

It is expected that any future mining operations will be able to be conducted year-

round.





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6.0 HISTORY


discovery of vermiculite bearing carbonatite rock in 1949. A summary of exploration

activities on the Project is described below and summarized in Table 6-1.

6.1 Commerce Exploration

In 2000, Commerce acquired the Property, confirmed known tantalum mineralization at

the Fir and Verity carbonatites, and explored for new carbonatite deposits. During the

summer of 2002, the Upper Fir Carbonatite showing was discovered and defined by

core drilling between 2005 and 2008.

Additional work undertaken by Commerce included surface mapping, trenching, soil,

rock chip, grab and channel sampling, geophysics, metallurgical testing, bulk

sampling, and mineral resource estimation.

6.2 Commerce Mineral Resource Estimates

During 2001 and 2002, Commerce commissioned preliminary Mineral Resource

estimates on limited data for the Verity and Fir carbonatite deposits (McRae, 2001,

2002).

During 2009–2010, Commerce commissioned Caracle Creek International Consulting

Inc. (CCIC) to prepare a mineral resource estimate for the Upper Fir. This estimate

was based on the interpretation of 168 Upper Fir drill holes completed between during

2005 to 2008. As part of the estimate, the Verity and Fir Mineral Resource estimates

were audited. CCIC concluded the Verity and Fir estimates could not be verified, and

therefore should not be relied upon. An initial NI 43-101 compliant Mineral Resource

estimate for the Upper Fir tantalum and niobium rich carbonatite was completed in

early 2010 by CCIC (Stone and Selway, 2010).

During 2010, Commerce commissioned AMEC to complete a Preliminary Economic

Assessment (PEA). As part of the PEA, the Upper Fir Mineral Resources were

updated and initial Bone Creek Mineral Resources were established using drill holes

completed during 2005 to 2009 and new geological interpretations of the data. A

NI43-101 compliant Mineral Resource update for the Blue River project was completed

by AMEC in early 2011 (Chong and Postolski, 2011)





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Table 6-1: Blue River Exploration History Summary Year Company Exploration

1949–1951 Oliver E. French Staking and prospecting; discovered vermiculite-bearing carbonatite near

Blue River, discovered uranpyrochlore in dolomitic carbonatite

1952–1955 St. Eugene Optioned property, geological mapping, prospecting, stripping, trenching,

and sampling

1967-1968 Vestor Staking, reconnaissance surface mapping in the area south of Paradise

Lake

1976 J. Kruszewski Re-staked the area as the Verity and AR claims

1977–1978 J. Kruszewski / E.

Meyers Magnetometer and scintillometer surveys, trenching and sampling

1980 AMC Optioned property, discovery of Fir and Bone Creek carbonatites

1980–1982 AMC 3,954.2 m of NQ diamond drilling at Verity, Mill, Fir and Bone Creek

1989 Diegel et al. Government survey discovered two new carbonatite localities near

Serpentine Creek and Gum Creek

2000–Present Commerce Surface mapping, trenching, soil sampling, geophysics, diamond drilling,

metallurgical testing, bulk sampling

Abbreviations: St. Eugene = St. Eugene Mining Corporation Ltd.; Vestor = Vestor Exploration Ltd; AMC = Anschutz

(Canada) Mining Ltd.; Commerce = Commerce Resources Corp.





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7.0 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

British Columbia is divided into three discrete areas hosting carbonatites and alkaline

rocks:

• Eastern Area: the Foreland Belt, east of the Rocky Mountain Trench

• Central Area: the eastern edge of the Omineca Belt

• Western Area: the core of the Omineca Belt.

The age of emplacement for carbonatites and alkaline rocks of the Eastern and

Central Areas is Devonian–Mississippian (ca. 330–380 Ma). Some occurrences from

the core, or Western Area of the Omineca Belt might be older (ca. 570 to 770 Ma).

The Central Area carbonatite intrusions occur along the eastern edge of the Omineca

Belt. The carbonatites of the Omineca Belt commonly have high concentrations of

niobium but low rare earth element (REE) values. Known carbonatite complexes

include the Blue River and Mud Lake areas.

All of the alkaline and carbonatite complexes and their host rocks within the Omineca

Belt rocks were deformed and metamorphosed during the Jurassic-Cretaceous

Columbian Orogeny and have been subjected to upper amphibolite facies

metamorphism.

7.2 Project Geology

The Project is located in the Central Area within the north-eastern margin of the

Shuswap Metamorphic Complex. The area comprises polyfolded, metamorphosed

Late Proterozoic (ca. 700–550 Ma) supracrustal rocks and is bounded on the east and

west by steep, Eocene, west-side down normal faults in the southern Rocky Mountain

Trench to the east and the North Thompson valley to the west. The Malton gneissic

complex lies to the north.

The supracrustal rocks are part of a belt dominated by the Late Proterozoic Horsethief

Creek Group and the overlying Kaza Group. The belt is continuous from the northern

Selkirk Mountains in the southeast, through the Monashee Mountains, and into the

Cariboo Mountains in the northwest.

The Blue River carbonatites are hosted in the Mica Creek assemblage of the

Horsethief Creek Group (Figure 7-1).





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Figure 7-1: Local Geology Map

Note: Figure courtesy Dahrouge Consulting Ltd.





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7.2.1 Metasedimentary Rocks

Two units of the Mica Creek assemblage underlie much of the study area. The units

are at least 1,000 m thick and comprise the lower pelite unit, and the stratigraphically

overlying semipelite–amphibolite unit (Figure 7-2).

The Mica Creek metasedimentary rock types include biotite gneiss, muscovite–biotite

schist and gneiss, garnet–muscovite–biotite schist and gneiss, calc-silicate–biotite

gneiss, amphibolite, garnet amphibolite, and calc-amphibolite.

The high intensity of deformation precludes determination of tops in metasedimentary

rocks, thus relative ages of individual units are not clear. Layering in the gneiss is

interpreted as relict bedding, not metamorphic segregation.

7.2.2 Gneisses and Schists

Metamorphosed quartzo-feldspathic biotite gneiss is the most abundant lithology that

crops out at surface. Biotite gneiss is ubiquitous and is inter-layered with all other

lithologies on the property.

Outcrops are moderately weathered with characteristic 0.2 to >1 m thick layers of

uniform, massive, medium grained quartz–feldspar–-biotite ± muscovite divided by

recessive schistose bands or fine partings. Fresh surfaces have a uniform,

equigranular, salt and pepper texture of quartz, feldspar and biotite. Muscovite occurs

as thin schistose partings from trace to abundant amounts. Sub-one mm diameter red

garnet occurs in varying amounts. These units are interpreted to represent deformed

and moderately re-crystallized turbidite. Calc-silicate-bearing biotite gneiss has pale

green bands a few centimetres thick that may be related to microscopic traces of

actinolite ± diopside.

7.2.3 Amphibolites

The amphibolite units occur as lenses within all gneiss and schist units. They are

typically medium-grained, massive to moderately foliated amphibolites and can contain

red garnets (almandine) typically <1 cm in diameter. Plagioclase and hornblende

occur in varying proportions forming rocks ranging from tonalite to hornblendite

composition (<10% hornblende to >90% hornblende respectively). Weak mineral

lineations are present as observed by the alignment of hornblende. Rare banding is

observed at centimetre scale.





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Figure 7-2: Deposit Area Surface Geology Map






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Locally, calc-amphibolite units are distinguished by an increase in mineral grain size, a

strong contrasting black and white colour, local presence of garnets, and effervescent

reaction with dilute hydrochloric acid. The amphibolites units are interpreted as

metamorphosed mafic sills, dikes, and possibly subaqueous flows.

7.2.4 Intrusive Rocks

Ultramafic Rocks

Ultramafic rocks associated with the carbonatites include fine- to medium-grained

pyroxenites and cumulate pyroxene–hornblendites. The ultramafic units likely

represent a metamorphosed ultramafic intrusion associated with mafic volcanism

(amphibolite) roughly the same age as the intruded metasediments (Figure 7-2).

Carbonatite

The Blue River carbonatite is approximately 330 million years old (and possibly older)

based on U–Pb geochronology data. The carbonatite was emplaced as dikes or sills

into the metasedimentary rocks prior to the regional deformation and metamorphism

that occurred c.a. 200 Ma.

The carbonatite forms sill-like bodies with average thicknesses of 30 m, ranging

between 5 m to about 90 m thick, and with strike lengths ranging between 50 m to

1,100 m (Figures 7-3 to 7-6).

Bedding-parallel foliation in metasedimentary gneisses and the contacts of carbonatite

intrusions generally strike 335° and 155° with shallow to moderate northeast and

southeast dips.

Both dolomitic carbonatite and calcitic carbonatite occur at Blue River. Dolomitic

carbonatite is often referred to as magnesio-carbonatite, rauhaugite, or beforsite.

Coarse-grained, calcitic carbonatite is also often referred to as calcio-carbonatite or

sövite.

Dolomitic and calcitic carbonatite usually form separate bodies but can occur together

within single intrusions. At Blue River, dolomitic carbonatite typically makes up the

cores of the carbonatite bodies. Crosscutting or gradational relationships can be

observed from one variety of carbonatite into another.





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Figure 7-3: Drill Collar and Vertical Section Locations

Note: See Figure 7-4 for the Lower Road longitudinal section; see Figure 7-5 for section 5796737 N; see Figure 7-6 for

section 5796425 N. Bulk sample locations are noted at BS1, BS2 and BS3. Figure courtesy Dahrouge Consulting Ltd.





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Figure 7-4: Lower Road Longitudinal Section 352800 E

100 m

North South

Fenite

Upper FirCarbonatite

Bone Creek

Carbonatite

Undifferentiatedmetasediments

Ta2O5

Note: The figure illustrates the carbonatite geometry and its north-south geological continuity. View is to the east. Section influence is +/- 25 m.

Carbonatite = blue coloured domains; fenite = dashed green outlines; undifferentiated metasediments = non-filled area below topography. Drill hole

2 m composites are colour-coded for Ta2O5 grade in ppm.





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Figure 7-5: Geology Section 5796737 N

West East

Bone CreekCarbonatite


Amphibolite

Fenite

Undifferentiatedmetasediments 100 m

Ta2O5

Note: Illustrates the carbonatite geometry, its east-west geological continuity, and relationship between drilled thickness versus true thickness. View

is to the north. Section influence is +/- 25 m. Carbonatite = blue coloured domains; fenite = dashed green outlines; undifferentiated metasediments =

non-filled area below topography. Drill hole 2 m composites are colour coded for Ta2O5 grade in ppm.





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Figure 7-6: Geology Section 5796425 N

West East


Fenite

Fenite

Undifferentiatedmetasediments 100 m

(A)

(B)

(C)

Ta2O5

Note: Illustrates the carbonatite geometry, east-west geological continuity, folding, and relationship between drilled thickness versus true thickness.

View is to the north. Section influence is +/- 25 m. Carbonatite = blue coloured domains; fenite = dashed green outlines; undifferentiated

metasediments = non-filled area below topography. Drill hole 2 m composites are colour coded for Ta2O5 grade in ppm.





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Dolomitic and calcitic carbonatites are medium to coarse-grained and have secondary

tectonically-imposed textures. A cataclastic (porphyroclastic) texture is common in all

the carbonatites. Most exposures display layering defined by varying quantities of

accessory minerals. Accessory minerals include amphibole, pyroxene, phlogopite,

olivine, magnetite, apatite, pyrite/pyrrhotite, ilmenite, zircon, and various tantalum and

niobium bearing minerals.

Contacts between carbonatite and the host metasediments are typically sharp and

mantled by zones of metasomatized host rock, known as fenite.

At Blue River, the fenite rocks commonly, but not always, envelope the carbonatite

rocks and can extend up to 50 m from the carbonatite intrusions. The

metasedimentary host rocks are characterized by foliated calcite-richterite-biotite

(± apatite, ± vermiculite) rock. Of lesser importance are contact metasomatic veins

commonly less than 1 m thick that comprise amphibole-pyroxene (± vermiculite

± carbonate).

7.2.5 Pegmatite Dykes

Pegmatite dykes and pods up to 500 m long and 15 m thick crosscut all lithologies

throughout the property. At least some of the pegmatites are folded. Among the

pegmatites, two mineralogically-distinct types exist:

• Two-mica (± garnet, ± tourmaline) granitic pegmatites

• Syenitic pegmatites with minor biotite (± amphibole, ± pyroxene).

7.3 Structural Geology and Metamorphism

The structural geology is summarized largely from Kraft (2010), Ghent et al. (1977),

Simony et al. (1980), and Raeside and Simony (1983).

The style of structural deformation at the Project directly impacts the carbonatite

geometry. A deformation model was developed on behalf of Commerce by J. Kraft

during 2009 and early 2010. The structural deformation model was confirmed and

enhanced by field work completed by J. Kraft during July and September of 2010. The

following descriptions include the 2010 supporting observations and interpretations.

Regionally, three phases of compressional deformation have been mapped throughout

most of the region, from the northern Selkirk Mountains into the Cariboo Mountains. At

the Project, at least two additional deformation events are observed.





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The first deformation event (D1) produced large recumbent folds (F1) with limbs

approximately 50 km long and an associated early foliation (S1). Features from F1 are

not observed within the immediate deposit area.

The second deformation event (D2) is associated with peak, mid-amphibolite facies

metamorphism with an associated foliation (S2). In general the S1 and S2 foliations

can rarely be distinguished from one another in the field and are mapped as S1+S2.

D2 has created boudinage, or pinch and swell features attributed to competency

contrasts between rock type layers.

The third deformation event (D3) is characterized by centimetre to decametre scale

recumbent folds (F3) that deform the S2 foliation. Axial planar schistosity that deforms

S1+S2, within micaceous lithologies such as fenite, is known as S3. The style and

attitude of F3 folds are variable, but axial planes are generally southwest-dipping.

The fourth deformation event (D4) is characterized by inclined, southwest trending

folds that re-fold larger F3 folds in the deposit area. Open to tight upright folds with

east to southeast hinges occur sporadically. D4 is suggested to include thrust faulting

with top to the southwest vergence.

The fifth deformation event (D5) is described as a brittle extensional event

characterized by normal faults with slickensides, weak quartz-pyrite alteration of wall

rocks, and cross-cutting relationships with D4 structures.

Folding is observed in waste rocks adjacent to the carbonatite in outcrop and in drill

core, and is interpreted for carbonatite intercepts on large scale geological sections.

Within carbonatite, compressional deformation with weak southeast elongation is

suggested by zones with cataclastic to mylonitic foliation (Chudy, pers. comm.) and

weakly to moderately developed mineral lineations defined by amphiboles.

Carbonatite bodies are folded at metres to deposit scale, however their thickness and

massive (non-layered) nature makes observation of folding indicators within the

carbonatite comparatively difficult to observe in outcrop or drill core.

7.4 Geochronology

The geochronology is summarized largely from Pell (1994), Simonetti (2008), and

Gervais (2009).

A uranium–lead date of about 325 Ma was obtained from zircons from the Verity

carbonatite. A lead–lead date of 332.5 ± 5.7 Ma age was obtained from zircons for the





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Upper Fir carbonatite. A preliminary uranium–lead date of 328 ± 30 Ma was obtained

from zircons from the Mud Lake area carbonatite.

Zircons separated from syenite at Paradise Lake yielded a uranium–lead age of about

340 Ma and lead-lead ages of about 351 and 363 Ma. Ongoing research on U/Th/Pb

dating of zircons and monazites from the property shows a complex thermal history,

indicating that the age of emplacement of the Blue River area carbonatites may be

older than initial results have shown.

7.5 Carbonatites

The Upper Fir and Bone Creek carbonatites are assessed in the mining plan

discussed in Section 16 of this report. However, for Report completeness purposes,

all of the significant carbonatite deposits are described in this subsection.

7.5.1 Fir Carbonatite

Information summarized in this subsection on the Fir deposit is from the British

Columbia Geological Survey website.

The Fir showing is located 1.25 km north of the Bone Creek carbonatite. Carbonatite

consisting of dolomitic and lesser calcitic carbonatite occurs as sills within the quartz-

hornblende-mica schist of the Semipelite Amphibolite division of the Horsethief Creek

Group. Other lithologies include amphibole-biotite schist, biotite-muscovite gneiss and

amphibole-biotite-garnet gneiss.

The Fir carbonatite likely strikes 400 m in a northerly direction based on outcrop

exposures. A 2 m exposure of dolomitic carbonatite was located 400 m north of the

discovery outcrops. Dolomitic outcrops are coarsely crystalline and typically weather

white. Accessory minerals in the carbonatites include apatite, amphibole, olivine,

magnetite, pyrite, pyrrhotite, pyrochlore, and columbite. The dolomitic carbonatite is

almost devoid of biotite and magnetite. Three distinct textures were observed:

breccias composed of tightly-packed dolomite fragments within a finely crystalline

dolomite groundmass; a porphyritic texture with ghost dolomitic crystals in a fine-

grained matrix; and a massive texture with local banding of accessory minerals.

Tantalum and niobium mineralization in the Fir carbonatite occurs in the minerals

pyrochlore and columbite. The Fir carbonatite has the highest background niobium

and tantalum values of all carbonatites in the area. Tantalum averages greater than

0.015 per cent. Sampling of the discovery outcrops returned assays of 1.02% Nb2O5,

0.06% Ta2O5, and 6.31% P2O5. A sample from drill hole BC19 returned values of





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0.18% tantalum and 8.51% phosphate from 119.0–119.6 m. The anomalous value

may be the result of a pyrochlore–apatite band in the sample.

7.5.2 Verity Carbonatite

The Verity carbonatite is located about 40 km north of the community of Blue River.

The Verity carbonatite has the most varied stratigraphy of all the carbonatites in the

area.

The Verity carbonatite consists of banded dolomitic and calcitic carbonatite that locally

intrude each other. It occurs as a 15 m to 30 m thick sill within quartz-hornblende-mica

schist of the Horsethief Creek Group. It can be traced up the hillside for 800 m to the

east–northeast.

A tectonic breccia showing hairline fractures is common in the dolomitic carbonatite. A

banded texture caused by layering of the accessory minerals apatite, amphibole,

olivine, magnetite, vermiculite, biotite, pyrite, pyrrhotite, pyrochlore, columbite, and

zircon is common in calcitic carbonatite and less developed in the dolomitic

carbonatite. Coarse olivine and apatite in calcitic dolomite form bands 1 cm to 5 cm

thick. Magnetite occurs as discontinuous lenses in calcitic carbonatite layers; the

lenses can be much as 20 cm in diameter.

Tantalum and niobium mineralization in the Verity carbonatite occurs in the minerals

pyrochlore and columbite. The pyrochlore and columbite crystals occur as

octahedrons that can reach 4 cm diameter. Calcitic carbonatite at the Verity

occurrence also contains greater than 10.8% phosphate and has abundant apatite

relative to other nearby carbonatites at the Project. Rare earth elements are

interpreted to be hosted in fluorine-rich carbonate.

7.5.3 Exploration Targets

Geochemical sampling has outlined a number of potential exploration targets,

including:

• Upper Fir Extension target: strong tantalum anomalies on four adjacent soil lines

north–northwest of Bulk Sample Pit #2 indicate that the carbonatite subcrop likely

extends to the north, past the current drill coverage





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• Bone Creek Extension target: strong tantalum-in-soil anomalies on two widely-

spaced lines centered at UTM 5,797,000 N indicate a near surface carbonatite

body that is on strike with the Bone Creek carbonatite

• Fir Exploration target: strong tantalum-in-soil anomalies on widely-spaced lines

located north, south and above the known Fir showing indicate possible extensions

of the Fir carbonatite

• Mt. Cheadle Exploration target: a large diffuse tantalum-in-soil anomaly, with

several spikes, stretching over 2 km, is located north of Gum Creek and along

strike from the Upper Fir carbonatite

• 3050 Road target: Strong tantalum anomalies on soil lines to the north and east of

current drilling on the Upper Fir deposit near 3050 Road indicate another

carbonatite body may be located above the known extents of the deposit.

Target locations are indicated on Figure 7-7.

Soil sample results indicate the Upper Fir carbonatite has exploration potential directly

northward of known deposit extents. Additional resource definition drilling is

warranted.

Soil sample results indicate the Bone Creek and Fir carbonatites have additional

exploration potential along, and across, strike. Additional in-fill soil sampling is

warranted prior to diamond drilling to assess for potential connections with the Upper

Fir carbonatite.

7.6 Mineralogy

There are two principal and one minor niobium- or tantalum-bearing minerals known at

the Project.

The minerals, using generic end-member compositions, are:

• Ferrocolumbite: (Fe,Mn,Mg)(Nb,Ta)2O6

• Pyrochlore: (Ca,Na,U)2(Nb,Ti,Ta)2O6(OH,F)

• Fersmite: (Ca,Ce,Na)(Nb,Ta,Ti)2(O,OH,F)6.





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Figure 7-7: Exploration Target Location Plan






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7.6.1 Ferrocolumbite

Ferrocolumbite occurs predominantly in medium to coarse-grained, granoblastic

dolomitic carbonatites which typically form thin intervals (<6 m) or occur at margins of

thicker intervals of carbonatite. Ferrocolumbite forms subhedral to anhedral,

sometimes strongly poikilitic, individual grains or agglomerates of grains.

Mineral liberation analyses show that the majority (~80%) of liberated grains are less

than 110 µm in diameter. Locally, individual grains and agglomerates of

ferrocolumbite may exceed 2 cm in diameter.

Ferrocolumbite grains from marginal zones may contain large amounts of tiny

inclusions such as thorite (Th-silicate), monazite (La, Ce-phosphate) and pyrochlore.

Ferrocolumbite may also occur sporadically as inclusions in apatite and amphibole. It

is often associated with layers and micro-veins of apatite that fill the interstices

between anhedral ferroan-dolomite grains.

7.6.2 Pyrochlore

Pyrochlore occurs predominantly in the fine-grained and porphyroblastic dolomitic

carbonatite which is commonly developed in the central portions of carbonatite

intervals greater than 10 m thick. Such zones are less abundant or absent in thinner

carbonatite intersections. Pyrochlore is the only tantalum mineral in the calcitic

carbonatites that occurs in accessory amounts. Black and brownish-yellow coloured

varieties of pyrochlore are present.

The majority of the pyrochlore occurs as liberated grains in the dolomitic matrix. The

vast majority (~ 85%) of pyrochlore forms subhedral to anhedral, rounded grains less

than 200 µm in diameter. There are local larger grains and agglomerates, as well as

accumulations or veins less than a few tens of centimetres in width with high

pyrochlore abundance. This style of mineralization can result in high tantalum values

(> 450 ppm Ta).

Pyrochlore also occurs as inclusions in amphiboles (richterite), fluorapatite, and in

ferrocolumbite. In some rare cases the pyrochlore grains can be coated with a thin

film of pyrrhotite or pyrite.

7.6.3 Fersmite

Fersmite occurs as anhedral inclusions in apatite and is considered a minor economic

mineral at the Project.





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7.6.4 Fenite Mineralization

Mineralization in the fenite is dominantly ferrocolumbite, concentrated in apatite-rich

layers. Ilmenite, with ferrocolumbite inclusions, appears to be a subdominant source

of both niobium and tantalum. Niobium and tantalum grades within fenite at Blue River

are considered to be sub-economic, but locally fenite may provide grade-bearing

mining-dilution material.

7.6.5 Mineral Zoning

Mineral zoning, or distribution, of ferrocolumbite and pyrochlore within the carbonatites

is not clear due to the variable thicknesses and polyfolded geometry of the carbonatite.

Further work is required to improve the understanding of the mineral zoning and to

locate potential material types required to support metallurgical testwork.

7.7 Comments on Section 7

In the opinion of the QPs, knowledge of the deposit settings, lithologies, and structural

and alteration controls on mineralization is sufficient to support Mineral Resource

estimation.

In the opinion of the QPs, the mineralization style of the Project deposit is sufficiently

well understood to support Mineral Resource estimation.

Prospects and targets are at an earlier stage of exploration, and their lithologies,

structural, and alteration controls on mineralization are at present not well understood

and hence more work is required to support estimation of Mineral Resources.





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8.0 DEPOSIT TYPES The mineralization identified to date at the Project is consistent with magmatic,

carbonatite associated deposits. Carbonatite associated deposits are classified as

either magmatic, replacement, or residual. Global examples of magmatic carbonatite

complexes or deposits include Oka, Niobec, and St. Honore (Quebec), Kovdor

(Russia), Iron Hill (Colorado) and Gardiner (Greenland) (Mitchell, 2010). Examples of

replacement carbonatite deposits are Rock Canyon (B.C.), Bayan Obo (China), and

Palabora (South Africa). Araxa and Catalao (Brazil) are classified as residual

carbonatite deposits due to the degree of lateritic weathering. Carbonatites are the

main source of niobium +/- tantalum, and important sources of rare earth elements.

Magmatic carbonatite deposits have the following common features (Birkett and

Simandl, 1998).

• Commodities: niobium, tantalum, rare earth elements, phosphate, vermiculite,

copper, titanium, strontium, fluorine, thorium, uranium, magnetite.

• Geological setting: Carbonatites intrude all types of rocks and are emplaced at a

variety of depths. Carbonatites occur mainly in a continental environment, rarely in

oceanic environments (Canary Islands) and are generally related to large-scale,

intra-plate fractures, grabens or rifts that correlate with periods of extension and

may be associated with broad zones of uplift.

• Age of Mineralization: Carbonatite intrusions are early Precambrian to Recent in

age; they appear to be increasingly abundant with decreasing age. In British

Columbia, carbonatites are mostly upper Devonian, Mississippian or Eocambrian

in age.

• Host Rocks: Host rocks are varied, including calcite carbonatite (sövite), dolomite

carbonatite (beforsite), ferroan or ankeritic calcite-rich carbonatite

(ferrocarbonatite), magnetite-olivine-apatite ± phlogopite rock, nephelinite, syenite,

pyroxenite, peridotite and phonolite. Carbonatite lava flows and pyroclastic rocks

are not known to contain economic mineralization. Country rocks are of various

types and metamorphic grades.

• Deposit Form: Carbonatites are small, pipe-like bodies, dikes, sills, small plugs or

irregular masses. The typical pipe-like bodies have sub-circular or elliptical cross

sections and are up to 3-4 km in diameter. Magmatic mineralization within pipe-

like carbonatites is commonly found in crescent-shaped and steeply-dipping zones.

Metasomatic mineralization occurs as irregular forms or veins. Residual and other

weathering-related deposits are controlled by topography, depth of weathering and

drainage development.





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• Deposit Mineralogy:

Magmatic: bastnaesite, pyrochlore, columbite, apatite, anatase, zircon,

baddeleyite, magnetite, monazite, parisite, fersmite.

Replacement/Veins: fluorite, vermiculite, bornite, chalcopyrite and other sulphides,

hematite.

Residual: anatase, pyrochlore and apatite, locally crandallite-group minerals

containing rare earth elements.

• Gangue Mineralogy: Calcite, dolomite, siderite, ferroan calcite, ankerite, hematite,

biotite, titanite, olivine, quartz.

• Alteration: A fenite halo (alkali metasomatized country rocks) commonly surrounds

carbonatite intrusions; alteration mineralogy depends largely on the composition of

the host rock. Most fenites are zones of desilicification with addition of Fe3+, Na

and K.

• Ore Controls: Intrusive form and cooling history control primary igneous deposits

(fractional crystallization). Tectonic and local structural controls influence the

forms of metasomatic mineralization. The depth of weathering and drainage

patterns control residual pyrochlore and apatite deposits, and vermiculite deposits.

Many features of the mineralization identified within the Project to date are analogous

to magmatic carbonatite deposits, in particular the Oka (Husereau Hill) and St. Honore

deposits, Quebec.

Key features of the Blue River deposits supporting a magmatic carbonatite model are:

• Commodities: niobium and tantalum

• Geological Setting: occurs along the eastern portion of the Omineca Crystalline

Belt and hence its tectonic setting is along a large scale zone with associated uplift

• Age of Mineralization: data yields results of about 330 Ma which is consistent with

other British Columbia carbonatite deposits

• Host Rocks: dolomite and calcite-rich carbonatite intrusion rocks

• Deposit Form: the Blue River carbonatites occur as sills and dykes

• Deposit Mineralogy: ferrocolumbite and pyrochlore

• Gangue Mineralogy: dolomite, calcite, amphibole (richterite), quartz, pyroxene,

phlogopite, olivine, magnetite, apatite, pyrite/pyrrhotite, ilmenite, and zircon

• Alteration: Fenite halos occur around most carbonatites at Blue River





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• Ore Controls: The Blue River carbonatites have been deformed by multiple episodes

of folding and faulting. The internal cooling history of the deposit is not clear.


In the opinion of the QPs,

• A polyfolded sill-like carbonatite model suitably describes the Blue River deposits

• The deposit concepts being applied as the basis for exploration planning at the

Project are reasonable.





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9.0 EXPLORATION


discovery of vermiculite-bearing carbonatite rock in 1949.

Commerce acquired the property in 2000 and initiated exploration for new carbonatite

deposits which culminated in the discovery and delineation of the Upper Fir and Bone

Creek carbonatites.

9.1 Grids and Surveys

All surveys to date are in UTM NAD83 Zone 11 coordinates.

In 2007, orthophotography and Lidar surveys were flown to create a 1:2,000 base map

of the Upper Fir area. A topography map with 2 m contour intervals was created by

Eagle Mapping Ltd.

9.2 Geological Mapping

Geological and structural mapping has been completed at 1:2,500 scale on a

continuous basis since 2006. The mapping area coverage is between Bone and Gum

Creeks; and from the North Thompson River to the ridge top located about three km

east on the slope that is known locally as either Fir or Cedar Mountain. The mapping

area includes the Fir, Upper Fir and Gum carbonatites plus the nearby Hodgie Zone.

Mapping was used to determine the outcrop of carbonatite and provide geological and

structural data.

9.3 Geochemical Sampling

9.3.1 Stream Sediment Sampling

Reconnaissance stream sediment sampling was completed during 2001 to 2003, and

2006 to 2007. During 2008, 531 stream sediment samples were collected and

analysed for the streams throughout the entire property. The key exploration

pathfinder elements at Blue River are tantalum, niobium, and rare earth elements.

Detailed sample analysis using microscopic mineral characterization was utilized,

focusing on identifying pyrochlore, apatite, richterite, and monazite as pathfinder

minerals.





Project No.: 162230


During the 2009 field season, a total of 20 stream pan concentrate samples were

taken in the Fir and Mud Creek areas to follow up on creek-mouth areas inaccessible

during the 2008 field season. Samples were analyzed at Acme Laboratories in

Vancouver B.C. (Acme) primarily for tantalum, niobium, rare earth elements,

phosphate, and carbonate using Acme’s 4B02 package (lithium metaborate fusion-

ICP-MS technique).

Several samples anomalous in tantalum and niobium indicate that the Fir carbonatite

likely extends further south.

9.3.2 Soil Sampling

Soil sampling has proven the best way to follow up stream pan concentrate sampling

in the Blue River area as the niobium–tantalum-bearing ferrocolumbite and pyrochlore

are residual in soils. The key exploration pathfinder elements from soil sampling are

tantalum and niobium. During 2002, follow-up on an anomalous stream sediment

sample led to the discovery of the Upper Fir carbonatite.

Reconnaissance soil sampling was completed during 2001 to 2003, and 2006 to 2007.

During the 2008 season, 4,181 soil samples were collected from several area grids.

Sample grids typically have 200 m spaced lines and samples are taken at 25 m

intervals. Soil sampling in 2009 followed up on 2008 stream or soil sampling

anomalies. Sample grids were laid out at Switch Creek, and in the area of the Fir

deposit to test lateral extensions of known carbonatite showings. The Hellroar soil

sampling grid is located approximately 8 km south of the area of the Mineral Resource

estimate, near Hellroar Creek. Sampling on the Hellroar soil grid occurred during 2008

and 2009 to follow up on tantalum anomalies identified from stream sediment samples.

A total of 1,694 samples was taken by Dahrouge and analyzed by Acme using Acme’s

4B02 package (lithium metaborate fusion-ICP-MS technique).

9.3.3 Rock Chip, Grab, and Channel Sampling

Rock samples from various new and known localities were taken during 2009

prospecting to test for or verify the presence and abundance of tantalum–niobium

mineralization. A total of 100 in situ bedrock grab, chip, and channel samples were

taken at the Paradise, Roadside, Howard Creek, Gum Creek, and Mud Creek area

carbonatites.

Five continuous chip samples from the upper 5 m of the Roadside carbonatite

averaged 1,373 ppm Nb and 101 ppm Ta. At the Paradise carbonatite, the best





Project No.: 162230


channel sample assay result was 82 ppm Ta and 375 ppm Nb for one sample over a

1 m continuous interval.

Southeast of Mud Creek, a new carbonatite called the “RD” occurrence was

discovered during the 2009 regional prospecting program. This carbonatite is

approximately on strike with the Mud Creek carbonatite and may be part of the same

system. Grab samples from the RD carbonatite ranged up to a maximum of 118 ppm

Ta and 4,703 ppm Nb.

9.4 Bulk Sampling

A bulk sample program was undertaken in the fall of 2008 by Dahrouge as part of on-

going evaluation of the Upper Fir carbonatite. Approximately 2,000 t from a 10,000 t

permitted volume were extracted from three bulk sample pits (BS-1, BS-2, and BS-3;

refer to locations shown in Figure 7-3) and placed into 75 t to 150 t stockpiles that

were comprised largely of -50 cm carbonatite material. The stockpiles were stored on

a lined, well-drained pad at the Project site for later metallurgical testing.

For each pit area, geological mapping was completed along with sampling of blast-

hole material and channel samples of the bench walls. Both primary igneous and

metamorphic textures and structures were exposed in the sample pits. Microscopic

examination of oxide phases in the bulk sample material indicated that pyrochlore was

the dominant mineral in pit BS-1 with the exception of benches at the upper and lower

contacts.

Pit BS-1 was excavated in dominantly fine- to medium-grained, granular, apatite-

bearing, dolomitic carbonatite. Pits BS-2 and BS-3 were excavated in dominantly light-

grey, coarse-grained, porphyroclastic, apatite-bearing, dolomitic carbonatite.

Crosscutting veins of dark green actinolite–calcite–diopside that are as much as 20 cm

wide are common. Contacts in each pit are marked by approximately 1 m of fenite,

with contorted layers of dolomitic carbonatite up to 10 cm thick.

Material from the 2008 bulk sampling program appear to provide a sufficient range of

tantalum and niobium grades to represent the Upper Fir carbonatite mineralization for

initial metallurgical testing.

Both pits, BS-1 and BS-2, were stabilized. The bulk sampling permit expired on 31

December 2009 and has not been renewed.





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9.5 Research Programs

Doctoral and post-doctoral studies on the geology, petrology and microscopy of the

carbonatite mineralization at Blue River are underway at the University of British

Columbia.


In the opinion of the QPs, the exploration programs completed to date are appropriate

to the style of the deposits and prospects within the Project. The exploration and

research work supports the genetic and affinity interpretations.

The project retains significant exploration potential for carbonatite-hosted tantalum–

niobium mineralization.





Project No.: 162230


10.0 DRILLING Core (diamond) drilling is the most extensively used exploration tool at Blue River. As

of 30 June 2010, there were a total of 215 core drill holes within the Upper Fir, Bone

Creek and Fir (Lower) carbonatites comprising 41,115 m of HQ (63.5 mm) and NQ

(47.6 mm) diameter drill holes.

Table 10-1 lists the core drilling campaigns at the Upper Fir and Bone Creek deposits.

Core drilling by Commerce commenced in 2005 and continues to the present. Drill

locations were included in Figure 7-3.

Of the 215 core drill holes, 183 drill holes totalling 37,446 m of HQ diameter core, and

8,218 samples are used to support the Mineral Resource estimate. Of the 215 core

holes, Commerce drilled 11 holes in the Fir carbonatite area which are not part of this

Mineral Resource update. Of the 215 core holes, 21 legacy holes were drilled by

operators prior to Commerce’s involvement in the project. No sample intervals are

present in the database for the 21 legacy holes and hence these holes have not been

used to estimate grades in this Mineral Resource update, but they were used to

interpret geology.

An additional 54 core holes, totalling 12,949 m of HQ drill core were drilled in 2010 and

34 holes totalling 8715 m of HQ drill core were drilled in 2011. The 2010 and 2011

holes were not used in the preparation of the 2010 Mineral Resource estimate used to

support the 2011 PEA.

Preliminary results from 54 holes drilled in 2010 were reviewed by AMEC after

completion of the 2010 Mineral Resource estimate update. Core and logged lithology

information from these holes was available for review. The carbonatite intercepts from

these holes were compared on screen against the 3D carbonatite model used in the

resource estimation and generally support the geologic interpretation. Some

discrepancies were observed which warrant local re-interpretation. Some holes

expand carbonatite volume and some holes reduce it where the carbonatite/wall rock

contact is off by 5 m to 10 m. Laboratory analysis of the 2010 drill-core samples has

been completed and quality control results are being reviewed.

Preliminary results from 30 holes drilled in 2011 were reviewed by AMEC in

September 2011. Core and logged lithology information from these holes was

available for review. The logged lithology was examined on lithology cross sections

prepared by Commerce. The drill holes intersected carbonatite where expected and

generally support the geologic interpretation. Laboratory analysis of the 2011 drill core

samples is in progress.





Project No.: 162230


Drill-collar locations are shown in Figure 10-1. The locations of the 2011 drill collars

are provisional, as some drill-collar locations may have been moved in response to

results from earlier 2011 holes.

Table 10-2 lists the three bulk sample pits (BS-1, BS-2, and BS-3) and four trenches

(TR0, 0A, 1, and 1A) which have been sampled. These were only used for geology

interpretation and domain modeling, not grade interpolation.

10.1 Core Drilling Strategy

The holes are collared on three primary drill roads (Upper, Middle, and Lower roads)

that are oriented sub-parallel to the Upper Fir carbonatite along the hillside. Set-ups

are spaced about 50 m apart along drill roads. The majority of the known portions of

the Upper Fir deposit are defined on 50 m centres. The at-depth Bone Creek

carbonatite has only been intersected by a limited number of drill holes.

The drill hole orientations appear to be oriented approximately sub-perpendicular to

the carbonatites. The relationship between sample length and true thickness varies

with the dip of holes. True thicknesses are slightly less than drilled intercepts.

10.1.1 Core Sizes

Core holes are typically HQ diameter (96 mm) producing core with a diameter of

(63.5 mm). Hole-diameter reductions due to poor ground conditions generally are not

an issue at the Project. Drill-hole orientations have typical azimuths of vertical, 090°,

or 270° and inclinations that range from vertical to -60°. Drill-hole depths range from a

minimum of 32 m and a maximum of 388 m, averaging about 200 m.

10.1.2 Collar Surveys

The drill-hole collars are spotted in the field with a hand-held GPS and oriented with a

Brunton compass.

During 2008 and 2009, collars were surveyed using a laser theodolite system by Steve

Mosdell of Align Surveys of Louis Creek, B.C.

In 2010, McElhanney Associates of Vancouver undertook a differential GPS survey of

all 2010 drill collars, as well as all historic collars still visible, including all Steve

Mosdell’s work that could be verified, and all roads. Six reference markers were also

placed on the property to support future surveys. Commerce advised AMEC that

some earlier drill collar locations have been lost due to subsequent construction

activities.





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Table 10-1: Drill Campaign Summary Category Deposit Operator Year # Holes Series Type # Metres # Samples % Samples

Resource Bone

Creek

Commerce 2005 4 CF05-01 to CF05-04 HQ 300 14

0%

Resource Upper Fir Commerce 2005 4 CF0505 to CF0508 HQ 505 44 1%

Resource Upper Fir Commerce 2006 17 CF0601 to CF0617 HQ 3,021 1,139 14%

Resource Upper Fir Commerce 2007 18 F0718 to F0735 HQ 4,310 1,053 13%

Resource Upper Fir Commerce 2008 118 F08-36 to F08-153 HQ 23,723 5,126 62%

Resource Upper Fir Commerce 2009 22 F09-154 to F09-176 HQ 5,587 842 10%

Resource Subtotal 183 37,446 8,218 100%

Historical Bone

Creek

AMC 1980-1981 17 BC-1 to BC-17 NQ 697 na-

Historical Fir AMC 1981 4 BC1-18 to BC-21 NQ 829 na-

Target Fir

(twins)

Commerce 2001-2002 11 F-01 to F-11 HQ 2,144 na-

Target Subtotal 32 3,670

Total Drilling 215 41,115

Notes:

Abbreviations: AMC = Anschutz Mining Corp.

The Commerce 2010 campaign comprises 54 HQ diameter drill holes totalling 12,949 m. These holes were not completed during the database audit

or block modeling for this Mineral Resource estimate. Assay data for the 2010 sampling are still pending.

Table 10-2: Upper Fir Deposit Trench and Bulk Samples

Category Deposit Operator Year

Number Series Type

#

(m)

Metallurgy Upper Fir Commerce 3 BS01 to BS03 Bulk Sample 138

Exploration Upper Fir Commerce 4 TR0, 0A, 1, 1A Trench-Chip 73





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10.1.3 Down Hole Surveys

Vertical holes were generally not surveyed down hole. The dip and azimuth of inclined

drill holes were typically tested at three points in each hole using Flexit Single Shot

down-hole orientation tools. The instrumentation records magnetic inclination,

azimuth, temperature and magnetic susceptibility at each survey depth. The Flexit

instruments were calibrated at the start of each field season from 2005 to 2010.

10.1.4 Oriented Drill Core

Six drill holes consisting of 1,271 m of HQ diameter oriented core were completed in

2010. The drilling was performed for geotechnical purposes.

10.1.5 Core Handling

Core was obtained using wire-line methods and was washed prior to placement in

wooden core trays. Core trays were placed near the core barrel so that the core was

placed in the tray in the same orientation as it came out of the barrel. Rubble, which

was rarely encountered, was piled to about the length of the whole core that its volume

would represent. Trays were marked with drill-hole name and box number. The end

of every run is marked by a wooden block depth marker.

The core trays are transported by pick-up truck down to the core logging facility at the

community of Blue River, B.C.

10.1.6 Core Recovery

Core recovery was determined prior to sampling. Typically, recovery measurements

were completed before detailed logging was initiated. Standard core recovery forms

were usually completed for each hole by the field assistant or geologist.

Core recovery is very good within the waste and carbonatite rocks (typically >95%).

The only area that may have core recovery issues would be within the fenite rocks

located in the immediate hanging wall to the carbonatite.





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10.2 Drill Intercepts

Table 10-3 contains examples of the types of drill intercepts that have been returned

for the Blue River deposit areas. Typical drill-hole orientations are indicated on the

cross-sections included in Section 7 of this Report. Due to the dip of the carbonatite,

drilled thicknesses reported in the table are slightly longer than true thicknesses.


In the opinion of the QPs, the quantity and quality of the lithological, geotechnical,

collar location and down-hole survey data collected in the exploration and infill drill

programs completed by Commerce are sufficient to support Mineral Resource

estimation.

• Core logging meets industry standards for tantalum and niobium exploration within

a carbonatite setting

• Collar surveys have been performed using industry-standard instrumentation

• Down-hole surveys were performed using industry-standard instrumentation

• Recovery data from core drill programs are acceptable

• Drill-hole orientations are generally appropriate for the mineralization style, and

have been drilled at orientations that are optimal to the orientation of mineralization

for the bulk of the deposit area

• Drill-hole intercepts appropriately reflect the nature of the tantalum and niobium

mineralization. Sampling is representative of the grades in the deposits, reflecting

areas of higher and lower tantalum and niobium grades

• No material factors were identified with the data collection from the drill programs

that could affect Mineral Resource estimation.





Project No.: 162230


Figure 10-1: Drill Collar Plan

Note: Figure courtesy Commerce Resources Corp.





Project No.: 162230


Table 10-3: Example Drill Hole Intercept Summary Table

Drill Hole Easting Northing Elevation Azimuth Dip From To Length (m) Ta (ppm) Nb (ppm)

F0718 352817.7 5796722 1207.19 0 -90 34.5 76.4 41.9 158 2,488

F0718

113.5 141.7 28.2 153 1,266

F0718

210.9 215.9 5 199 1,396

F0718

218.4 221.5 3.1 148 1,173

F0719 352818.7 5796722 1207.19 94.8 -60 34.8 37.3 2.5 182 3,528

F0719

39.8 105 65.2 214 2,258

F0719

112.1 138.3 26.2 174 1,472

F0719

231.8 236.8 5 126 940

F0719

239.3 240.8 1.5 136 1,750

F0720 352816.7 5796722 1207.19 272.8 -60 38.4 86.4 48 120 2,032

F0720

146.7 158.5 11.8 129 1,456

F0720

244.1 248.8 4.7 150 1,240

F0727 352810.5 5796442 1214.98 0 -90 16.4 17.5 1.1 14 906

F0727

109.3 142.1 32.8 211 2,255

F0727

222.6 225.8 3.2 167 945

F0728 352811.4 5796442 1214.32 91.8 -60 27.3 30 2.7 13 745

F0728

93 150.4 57.4 188 1,417

F0729 352813.5 5796442 1214.14 271.8 -60 13.7 17.3 3.6 32 1,243

F0729

132.7 137.9 5.2 141 1,293

F0730 352905.3 5796762 1240.53 0 -90 75.1 149.5 74.4 163 1,583

F0731 352906.2 5796762 1240.4 91.8 -60 79.8 173.2 93.4 172 1,518

F08-064 352782.3 5796404 1211.64 0 -90 95.7 130.4 34.7 172 1,754





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11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

11.1 Sampling Methods

Samples were collected from an area approximately 1,600 m north–south by 1,000 m

east-west. Sampling is from a combination of vertical and inclined holes drilled from

common collar locations. This results in a drill hole or sample spacing which increases

with depth. Average spacing between drill-hole intercepts in the Mineral Resource

area is 50 m.

Core sampling method and approach has been consistent through the 2005 to 2010

drill programs. Core was boxed onsite and delivered each day to a core facility in Blue

River where the core was logged and sawn. Core logging involved both geotechnical

and geological information. Geotechnical logging involved measuring core recovery

per run, rock quality designation (RQD), fracture roughness and orientation. Core

recovery and RQD were generally good for most drill core, with typically greater than

95% recovery. The geological logging included observations of colour, lithology,

texture, structure, mineralization, and alteration. All drill core was photographed prior

to splitting.

The sampling procedure used to collect core at Blue River is as follows:

• The entire carbonatite intersection and shoulder samples on each side of the

intersection are sampled

• Samples intervals, generally 1 m in length, are marked on the core by a geologist

• Sample intervals are assigned a unique sample number

• The geological contacts are generally respected

• Specific gravity measurements for the carbonatite are collected approximately

once every 3 m

• Carbonatite samples are checked at regular intervals with a GR-130 miniSPEC

gamma ray spectrometer for the presence of U and Th

• Core is sawn in half by diamond saw

• Half of the core is sent for analysis

• Half of the core is retained for reference or further sampling.





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11.2 Metallurgical Sampling

The bulk sampling program conducted in 2008 to provide metallurgical samples is

discussed in Section 9.3. Additional information on the metallurgical sampling is

included in Section 13.1.

Metallurgical samples were collected from bulk sample material originating from BS-2

(approximately 173 t) during January 2009 and from BS-1 (approximately 6 t) during

November 2009. BS-2 samples were ferrocolumbite dominant and selected to best

average the tantalum-niobium grades for the carbonatite. BS-1 samples were selected

to best reflect pyrochlore dominant mineralization.

The two bulk samples were crushed to a particle size of <1 inch diameter. After

crushing, each group of samples was homogenized separately by a standard coning

and quartering procedure. The well-mixed samples were bagged into one tonne bags

and put into storage. One tonne of each sample was delivered to MetSolve Laboratory

(Metsolve) in Burnaby in BC to air dry and further reduce the size to -10 mesh for

bench testing.

11.3 Density Determinations

Commerce regularly collected density measurements using a water displacement

method. Five to 10 cm pieces of whole core were weighed with a balance beam scale

and then were placed into a beaker of water. The volume of water displaced was

measured with a clear plastic ruler to approximate the volume of the core. AMEC

recommended a check on the density data and additional measurements for waste

rock types. A water immersion specific gravity (SG) method was used in addition to

the water displacement method. No check samples were completed.

A total of 425 pieces of whole and half-sawn HQ diameter drill core measuring at least

10 to 20 cm long were checked. The samples chosen covered the spatial and time

aspects of the different Commerce drill campaigns.

Thirty-nine of the 425 samples were sent to Met-Solve Laboratories Inc. of Burnaby,

B.C. for measurement of SG by a wax coated water immersion method to assess for

possible water porosity. Original water displacement, check water displacement, and

check water immersion measurements were compared to the 39 MetSolve wax coated

water immersion SG measurements, with the following results:

• Original water displacement density measurements had poor correlation

• Check water displacement density measurements had adequate correlation





Project No.: 162230


• Check water immersion SG measurements had excellent correlation.

Based on this, the 425 water immersion check measurements were used to calculate

SG values for each major rock type at Blue River. The means are used as constants

to support the Mineral Resource estimate (Table 11-1).

11.4 Analytical and Test Laboratories

Acme Analytical Laboratories (Acme) in Vancouver were the primary laboratory for the

2005–2008 drill programs. Acme is independent of Commerce and is a recognized

analytical laboratory. Acme has implemented a quality system compliant with the

International Standards Organization (ISO) 9001 Model for Quality Assurance and

ISO/IEC 17025 General Requirements for the Competence of Testing and Calibration

Laboratories. The Vancouver hub laboratory is working toward ISO 17025:2005

accreditation, and Acme notes on its website that the laboratory is expected to

complete the accreditation process within the next year. Acme’s website notes:

Samples submitted are analyzed with the strictest quality control. Blanks (analytical

and method), duplicates and standard reference materials inserted in the sequences of

client samples provide a measure of background noise, accuracy and precision.

QA/QC protocol incorporates a granite or quartz sample-prep blank(s) carried through

all stages of preparation and analysis as the first sample(s) in the job. Typically an

analytical batch will be comprised of 34-36 client samples, a pulp duplicate to monitor

analytical precision, a -10 mesh reject duplicate to monitor sub-sampling variation (drill

core only), a reagent blank to measure background and an aliquot of Certified

Reference Material (CRM) or Inhouse Reference Material to monitor accuracy. In the

absence of suitable CRMs Inhouse Reference Materials are prepared and certified

against internationally certified reference materials such as CANMET and USGS

standards where possible and will be externally verified at a minimum of 3 other

commercial laboratories. Using these inserted quality control samples each analytical

batch and complete job is rigorously reviewed and validated prior to release.

PRA/Inspectorate Laboratories (Inspectorate) in Richmond, BC, was the primary

laboratory for sample preparation of the 2009 core drill program. Inspectorate is

independent of Commerce and is a recognized analytical laboratory that holds

ISO9001:2000 accreditation.

Global Discovery Laboratory (Global) in Vancouver, BC, was initially used for the 2009

drill program. Global was independent of Commerce and was a recognized analytical

laboratory. Global was subsequently purchased by Acme during the second quarter of

2009.





Project No.: 162230


Table 11-1: Specific Gravity Measurements for Blue River Rock Types

Domain Count Min Max Mean StDev CV

Gneiss 81 2.69 3.16 2.82 0.09 0.03

Amphibolite 30 2.93 3.19 3.05 0.07 0.02

Fenite+CalcAmphibolite 55 2.87 3.16 2.97 0.05 0.02

Carbonatite 168 2.85 3.24 3.01 0.08 0.03

Pegmatite 44 2.57 2.68 2.62 0.02 0.01

Other 29 2.81 3.20 3.03 0.10 0.03

Fault Zone (mixed) 17 2.54 3.06 2.85 0.15 0.05

Total 424 2.54 3.24 2.92 0.15 0.05

11.5 Sample Preparation and Analysis

Between 2005 and 2008 sawn core samples were shipped to Acme where the entire

sample was crushed in a jaw crusher to 70% passing 10 mesh (2 mm) and a 250 g

riffle split sample of the crushed material was pulverized in a ring-and-puck mill to 85%

passing -200 mesh (75 µm).

Split core samples from the 2009 drill program were shipped to PRA/Inspectorate

Laboratories in Richmond, B.C. where the entire sample was crushed to 80% passing

10 mesh and a 300 g split of the crushed material was pulverized to 100% passing -

200 mesh.

Between 2005 and 2008 primary samples were analyzed at Acme by packages 4A

and 4B. Package 4A allows reporting of total abundances of the major oxides and

several minor elements using a 0.2 g sample analyzed by inductively-coupled plasma

(ICP)-emission spectrometry following a lithium metaborate/tetraborate fusion and

dilute nitric digestion. Package 4B provided reporting of rare earth and refractory

elements determined by ICP mass spectrometry (MS) following a lithium

metaborate/tetraborate fusion and nitric acid digestion of a 0.2 g pulp. Primary

samples from the 2009 drill program were initially analyzed at Global Discovery

Laboratory in Vancouver and later at Acme by packages 8X, 4A, and 4B. Package 8X

is an X-ray fluorescence (XRF) analysis following a lithium metaborate fusion.

11.6 Quality Assurance and Quality Control

Quality control procedures used by Commerce to monitor laboratory results have

evolved over the life of the project. Between 2005 and 2007 there was minimal

insertion of blank, duplicate, or standard reference material (SRM) control samples.





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During this period, analysis of several pulp check samples was completed at six

different laboratories. In 2008 the control sample insertion rate was increased to an

average of 3% for each of blanks, quarter core field duplicates, and SRM control

samples. In 2009 control sample insertion rates were increased to an average of 5%

per control sample type and included pulp duplicates.

11.6.1 Assessment of Accuracy with SRM Control Samples

In 2005 a SRM control sample, BR-01, was prepared for Commerce by Acme, using

sample material from the nearby Verity Carbonatite. This SRM was inserted by Acme

into primary sample batches submitted between 2005 and 2008. Control sample

results were generally within ±5% of the best value. Less than 5% of the results

exceeded ±2 standard deviations

In 2008, fifteen SRMs were prepared for Commerce by Process Research Associates

(PRA) by pulverizing core samples from within the resource area to 100% passing

-200 mesh. Three samples of each standard were sent to six laboratories, for Ta, and

seven laboratories for Nb analysis. Analytical procedures included XRF/fusion and

XRF/pellet, ICP-AES, ICP-MS, ICP-M, and AA.

Results were compiled by PRA. The results indicate that, with the exception of a few

higher grade Nb samples, the Blue River SRMs do not achieve typical tolerance

thresholds for certified standards. This may be in part due to the difficulty in assessing

low-grade Ta and Nb.

SRMs 3, 4, 6, 7, 13, and 15 are considered suitable for use in assessment of Nb

grades greater than 900 ppm. SRMs 8, 9, 10, 12, and 14 are considered acceptable

for assessing Ta grades greater than 170 ppm.

All but two SRM Ta results returned for 2008 core analysis by Acme (ICP-MS) were

within ±10% of the best values. Only four out of 10 Nb results were within ±10% of the

best values. A negative bias was noted for SRM Ta values for 2009 core analysed by

Acme XRF (fusion) methods. No bias was observed for Nb.

11.6.2 Assessment of Accuracy with Secondary Laboratory Pulp Checks

Pulp checks were submitted to secondary laboratories by Commerce for all drill

campaigns. Secondary laboratory pulp checks are used to assess primary laboratory

accuracy by checking for biases between the laboratories.





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The 2005 pulp check samples were analyzed for Ta and Nb by Global Discovery

Laboratories in Vancouver using an XRF (fusion) method. An RMA chart of the

primary and check results shows a 44% negative bias and an 11% positive bias for

primary laboratory Ta and Nb respectively. There is no significant bias for samples

collected in 2008 after removal of outliers.

Pulp check sample analyses were not systematically performed between 2005 and

2008, and SRMs were not consistently included with the checks. Therefore the results

are considered an indicator of laboratory bias, but in isolation are not considered to be

definitive. The pulp check results indicate that there was improvement in agreement

between laboratories between 2005 and 2009.

11.6.3 Assessment of Precision with Duplicates

Between 2006 and 2009, 716 quarter-core field duplicates were submitted as control

samples to support primary assay results. Duplicates are used to assess for sampling

bias and the ability of the laboratory to reproduce their results. A duplicate result is

considered a failure if the absolute relative difference (ARD) between the pairs

exceeds a given threshold.

Field duplicate pairs sampled between 2006 and 2009 have a 7.2% and a 18.8%

failure rate for Ta and Nb respectively. All but the 2007 and 2008 Ta results are above

the acceptable 10% failure rate threshold. Ta precision failure rate increased in 2009,

which is coincident with a move to XRF (fusion) analysis. A better assessment of

precision would be to use pulp duplicate pairs.

In 2008, 268 pulps were resubmitted to Acme for ICP-MS analysis. Min-Max plots of

these data indicate excessive failure rates, suggesting analytical precision was not

under control for Ta or Nb.

ARD is commonly related to grade. As sample grades approach the lower detection

limits of the given analytical procedure, the ARD typically increases. A practical

detection limit (PDL) for an analytical procedure is the grade at which the ARD

exceeds 100%. The PDL is generally greater than the lower detection limits reported

by laboratories for a given analytical procedure. Based on the PDL for the 2008 pulp

duplicates the ARD generally exceeds the 10% threshold for both Ta and Nb at all

grades. However, there is no significant bias between the duplicate pairs for Ta or Nb.

In 2009 Commerce initiated an extensive re-assay program which culminated in a

decision to move to XRF (fusion) analysis for Ta and Nb. Analysis of pulp duplicates

from these tests showed that despite Ta and Nb detection limits being reported down





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to 5 and 10 ppm, the PDL was much higher. Forty-eight pulp duplicates were inserted

with the primary assay submission of 2009 core. There is an excessive failure rate for

Ta pairs but an acceptable failure rate for Nb. The PDL may be near 50 ppm for both

elements.

11.6.4 Assessment of Contamination Using Blanks

A total of 367 coarse blanks were submitted over the duration of the assay programs.

In general the Ta and Nb values returned for the blanks are at or near detection limits

suggesting little or no carry-over contamination. Sporadic carry-over contamination is

indicated and could be a result of sample swaps or poorly prepared blanks. The

abrupt jump in the Ta values in 2009 reflects a change in the lower detection limit as a

result of Commerce switching from ICP-MS to XRF(F).

11.7 Databases

Collar, down-hole survey, geology, specific gravity and assays were stored in a

Gems™ database format. Prior to GEMS™, data capture occurs in a variety of

formats from hand logs and Excel files.

Assay data is received from the laboratories via comma-separated value (CSV) data

files. These files are compiled and imported into the Gems database. After data are

imported, visual checks are done to ensure that data placement was correct within the

various database fields.

Commerce has migrated to a Century database during 2011 for all data captured

during 2011 exploration activity. Historic data is being migrated to the Century

database on an incremental basis with completed anticipated by 2012.

11.8 Security

Core is delivered by the drillers in the back of a pick-up truck to the Commerce field

office in Blue River. The boxes are laid out in order on saw horses and inspected by

the project manager. Core is logged by Dahrouge geologists. The core logging

supervisor completes spot checks for quality on a daily basis. The core storage,

logging and sampling facilities are not secured. Samples are placed in pails and

stored in the locked Quonset hut for security prior to shipping.

Core samples are commercially transported via Monashee Painting and Services of

Blue River, BC, to the preparatory laboratory. Sample sheet manifests are submitted

with the core samples. The manifests include information on the operator, sample





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preparation laboratory, and a sample list. Sample rejects returned from the laboratory

are stored in the onsite Quonset hut.


Sample preparation for samples that support Mineral Resource estimation has

followed a similar procedure for all of Commerce’s drill programs. The preparation

procedure is in line with industry-standard methods for sampling within carbonatite

deposits.

Exploration and infill core samples were analysed by independent laboratories using

industry-standard methods for tantalum and niobium analysis. Delay in submitting

2010 samples for analyses was related to addressing precision and accuracy issues.

To ensure appropriate quality and consistency in pulp grind size, all 2010 drill program

samples are being prepared at an independent third-party laboratory prior to shipping

samples to primary lab where sieve checks are completed prior to preparation for

analysis. During preparation a sanding step is completed to ensure no sample

carryover; this is extremely important during sample preparation of hygroscopic

carbonatites for analysis.

Typically, drill programs included insertion of blank, duplicate and standard reference

material samples. QA/QC submission rates meet industry-accepted standards of

insertion rates.

During preparation of the Mineral Resource estimate reported in Section 14, AMEC

and Commerce examined the sample preparation and analysis of Blue River core.

The principal findings of this work were as follows:

• Significant inter-laboratory grade biases are evident for Ta and Nb in the initial

sampling

• Acceptable inter-laboratory biases were achieved through the remaining sampling

programs

• SRM control samples indicate acceptable levels of accuracy are occasionally

achieved for Ta and Nb. Where acceptable levels are not achieved the SRM

results suggest a low bias for the primary results may exist

• Poor precision is evident for Ta and Nb results collected in 2008 and 2009.

However no consistent bias is evident.





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AMEC concludes that the Blue River sample results show imprecision but no

consistent bias and that those analyses are suitable for use in Mineral Resource

estimation.

Data that were collected were subject to validation, using in-built program triggers that

automatically checked data on upload to the database. Verification is performed on all

digitally-collected data on upload to the main database, and includes checks on

surveys, collar co-ordinates, lithology data, and assay data. The checks are

appropriate, and consistent with industry standards. Independent data audits have

been conducted, and indicate that the sample collection and database entry

procedures are acceptable.

Sample security has relied upon the fact that the samples were always attended or

locked in appropriate storage facilities. Chain-of-custody procedures consist of filling

out sample submittal forms that are sent to the laboratory with sample shipments to

make certain that all samples are received by the laboratory;

Current sample storage procedures and storage areas are consistent with industry

standards.

The QPs are of the opinion that the quality of the tantalum and niobium analytical data

are sufficiently reliable to support Mineral Resource estimation and that sample

preparation, analysis, and security are generally performed in accordance with

exploration best practices and industry standards. Caution should be applied in

assigning a high level of confidence to the results until precision and accuracy issues

are resolved.





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12.0 DATA VERIFICATION Commerce implemented an industry-acceptable quality control program to manage

logging, sampling, and analysis. Check samples for initial sample batches identified

discrepancies. In 2008 Commerce prepared matrix-matched standard reference

material control samples to monitor accuracy and initiated insertion of intra-laboratory

pulp duplicates in addition to secondary pulp check control samples. Primary

laboratory precision and accuracy has been poor but no significant biases are

apparent for the bulk of results.

AMEC completed a database verification check and concluded the collar coordinates,

down-hole surveys, lithologies, and assay databases were sufficiently free of error and

that the data are suitable to support mineral resource estimation.

AMEC collected and submitted 15 quarter-core samples to Acme in Vancouver for

preparation and analysis by Package 4B ICP-MS methods. A comparison of AMEC

results with matched interval results reported in the resource database supported the

grades reported in the resource database.

12.1 Comments on Section 12

In the opinion of the QPs, and based on the database verification performed by AMEC,

the collar coordinates, down-hole surveys, lithologies, and assays are considered

sufficiently free of error and that the data are suitable to support Mineral Resource

estimation.





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13.0 MINERAL PROCESSING AND METALLURGICAL

TESTING Testwork began in 2009 and continued into 2010 to develop a process flowsheet for

the Blue River Project. The testwork was based on material produced from two bulk

samples, BS-2F and BS–2G. Mineralogical analysis was performed to obtain

knowledge regarding the occurrence of the tantalum and niobium within the material.

Given the complexities with assaying for the tantalum, a fair amount of effort also went

into developing the appropriate routine for the assaying of samples.

The testwork primarily took place in two phases:

• Phase I – concentrated on the recovery of the tantalum–niobium minerals by

gravity although grinding and mineralogy investigations were also performed.

• Phase II – concentrated on the recovery and upgrading of the tantalum–niobium

minerals by flotation.

A large amount of work was performed in Phase I that showed gravity could

concentrate the material to a low-grade product, but that upgrading increasingly gave

lower levels of benefit as grade was sought. Mineralogical work completed before and

during this phase of work showed that the tantalum was not present as tantalite but

rather as the minerals ferrocolumbite and pyrochlore, which limits recovery by the

gravity route due to the low differential specific gravity between pyrochlore and gangue

minerals.

Work in Phase II saw the use of flotation concentration technology similar to that being

used for niobium-bearing carbonatites at Iamgold’s Niobec Mine in Quebec, Canada.

There was immediate success in the first phases of the work. Although there are

several stages to the concentration, the overall level of equipment, risk, and complexity

to produce a saleable or treatable concentrate is lower than the gravity route.

Process development work is continuing in this area, but for the purposes of the PEA,

the process suggested by Test F81 was selected as the basis of initial concentration

design because recoveries were good (approx. 70% for Ta) for this type of

mineralization and because a combined grade of 10% Ta–Nb was achieved. It is

expected that with further work, a combined grade of 30% Ta–Nb should be

achievable.

In both work phases, the emphasis of concentration techniques was to create a

material which would be easily upgraded by hydrometallurgical methods,

pyrometallurgical methods, or a combination of both. These processes would permit





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the separation of Ta from Nb, allowing payment for both products. To this end, an in-

depth review was completed of those technologies for the production of high-value

intermediate products and final products. There is confidence that the concentrate

could be reduced to metal by the aluminothermic process. Subsequently there would

be chlorination of the granulated metal alloy product and distillation of the anhydrous

metal chloride products to produce high purity Nb and Ta chlorides. Tantalum chloride

is the precursor to capacitor grade Ta powder and can be marketed as such.

However, both Ta and Nb chlorides can also be hydrolyzed and calcined to generate

high purity oxide products for other applications.

13.1 Head Samples for Initial Testing

In 2009, two bulk samples, BS-2F and BS–2G, sourced from a small pit in the Upper

Fir zone, and weighing approximately 200 t in total, were contract-crushed to a particle

size of <1 inch diameter. After crushing, each group of samples was homogenized

separately by a standard coning and quartering procedure. The well-mixed samples

were bagged into one tonne bags and put into storage. One tonne of each sample

was delivered to Met-Solve Laboratory (Met-Solve) in Burnaby B.C to air dry and

further reduce the size to -10 mesh for bench testing.

Met-Solve is a commercial mineral and metallurgical testing facility that is independent

of Commerce, and specializes in mineral beneficiation and hydrometallurgical

testwork.

The mineralogical examinations of all the bulk samples taken during the 2009

exploration program are described by Chudy (2009). Additional mineralogical

examinations were performed on some of the test products during the mineral

processing investigations.

The head assays, established using X-ray fluorescence (XRF) analysis, for the two

bulk samples are tabulated in Table 13-1.

13.2 Phase I Testing

13.2.1 Grinding Size

Each sample was subjected to gravity separation tests at five different grind sizes of

80% passing 500 µm, 230 µm, 100 µm, 74 µm and 45 µm to determine the liberation

size using a centrifugal concentrator. A standard seven-pass procedure was used to

simulate continuous gravity concentrator action.





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Table 13-1: Head Assay Grades, Bulk Samples BS-2F and BS–2G Sample Ta (ppm) Nb (ppm)

BS-2F 194 1,300

BS -2G 114 764

This work indicates that the liberation size for both samples is coarser than P80 of

76 µm. The relative position of the curves (Figures 13-1 and 13-2) indicates that

effective liberation for gravity is likely achievable at a grind size slightly coarser than

120 µm. The results for niobium are similar.

Assaying of the individual size fractions of the tailings from the BS-2G tests indicate

that there are still a few locked particles between 74 µm and 106 µm when ground to

P80 112 µm but that material coarser than 150 µm does not contain any tantalum.

Given the natural size distribution obtained in grinding, this implies that effective

liberation for processing, is about P80 of 125 µm for gravity treatment and slightly

coarser for flotation (P80 up to 160 µm). These numbers are in line with independent

findings from the mineralogical examination of all bulk samples of the 2009 exploration

program.

13.2.2 Roughing and Cleaning Gravity Concentration

With the establishment of the grind size and initial gravity results, it was decided to

progress with the gravity concentration work. The two samples were treated with a

centrifugal concentrator, using 10 consecutive stages for rougher concentration

followed by three cleaning stages of the combined rougher concentrates. Four

different grind sizes were tested for each sample. All results were similar, with

recoveries falling off quickly in cleaning and inability to raise the grades any higher

than Ta 3,500 ppm (0.35% Ta). Results from sample BS-2G are shown on Figure 13-

3.

Large batch samples of 60 kg were tested using a Falcon Centrifugal Gravity

Concentrator in 10 consecutive stages to produce a rougher and a scavenger

concentrate at a grind size of P80 100 µm. The rougher concentrate only was

screened to produce three size fractions as follows:

• +74 µm

• 37 to 74 µm





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• -37 µm.

Figure 13-1: Sample BS-2F – Gravity Separation (Different Grinds)

Figure 13-2: Sample BS-2G – Gravity Separation (Different Grinds)





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Figure 13-3: Rougher and Cleaners by Centrifugal Gravity Concentration

Each fraction was then cleaned by gravity using a Wilfley shaking table, with a

medium-size deck. Results were similar to the gravity separation using centrifugal

separator only, with no improvement in recoveries or grades. These fractions were

also tested using a Mozley table concentrator to determine the upgrading

characteristics of the products. Results showed that while it would be possible to

increase the grades by up to six times at the laboratory level, the recoveries would

drop accordingly. The results are shown in Figure 13-4.

Tests were also performed to determine the benefits of additional steps, such as de-

sliming and de-sulphidization. These procedures were incorporated into the testwork

but essentially AMEC was of the opinion that concentration by gravity as the primary

method was not the optimum choice for Project development.

13.3 Phase II Testing

13.3.1 Flotation Tests

A series of tests were performed by Met-Solve using varying de-sliming procedures

followed by flotation with various combinations of reagents. Results were not

successful, with recoveries generally below 15%.

A few tests were then performed using the flotation procedures used at the Niobec

Mine, operated by Iamgold Corporation in Quebec.





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Figure 13-4: Upgrading by Wilfley and Mozley Units

The first tests were immediately successful with higher recoveries and grades in the

roughers than the gravity method. While the rougher stage gave good results, the

cleaning stage was problematic due primarily to non-optimized conditions at this

preliminary stage of testing. Cleaning tests were performed and it was shown that a

total oxide grade of more than 30% combined Nb2O5 and Ta2O5 is achievable although

not at high recoveries. Until such time as the procedures have been well established

and stabilized, and are reproducible on larger test weights, the recoveries indicated

should be considered semi-quantitative and indicative of the possibilities.

Once improved de-sliming equipment was purchased, a series of de-sliming tests were

performed to determine the best range of operating parameters which indicated

optimum ranges are similar to those obtained at Niobec. Tests were then performed to

optimize the kinetics of the rougher tantalum–niobium flotation. It has been shown that

control of the pH through the stages is critical. Also important is the formulation of the

main collector, a tallow diamine acetate. Previously available as a commercial

product, Duomac-T, it is no longer available and current practitioners such as the

Niobec Mine now purchase the two main reagents (the amine and acetic acid) and

prepare the collector at site.

Process development work is continuing in this area, but for this report, the process

suggested by Test F81 (see Table 13-2) has been chosen as the basis of initial

concentration design as recoveries were good (approx. 70% Ta) for this type of

mineralization and due to a combined grade of 10% Ta–Nb (equivalent to

approximately 14% combined oxides) being achieved. It is expected that with further

work, a combined grade of 30% combined oxides should be achievable.





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Table 13-2: Results from F81

Mass Assay Recovery

Products Ta Nb S Ta Nb S

% ppm ppm % % % %

Cyclone Overflow #1 16.5 49 338 0.35 6.9 7.4 9.0

Cyclone Overflow #2 7.1 73 410 0.41 4.4 3.8 4.5

Carbonate Concentrate 28.0 25 151 0.17 5.9 5.6 7.3

Pyrrhotite Concentrate 1.7 152 275 27.38 2.2 0.6 73.5

Magnetic product 0.1 56 395 21.59 0.0 0.0 2.3

Stage 5 Pyrochlore Cleaner Con 0.6 12839 86732 1.14 69.8 72.7 1.1

Stage 5 Pyrochlore Cleaner Tail 0.4 228 1816 0.15 0.7 0.9 0.1









Total Pyrochlore Rougher

Concentrate 25.1 367 2492 0.07 78.7 82.3 2.8

Flotation Tails 21.5 10 10 0.02 1.8 0.3 0.7

Calculated Feed 117 760 0.64 100 100 100

Assayed Feed 113 764

Results of preliminary dilute hydrochloric acid leaching tests indicated that low- to

intermediate-grade gravity and flotation products can be upgraded significantly with

negligible loss of Ta+Nb. The final upgrading flowsheet will be based on an economic

comparison between pay metal losses from physical beneficiation and the cost of acid

plus stabilization/disposal of the leach products.

Table 13-3 presents results of a four-stage hydrochloric acid (pH 2, pH 1.2,

6N/1h/100ºC, 6N/5h/100ºC) on a flotation middling product.





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Table 13-3: Results of a Sequential Hydrochloric Acid Leach of Flotation “Middling” Products Weight Assay (ppm) Distribution (%)

(g) Ta Nb Ta Nb

Stage 1 Filtrate 180.0 0.002 0.00 0.0001 0.000004

Stage 2 Filtrate 220.0 0.031 0.34 0.001 0.002

Stage 3 Filtrate 255.0 0.024 0.63 0.001 0.003

Stage 4 Filtrate 210.0 1.491 12.27 0.056 0.053

Filter Cake 10.7 51,813 453,168 99.9 99.9

Calculated Head 26,824 234,609 100.0 100.0

Assayed Head 20.0 27,663 245,813

Most of the weight loss (>80%) was achieved in the first two stages, clearly indicating

the technical feasibility of upgrading by acid leaching with negligible solution loss of

Ta + Nb. The final leach residue assay is >50% Ta + Nb. The stage 3 and 4 strong

acid leaches were designed to investigate the possibility of dissolving Ta + Nb, but the

minerals appear to be entirely resistant to this relatively aggressive leach.

13.4 Review of Concentrate Treatment Options

In both phases, the emphasis of concentration techniques is to create a material which

would be easily upgraded by hydrometallurgical methods, pyrometallurgical methods,

or a combination of both. This has led to an in-depth review of those technologies for

the production of high value intermediate products and final products. There is

confidence that the concentrate could be reduced to metal by the aluminothermic

process. Subsequently there would be chlorination of the granulated metal alloy

product and distillation of the anhydrous metal chloride products to produce high purity

Nb and Ta chlorides. Tantalum chloride is the precursor to capacitor-grade Ta

powder, so would be marketed in this form. Niobium chloride can be sold as a

chemical precursor. Both Ta and Nb chloride products can be readily converted and

marketed as high purity Ta2O5 and Nb2O5 oxides respectively.

13.5 Accuracy of Assaying

A review of all calculated and measured feed assay results for tests using bulk sample

BS-2G was performed to check on the accuracy of the chemical analysis and the tests

results. It was decided to continue the assaying of low values, such as tailings, in

duplicate on separate aliquots; this procedure will continue as these assays could

introduce variations to results.





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In the opinion of the QPs, the following conclusions are applicable:

• Tantalum and niobium occur as ferrocolumbite and pyrochlore, which are

amenable to conventional flotation and proven refining processes with estimated

recoveries of 65% to 70%. For the purposes of the financial analysis in Section 22

of this Report, it was assumed that the process plant will have a 65% recovery for

Ta and 69% recovery for Nb in the flotation stage. The filtration process will have

a 97% recovery for both Ta and Nb

• The metallurgical testwork has shown that it is possible to concentrate the tantalum

and niobium minerals into a concentrate suitable for extraction of the metals into

saleable products. The first step of the process uses typical grinding followed by

flotation. The secondary treatment or metal extraction of the material is possible

by an existing method such as aluminothermic reduction followed by chlorine

refining. These results are suitable to support estimation of Mineral Resources for

the deposits.





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14.0 Mineral Resource Estimates The PEA conducted was based on a mineral resource estimate announced in

February 2011 (“Blue River Ta-Nb Project NI 43-101 Technical Report, Blue River,

British Columbia” by AMEC with an effective date 31 January 2011). AMEC used the

drill results up to the end of 2009, which includes 183 drill holes comprising 37,446

metres of HQ drill core, 8,218 sawn core samples, and 8,434 lithology records to

develop the mineral resource estimate.


model and were found to be reasonably consistent with the geology predicted by the

model. As well, the 2011 preliminary drill results were inspected at site on vertical

cross-sections and were also found to be reasonably consistent with the predicted

geology model.

The new effective date of the mineral resource estimates, based on the review of

recent drill results, is now 29 September 2011. There has been no change to the

closure date of the database used in the estimate, which is confined to all drilling that

was completed up to the end of December 2009.

14.1 Geological Models

The resource block model was constructed inside carbonatite only using 183 diamond

drill holes. All surrounding lithologies including fenite carry fairly low Ta2O5 and Nb2O5

grade and are considered sub-economic. Generally assay data exists only for

carbonatite. There are some assay values for fenite and adjacent wall rocks but not in

sufficient numbers to create a block model for these lithologies.

Geological interpretations were provided by Commerce to AMEC as 2D vertical

sections and 3D solids in DXF format. The solids were created by Dahrouge

geologists for the major lithologies with the exception of gneiss which was left as a

default. The carbonatite solids were provided as 39 structural (different strike, dip

and/or pitch) domains. AMEC reviewed the geological interpretations and 3D solids.

Minor modifications were made to the solids to provide a better agreement with the

drilling and geological interpretation.

The block model consists of regular blocks and no rotation was used. The block

model framework parameters are listed in Table 14-1.

Blocks in the block model were coded by lithology solids, and the volume of each

lithology solid was compared with the volume of the blocks inside a particular solid. A





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block was tagged by a particular solid code if at least 50% of the block volume

belonged to this solid. The block model and corresponding lithology solid volumes

compare within ±1%.

Table 14-1: Block Model Dimensions Axis Origin* Block Size (m) No. of Blocks Model Extension (m)

X 352,350 5 250 1,250

Y 5,795,850 5 390 1,950

Z 925 2.5 244 610

Note: *Origin is defined as the bottom southwest corner of the model, located at the lowest combined northing and

easting coordinates and the lowest elevation

14.2 Exploratory Data Analysis

Exploratory data analysis (EDA) was performed on the composites to better

understand the data used in the resource estimation. Arithmetic and log histograms

and probability plots of Ta2O5 and Nb2O5 composites in carbonatite were constructed.

The distributions are skewed, and Nb2O5 distribution is approximately lognormal. The

coefficients of variation are low and support the use of linear grade interpolation

methods such as kriging or inverse distance methods.

AMEC calculated contact profiles on assay data in MineSight® software to analyze

grade behaviour at lithology boundaries. There were sharp differences in grade for

each of the variables on the carbonatite and fenite boundary, meaning that values from

outside the carbonatite should be disregarded in the interpolation process of Ta2O5

and Nb2O5 grade inside the carbonatite.

14.3 Density Assignment

Resource block model specific gravity was not estimated; instead it was assigned to all

blocks in the block model (including blocks outside of carbonatite) as follows:

• 3.01 value was assigned to all blocks in carbonatite

• 2.97 value was assigned to all blocks in fenite

• 2.82 value was assigned to all blocks in gneiss

• 3.05 value was assigned to all blocks in amphibolites





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• 2.62 value was assigned to all blocks in pegmatite.

14.4 Grade Capping/Outlier Restrictions

AMEC conducted grade capping on original samples that are mostly 1 m long.

Capping was required to limit the influence of outliers. The choice of capping was

based on visual inspection of histograms and probability plots. The amount of capping

was small; top-cuts of 1,000 ppm Ta2O5 and 10,000 ppm Nb2O5 were used in

carbonatite. Only four Ta2O5 samples and eight Nb2O5 samples were capped resulting

in an expected metal removal of 0.12% Ta2O5 and 0.26% Nb2O5.

14.5 Composites

Capped drill core assays were composited down the hole to a fixed length of 2.5 m.

Compositing of Ta2O5 and Nb2O5 was performed in MineSight® software honouring

geologic boundaries. Composites with length less than 1.25 m were merged with the

previous composite.

14.6 Variography

AMEC used both in-house software and commercially-available Sage2001 software to

produce variogram maps and to construct down-the-hole and directional correlograms

for carbonatite composites. Ta2O5 and Nb2O5 correlograms were created within the

entire carbonatite zone. Two spherical models were used to fit the experimental

correlograms.

14.7 Estimation/Interpolation Methods

Ta2O5 and Nb2O5 grade was estimated in the carbonatite using both Ordinary Kriging

(OK) and inverse distance to the third power (ID3) interpolation methods. A four-pass

interpolation approach was used with each successive pass having greater search

distances. A hard boundary was used, meaning that composites from outside the

carbonatite were ignored in the interpolation process. Estimation was done separately

within each limb of the carbonatite folds. A total of 39 different limbs / structural

domains were identified and used in the estimation process. The limbs differ in the

orientation of the search ellipse.

The rotation angles of the search ellipse are the same for each pass, but they are

different for each of the 39 structural domains. They reflect average strike, dip, and

pitch of each fold limb/domain.





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14.8 Block Model Validation

AMEC completed a visual inspection of composites and blocks in vertical sections and

plan views. The model generally honours both Ta2O5 and Nb2O5 data well, and grade

extrapolation is well-controlled where sufficient data exists.

The OK and ID3 block models were checked for global bias by comparing the average

grade (with no cut-off) from these models with that obtained from nearest-neighbour

(NN) model estimates. Global biases are well below the recommended AMEC

guidelines of ±5% (relative difference).

AMEC also estimated the impact of outlier capping on the estimated global mean of

the model. A comparison of global means of capped and uncapped OK and ID3

models showed the amount of metal removed by capping is minor (0.1% for Ta2O5 and

0.3% Nb2O5).

Checks for local biases were performed for Ta2O5 and Nb2O5 by creating and

analyzing local trends in the grade estimates using swath plots. Swath plot checks,

conducted by AMEC, show that there are no local biases between ID3 / OK and NN

models for estimated Ta2O5 and Nb2O5 in carbonatite.

Selectivity analysis for Ta2O5 and Nb2O5 was completed using the Discrete Gaussian

Model for change of support from composite size units to an SMU size. This was done

using AMEC in-house software (Herco). The aim of this analysis is to assess if the

estimated resource reasonably represents the recoverable resources (represented by

Herco curves) relative to the proposed mining method. The selectivity analysis

assumed a 10 m by 10 m by 5 m block as the smallest selective mining unit (SMU) for

Blue River. The Herco selectivity analyses show that the Ta2O5 and Nb2O5 ID3

models are properly smoothed. The ID3 model produces slightly higher average

grades of Ta2O5 and Nb2O5 compared to the OK models above cut-off grades of 140

ppm Ta and 900 ppm Nb. The ID3 model was chosen for tabulating the Blue River

Mineral Resources as the OK model is considered too smooth.

14.9 Classification of Mineral Resources

The Mineral Resources were defined taking into account the CIM Definition Standards

(2010).

Based on a grade drill-hole spacing study AMEC established the following criteria for

classification of Mineral Resources at Blue River:





Project No.: 162230


Inferred Mineral Resources:

• Minimum one hole

• Distance to the closest composite less than 110 m

Indicated Mineral Resources:

• Minimum two holes


• Distance to the second closest composite less than 70 m

Measured Mineral Resources:

• Minimum three holes


• Distance to the second closest composite less than 40 m

The Mineral Resource classification at Blue River was restricted to Indicated or

Inferred, based on the following:

• Confidence limits resulting from drill hole spacing studies

• Concerns of analytical precision and accuracy for the sample dataset

• Local discrepancies in the model identified with the 2010 drilling

• Lack of metallurgical testwork on the final stage of the proposed metallurgical

process.

Eighty percent of the carbonatite blocks are classified as Indicated. Fourteen percent

of the carbonatite blocks are classified as Inferred. Six percent of the block model in

carbonatite is unclassified. Blocks that fall into the unclassified category are within

carbonatite solids that were intersected typically by one isolated drill hole. The

geological continuity and volume of those solids cannot be reasonably assumed.

14.10 Reasonable Prospects of Economic Extraction

14.10.1 Tantalum Pricing

Tantalum is quoted as a number of different forms, but the two most common are:





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• Ta2O5 in tantalite concentrate: a non-refined, tantalum-bearing concentrate of

variable composition and trace element content

• Tantalum metal scrap (99.9% pure Ta): this form of tantalum product receives a

premium price in the market relative to tantalite concentrate.

Over the last six years tantalite concentrate prices ranged from US$75/kg contained in

Ta2O5 to US$100/kg contained in Ta2O5 (US$35/lb to US$45/lb). In the same period

tantalum metal scrap prices ranged from US$110/kg Ta to US$180/kg Ta metal

(US$50/lb to US$80/lb). In 2010, prices rose dramatically in response to changing

market conditions including reduced production, increased concerns about conflict-

tantalum production in Africa and depletion of known strategic stockpiles. In mid-

October 2010 the price for Ta2O5 in tantalite concentrate was US$195/kg and for

tantalum metal scrap was US$280/kg.

The higher price for tantalum metal scrap compared to the price for Ta2O5 in

concentrate is considered a proxy for the added value Commerce should recognize by

refining the Blue River concentrate to high-purity Ta2O5.

In AMEC’s opinion, the base case price for tantalum metal scrap is reasonable for

constraining Mineral Resources based on recent market conditions, but notes it is

significantly higher than historical prices. There is a risk that using current price

assumptions may not reflect the long-term price of Ta and Nb, particularly in the

present volatile market conditions.

14.10.2 Niobium Pricing

Niobium generally trades as Nb oxide, metal, or ferroalloy, and the price has remained

relatively constant at US$44.08/kg Nb metal in oxide (US$20/lb Nb) over the last

several years. A base case price of US$46/kg Nb metal in oxide was assumed.

14.10.3 Physical Assumptions

• Tantalum–niobium mineralization is hosted in carbonatite

• Continuous mineralization is found in moderately flat and wide carbonatite bodies

with modest dips

• Mineralized areas 20 m to 70 m in height are expected in several zones

• Steep topography provides access to the mineralized areas in the form of adits on

the hillsides

• Fair to good rock conditions are expected in the majority of the deposit





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• Identified faulted zones may require wider pillars to avoid unstable mining conditions.

14.10.4 Operational Considerations

• The underground mining method used to constrain the mineral resources is room

and pillar

• 70% of the resources could be recovered by mining with 30% of the resource

expected to remain as unrecoverable pillars for stability considerations

• A minimum stope size of 10 m x 10 m rooms with heights from 5 m to 15 m is

assumed

• An external dilution factor was not considered during this estimation

• The concentration method considered is flotation followed by a refining process on site;

global recoveries to obtain metal grade products were assumed to be 65% for tantalum

and 68% for niobium.

14.10.5 Economic Assumptions

Most of the operating cost assumptions were extrapolated from previous studies

prepared by AMEC:

• Mining and backfilling cost US$32.00/t

• Processing and refining cost US$17.00/t

• General and Administration (G&A) US$2.70/t

• Base case scenario price of tantalum US$317/kg Ta in oxide product

• Price of niobium US$46/kg Nb in oxide product

14.10.6 Economic Cut-off

Block Unit Value

The block model was adapted to represent the two payable metal contents in terms of

block unit value (BUV) in US$/t using the following formula:





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BUV = (Ta2O5 grade in ppm * Ta recovery factor * Ta price in US$/g *

proportion of 2Ta:Ta2O5) + (Nb2O5 grade in ppm * Nb recovery factor * Nb

price in US$/g * proportion of 2Nb:Nb2O5)1

For the base case scenario:

BUV = (Ta2O5 * 0.654 * 0.317 * 0.819) + (Nb2O5 * 0.682 * 0.046 * 0.699)

Considering the direct operating costs, a cut-off value of US$52/t was assumed for the

material susceptible to be mined by bulk methods and a cut-off value of US$59/t for

material to be mined by the selective methods.

The tool “Stope Analyzer” from Vulcan® was utilized to identify the blocks that exceed

the cut-off value while complying with the aggregation constraint of the specified

minimum stope size. This tool “floats” a stope with the specified dimensions and flags

each block when the average block unit value of the contained blocks within a stope

exceeds the designated cut-off value.

For constraining Mineral Resources deemed to be mined by underground methods,

the use of this tool as an alternative to a conventional economic grade-shell provides

an advantage based on the ability to aggregate blocks into the minimum stope

dimensions and the automatic elimination of outliers that do not comply with this

condition.

14.11 Mineral Resource Statement

The Mineral Resources were classified in accordance with definitions in the Canadian

Institute of Mining, Metallurgy, and Petroleum (CIM) Definition Standards for Mineral

Resources and Mineral Reserves (2010), incorporated by reference into NI 43-101.

The Indicated Mineral Resources are 36.35 Mt containing 195 ppm Ta2O5 and

1,700 ppm Nb2O5. Inferred Mineral Resources are 6.40 million tonnes containing 199

ppm Ta2O5 and 1,890 ppm Nb2O5.

Table 14-2 shows the estimated Mineral Resources for the Project. AMEC cautions

that Mineral Resources are not Mineral Reserves as they do not have demonstrated

economic viability.

1 Note: Ta2O5 = 1.2211 * Ta; Nb2O5 = 1.4305 * Nb





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Table 14-3 shows the sensitivity of the Blue River Mineral Resources to tantalum metal

price. Sensitivities are based on a fluctuating metal price.

The PEA is based on a mineral resource estimate announced in February 2011 (“Blue

River Ta-Nb Project NI 43-101 Technical Report, Blue River, British Columbia” by

AMEC with an effective date 31 January 2011). AMEC used the drill results up to the

end of 2009.


model and were found to be reasonably consistent with the geology predicted by the

model. As well, the 2011 preliminary drill results were inspected at site on vertical

cross-sections and were also found to be consistent with the predicted geology model.

The effective date of the Mineral Resource estimate has been updated to 29

September 2011 reflecting reviews of the latest drill information.

The Mineral Resource estimate is supported by base case price assumptions for Ta

and Nb which reflect current market prices, but are higher than historic average prices.

A review of publicly-available market analysts’ opinions shows a general agreement

that current market conditions support the probability of sustained higher prices. There

is no assurance the higher prices will be realized over the life of the Project.





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Table 14-2: Blue River Project Estimated Mineral Resources; Effective Date 29 September 2011, Tomasz Postolski, P. Eng.

Ta Price Confidence Mass Ta2O5 Nb2O5 Ta2O5 Nb2O5

(US$/kg) Category (tonnes) (ppm) (ppm) (1000s of kg) (1000s of kg)

317 Indicated 36,350,000 195 1,700 7,090 61,650

Inferred 6,400,000 199 1,890 1,300 12,100

Notes:

1. Assumptions include US$317/kg Ta, US$46/kg Nb, 65.4% Ta2O5 recovery, 68.2% Nb2O5 recovery,

US$32/tonne mining cost, US$17/tonne process and refining cost. Mining losses = 0% and dilution = 0%.


Analyzer”.

3. An economic cut-off was based on the Ta and Nb values per block which is variable based on the location of

blocks used in the Mineral Resource estimate. A block unit value cut-off ranged from US$52 to US$59.


5. In situ contained oxide reported.

Table 14-3: Blue River Project Sensitivity of Estimated Mineral Resources to Tantalum Price; Effective Date 29 September, 2011, Tomasz Postolski, P.Eng, (Base Case is bolded)

Ta price

(US$/kg)

Confidence

Category

Mass

(tonnes)

Ta2O5

(ppm)

Nb2O5

(ppm)

Contained

Ta2O5

(1000s of kg)

Contained

Nb2O5

(1000s of kg)

470 Indicated 51,130,000 188 1,410 9,610 72,300

Inferred 8,100,000 192 1,700 1,600 13,800

381 Indicated 44,430,000 192 1,530 8,530 68,020

Inferred 7,300,000 196 1,780 1,400 13,000

317 Indicated 36,350,000 195 1,700 7,090 61,650

Inferred 6,400,000 199 1,890 1,300 12,100

272 Indicated 29,990,000 197 1,850 5,910 55,480

Inferred 5,500,000 201 2,010 1,100 11,100

238 Indicated 25,130,000 197 2,000 4,950 50,240

Inferred 4,900,000 202 2,110 1,000 10,400

Notes:

1. Ta price was varied and all other assumptions remain the same as base case.

2. Base case is in bold.


Analyser”.


5. In-situ contained oxide reported





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14.12 Factors That May Affect the Mineral Resource Estimate

Factors which may affect the Mineral Resource estimates include:

• Commodity price assumptions

• Metallurgical recovery assumptions

• Cut-off criteria used to report the Mineral Resource estimates

• Stope dimensions and mining and recovery assumptions used to constrain the

estimates.


The QPs are of the opinion that the Mineral Resources for the Project, which have

been estimated using core drill data, have been completed using industry best

practices, and conform to the requirements of CIM (2010).

Mining assumptions may have an impact on the contained metal uncertainty and as

such on the Mineral Resource classification. The mining cost and method assumptions

used to constrain the Mineral Resource estimate in 2010 differ from those used in the

2011 PEA. During the next Mineral Resource update the PEA parameters should be

considered when classification and reasonable prospects criteria are evaluated.





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15.0 Mineral Reserve Estimates No Mineral Reserves have been estimated for the Project.





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16.0 MINING METHODS The mining section is partly based on 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 based on these Mineral Resources will be

realized.

16.1 Optimization

16.1.1 Assumptions

Mining assumptions used to define the Mineral Resource estimate in Section 14 were

modified during development of the mine plan.

The block model consists of regular blocks with dimensions of 5 m x 5 m in the

horizontal plane and 2.5 m vertically; no rotation was adopted.

The block model was adapted to represent the unit value per tonne of material

considering two payable metal contents. This value was named the block unit value

(BUV) in US$/t and was estimated using the following formula:

BUV = (Ta2O5 grade in ppm * Ta recovery factor * Ta price in US$/g *

proportion of 2Ta: Ta2O5) + (Nb2O5 grade in ppm * Nb recovery factor *

Nb price in US$/g * proportion of 2Nb:Nb2O5)

Price assumptions were US$317/kg tantalum metal and US$46/kg niobium metal,

contained in oxide product.

Tantalum and niobium mineral occurrences are amenable to conventional flotation and

refining processes with estimated overall recoveries of 65.4% and 68.2%.

The block model unit was valued using the following formula:

BUV = (Ta2O5 * 0.654 * 0.317 * 0.819) + (Nb2O5 * 0.682 * 0.046 * 0.699)

For purposes of estimating the Mineral Resources, an underground mining cut-off

grade of US$52/t was established for the material susceptible to be mined by bulk

methods and a cut-off value of US$59/t for material to be mined by the selective

methods. These figures were based on conceptual preliminary direct operating costs

prior to developing the PEA’s mine plan and its associated costs.





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Mineral Resources considered in the preliminary mine plan are those tabulated in

Table 14-2.

The “Stope Analyzer” tool from commercially-available Vulcan® software was utilized to

identify blocks within the resource model for which the block unit value (BUV)

exceeded the cut-off value while complying with an aggregation constraint of a

specified minimum stope size. This tool “floats” a stope with previously specified

dimensions and flags each block when the average block unit value of the contained

blocks within a stope exceeds the designated cut-off value.

The use of this tool is an alternative to a conventional economic grade-shell for

constraining Mineral Resources deemed to be mined by underground methods, and

provides an advantage based on the ability to aggregate blocks into the minimum

stope dimensions and the automatic elimination of outliers that do not comply with this

condition. Stope dimensions are detailed in Table 16-1.

16.1.2 Mining Method

Blue River Project is located largely along the steep, west-facing slopes of the

Monashee Mountains and it is closely positioned to, and just east of, the North

Thompson River. As mineralization is close to surface, extraction of the mineralized

material could potentially be by either open pit or underground methods, or a

combination of both.

The open pit method offers a potentially higher mining recovery and the opportunity for

extracting the mineralized material with lower operating and capital costs. On the

other hand, it presents technical challenges for:

• Placement of large volumes of waste rock in limited amenable areas in the vicinity

of the project

• Exposure to heavy snow loads on the dumps and excavations and the associated

risk of avalanches

• Need for surface water diversion and groundwater control infrastructure as well as

the mitigation of these impacts

Initial assessment identified technical challenges and increased costs for tailings and

waste rock disposition related to local topography, stream courses and climate. For

this reason a decision was made to consider an underground mining scenario for PEA

purposes.





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Table 16-1: Minimum Stope Dimensions for Constraining the Subset of Mineral Resources within Designed Stopes

Mining Method Width (m) Length (m) Height (m)

Sub-level Open Stoping 10 10 15

16.1.3 Production Rate

In 2009, AMEC carried out an “order of magnitude” financial analysis to evaluate a

range of processing throughput rates. This analysis indicated that the Project required

processing rates of greater than 5,000 t/d to provide a reasonable economic return on

investment.

A processing rate of 7,500 t/d was assumed for Mineral Resource estimation and for

the conceptual design of an underground mine for the Project.

It is AMEC’s opinion that this mining rate is reasonable based on the geometry of the

deposit and the mining method selected.

16.1.4 Backfill Option

Topography, drainage and precipitation aspects limit the options for surface tailings

storage facility locations. AMEC visited the Project site in July 2010 when potential

areas for infrastructure and tailings storage were identified. Due to the limited options

available for tailings storage, Commerce and AMEC agreed that underground mining

should consider the ability to dispose of filtered tailings in the mined-out areas, but that

this would not be the preferred option for the Base Case.

16.2 Geotechnical Conditions

In the second half of 2010 AMEC carried out a geotechnical program to provide design

guidelines for the Blue River Project. This program included:

• A geotechnical site investigation program

• Training for site geologists and geological technicians in oriented core logging

• QA/QC site visits during the geotechnical drilling program

• Engineering analysis and recommendations applicable to the underground mining

method.





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Rock types have been grouped into two main geotechnical domains: Intrusive and

Layered Rocks. The Intrusive group encompasses carbonatite and fenite rocks while

the Layered Rocks group encompasses gneiss and amphibolite.

Based on rock strength analysis and assessment of rock quality designation (RQD),

joint spacing, joint condition and groundwater condition assumptions, the rock mass

characteristics detailed in Table 16-2 were obtained for each rock group.

Table 16-2: Rock Mass Characteristics by Rock Group

A structural set analysis showed similarities between the Intrusive and Layered Rock

groups with a strong concentration of flat joints dipping between 0º to 30º to the

east/northeast and sub-vertical joints striking northwest/southeast. The major joint

sets were identified as shown in Table 16-3.

16.3 Conceptual Mining Method

The major items governing the selection of underground mining methods include the

geology and geometry of the deposit and geotechnical properties of the mineralized

material and country rock.

AMEC followed the technique proposed by Nicholas (1992) to rank the underground

mining methods suitable for the Blue River deposit. AMEC classified the deposit as a

thick, tabular, and flat deposit with relatively uniform low grades and moderate

geotechnical conditions in the deposit as well as in the host rock. The three methods

with higher ranking are:

• Sub-level stoping

• Cut-and-fill

• Sub-level caving.

The first two methods were considered for this study. The first was evaluated for

areas where the carbonatite is greater than 15 m in thickness and the second for

thinner areas towards the edges of the carbonatite folds.





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During economic analysis it was found that the cut-and-fill method was likely to prove

uneconomic and it was excluded from the mine plan.

The assumed mining method is sub level open stoping with no backfill and no pillar

recovery.

Table 16-3: Major Joint Sets

16.4 Stoping Design

The Upper Fir deposit geological model shows thicknesses between 20 m to 80 m and

strike lengths between 50 m to 200 m in the east–west direction. Transverse

dimensions in the north–south direction range between 100 m to 500 m. Vertical

stopes oriented east–west with maximum dimensions of 30 m high, 15 m wide and

60 m long were selected for preliminary analysis.

Based on these stope dimensions, the following stope faces and hydraulic radius are

defined (Table 16-4).

16.4.1 Stability Analysis and Ground Support

As part of geotechnical analysis, AMEC carried out stability analysis based on the

empirical Mathews/Potvin Stability Graph method.

Results of the analysis indicated:

• Face A – Longwall: This wall is expected to be composed by competent Intact to

Blocky rock mass. The main jointing is sub-horizontal crosscut by two sub-vertical

joint sets. The stope wall is vertical and should be inherently stable unless

disturbed by over-blasting. Sloughing up to 0.5 m thick is expected. Cable bolting

may be required in certain places.

• Face B – Endwall: Numerical modeling suggests that the stress is relatively high

compared with the rock strength and the joint set configuration favours sliding

failure. Cable bolts are required to provide stability. A square cable pattern of 2 m

by 2 m is recommended and will be installed from the sill drift fanning upwards to

cover the unsupported span between level drifts.





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• Face C – Back: This face is horizontal and as such is expected to be in a relaxed

stress state. Gravity driven rock fall is likely the dominant failure mode. Face

reinforcement is required with the development of a self supporting rock arch over

the back face. A 15 m wide face requires cable bolting in square pattern of 2 m by

2 m and cable bolts of 9 m long.





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Table 16-4: Stope Faces and Hydraulic Radius Face A Face B Face C

Longwall Endwall Back

Height (m) 30 30

Width (m) 15 15

Length (m) 60 60

Hydraulic radius (m) 10 5 6

16.4.2 Stope Geometry

AMEC made adjustments to the preliminary stope design to reduce the demand for

cable bolting; specifically the Back was reduced to a 5 m-wide entry. This requires the

stope drilling pattern to be changed from vertical to fanned holes.

16.4.3 Mining Sequence

The stope face will advance from the east end of each block to the west in a retreat

manner. Once the maximum unsupported stable length of the stope has been

reached, a pillar will be established, and mining will resume from a new stope.

16.4.4 Conceptual Mine Design

AMEC developed a conceptual design for one of the potential areas to be mined,

geological domain A110. This domain was selected because it was one of the thicker

areas within the deposit and represented a reasonable test area for mine design.

16.4.5 Mining Dilution and Recovery

Material deemed to be mined by bulk mining methods represents 84% of the

resources. Within the mineable shapes there was internal dilution of 2% waste rock. It

was assumed that during mining 2% of waste material would be added as external

dilution and 2% of the broken material would not be recovered from the stopes due to

operational conditions.

The geotechnical investigation indicates that an extraction ratio of 67.5% is

reasonable. Applying this to subset of the Mineral Resources considered in the mine

plan results in an overall mining extraction of 58% and provides 25.0 Mt of material as

run-of-mine (ROM) production to be processed. Applying internal and external mining

dilution, the overall subset Mineral Resource grades were diluted to 185 ppm of Ta2O5

and 1,591 ppm of Nb2O5 for mine planning purposes.





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16.5 Drilling and Blasting

AMEC assumed conventional drilling methods: electro-hydraulic jumbos for face drift

rounds and electro-hydraulic long-hole drills for stoping drilling.

Due to the high precipitation in the region and water continuity along fractures and rock

layers, wet conditions were assumed for development and stoping areas requiring the

use of emulsion type explosives.

AMEC assumed the utilization of specifically-designed tanks to store emulsion

explosives in trucks delivering and loading explosives at development faces and

stopes. A centralized blasting system where all blasts are initiated sequentially from a

single location at the end of each shift will be used to reduce potential safety and

ventilation risks.

16.6 Mine Development

The deposit will be accessed through two main portals, the Upper and Lower Portals,

located in where the deposit outcrops on the hillside. The Upper Portal will be located

at Elevation 1,150 m. It will be used as the main entry and will have most of the mine

services. The Lower Portal will be located at Elevation 1,030 m and will be used for

haulage trucks access. Access to the portals will be by a road upgraded from existing

exploration roads. Location plans for the portals are included in Section 18.

The mine will be accessed by adits driven in pairs from the portals. All entries, ramps,

drifts and crosscuts will be 5 m wide by 5 m high with semi-arched back to

accommodate the haulage trucks plus mine services, including pipelines for

compressed air, drill and drainage water, for ventilation duct and electric and

communication cables.

Ramps will be driven at grades to a maximum of 15% to provide access to the

production areas. Two ramps or adits will be driven to each area to provide single-way

traffic of haulage trucks and to facilitate ventilation.

Top access crosscuts are driven from the main ramps to each level on vertical

intervals between 20 m to 30 m. Stope access crosscuts are driven along the levels

from west to east. Bottom-access crosscuts are driven to function as mucking drifts.

Ground support will be by grouted bolts in all man-entry drifts with steel mats, wire

mesh and plates assumed to be installed in 20% of the areas.





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The length of development necessary was calculated by designing centrelines of

ramps and drifts to provide access to all potentially mineable areas of the mine.

Additional development for re-mucking cut-outs, sumps, substation rooms, storage and

any other excavation needed for infrastructure was included as an allowance of 20% of

the semi-permanent development entries. The total development was estimated at

92,500 m during the life-of-mine.

A conceptual preliminary overall mine plan is included as Figure 16-1. Figure 16-2

displays an aerial view of the mining area from the Upper Portal.

16.7 Mineralized Material and Waste Rock Haulage

The PEA envisages that tailings material will be dry-stacked and waste rock will be

stored in the same general area. For convenience, the combined storage area is

referred to as the “co-disposal facility”.

Radio remote-controlled load-haul-dumps (LHDs) will be used to extract the

mineralized material from beyond the safety of the stope brow. This material will then

be loaded directly into the haulage trucks that will be spotted at the end of each stope

crosscut.

Underground trucks will haul the mined material through the access drifts and ramps,

unloading into the primary crusher surface stockpile located close to the Lower Portal.

Crushed material will be transferred to the process plant by a belt conveyor.

Waste from development will initially be utilized for construction of a structural shell for

the tailings co-disposal site located between Elevations 1400 m and 1600 m in an area

east of the process plant. The conveyor will be used to transport this material in

batches from the mine to a stockpile by the plant site.

A road developed at +10% grade will connect the plant with the co-disposal site.

Surface trucks will haul waste from the plant to the co-disposal site.





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Figure 16-1: Conceptual Mine Layout Plan (plan view projection)

Note: image figure colours may appear darker than reference key colours due to over-plotting of design layers





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Figure 16-2: Aerial View of the Mining Area from Upper Portal

Note: image figure colours may appear darker than reference key colours due to over-plotting of design layers. Light brown blocks in background of

figure are the potentially-mineable blocks.





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16.8 Mine Services

Underground mine services such as ventilation and air heating, compressed air, water

for drilling and electric power supply will be provided to the mine via adits from the

portals. The main power substation and the air compressors will be installed in

facilities located adjacent to the Upper Portal.

Other mine services will include all the systems and supplies needed for the mining

operations, including explosives storage, communications, monitoring and control

systems, road maintenance and a mine equipment underground maintenance facility.

Portable self-contained refuge stations will be provided for the mine and will be located

at convenient locations. Refuge stations provide a common assembly area in the

event of a mine fire or other emergency and are portable so they can be easily

relocated to the next active area.

16.9 Mine Development and Production Forecasts

The forecast preproduction development is 12,000 m and the annual development

over the ten-year mine life is forecast to decrease from 15,000 m in the first full

production year to about 2,150 m in the final year.

Production was estimated at 2.7 Mt/a of mineralized material. Following the

preproduction development year full production is maintained for nine years followed

by decreased production in Year 10 as the subset of the Mineral Resources

considered in the mine plan are exhausted.

At this preliminary level of study the stope mining sequence was not defined and

therefore average grades were used for each year in the mine plan.

There is opportunity to increase the net present value (NPV) of the project by mining

higher-grade zones early in the mine life providing that the sequence and overall

recovery of the stopes is not negatively affected.

The development and production forecasts are shown in Table 16-5.





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Table 16-5: Mine Development and Production Forecasts

Production Schedule Units Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5

Nominal production rate tpd

7,500 7,500 7,500 7,500 7,500

Scheduled working days days/year 360 360 360 360 360 360

Production of mineralized material t/year

2,700,000 2,700,000 2,700,000 2,700,000 2,700,000

Ta2O5 Grade ppm

185 185 185 185 185

Nb2O5 Grade ppm

1,591 1,591 1,591 1,591 1,591

Development (total) m/year 12,000 15,000 12,834 10,980 9,394 8,038

Capital development m/year 12,000 5,000 1,216 1,040 890 761

Operational development m/year 0 10,000 11,618 9,940 8,505 7,276

Production Schedule Units Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Total

Nominal production rate tpd 7,500 7,500 7,500 7,500 7,500

Scheduled working days days/year 360 360 360 360 360

Production of mineralized material t/year 2,700,000 2,700,000 2,700,000 2,700,000 700,000 25,000,000

Ta2O5 Grade ppm 185 185 185 185 185 185

Nb2O5 Grade ppm 1,591 1,591 1,591 1,591 1,591 1,591

Development (total) m/year 6,877 5,884 5,034 4,307 2,153 92,500

Capital development m/year 651 557 477 408 0 23,000

Operational development m/year 6,225 5,326 4,557 3,899 2,153 69,500





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16.10 Mine Equipment Requirements

Table 16-6 shows the major mining equipment and the support equipment required

each year for the life of mine and includes allowances for equipment utilization and

availability due to maintenance downtime. The table includes equipment required for

the tailings and waste co-disposal facility.

The “Equipment additions” reflect additional requirements based on increased activity

levels and replacements assuming typical useful operating lives for the units.

16.11 Mine Infrastructure

The mine infrastructure planned for the Project is listed in the capital costs section of

this report and includes establishment of mine portals, access roads to portals

underground maintenance bays, ventilation and heating systems, air compressors, fuel

tanks, explosives magazines, pumps, electrical transformers and ancillary equipment.

16.12 Mining Personnel

The mine is scheduled for three 8-hour shifts per day, 360 days per year. This will

require an equivalent of 4.5 mine crews working a rotating schedule for activities

scheduled seven days a week and three crews for activities scheduled five days a

week. The annual personnel requirements are shown in Table 16-7. This includes

management and supervisory personnel and personnel to operate and maintain

equipment used to service the surface roads and the tailings disposal site.


In the opinion of the QPs, the following conclusions are appropriate:

• The deposits are amenable to underground mining, and the PEA has been

developed assuming a Base Case sub-level open stoping mining method with no

backfill

• Material deemed to be mined by bulk mining methods represents 84% of the

Mineral Resources; within the stopeable shape, an additional 2% of waste was

identified as internal dilution.





Project No.: 162230


Table 16-6: Mining and Tailings Facility Equipment Requirements

Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Major Equipment Units Required

Jumbo - 2 boom 4 4 4 4 3 3 3 3 2 2 2

Longhole dri ll - 5 5 5 5 5 5 6 6 6 4

Bolter 3 3 3 3 3 2 2 2 2 1 2

Emulsion Truck 3 4 4 3 3 3 3 3 2 2 3

Scissor Lift 4 4 4 4 3 3 3 3 2 2 2

LHD - Ejector - 7 m3 3 5 5 5 5 5 5 5 4 4 4

Trucks 3 9 9 9 8 8 8 8 8 8 7

Major Equipment Purchases

Jumbo - 2 boom 4 - - - - 2 1 - - - -

Longhole dri ll - 5 - - - 3 2 - - - -

Bolter 3 - - - - 1 1 - - - -

Emulsion Truck 3 1 - - - 2 1 - - - -

Scissor Lift 4 - - - - 2 1 - - - -

LHD - Ejector - 7 m3 3 2 - - - 3 2 - - - -

Trucks 3 6 - - - 4 4 - - - -

Support Equipment Purchases

Low profi le U/G Motor Grader 1 - - - - 1 - - - - -

Crane truck 1 - - - - 1 - - - - -

Fuel/lube vehicle 2 - - - - 2 - - - - -

Service truck w/ scissor l ift 2 - - - - 2 - - - - -

Forklift/Cable reeler 1 - - - - 1 - - - - -

Mancarrier - 16 person 2 - - - 2 - - - 2 - -

Shotcrete Machine 2 - - - 2 - - - 2 - -

Mechanics truck w/ flat deck 2 - - - 2 - - - 2 - -

Supply truck w/ flat deck 2 - - - 2 - - - 2 - -

Crew cab 4 - - - 4 - - - 4 - -

Mine rescue van 1 - - - 1 - - - 1 - -

Surface Equipment Required

Grader 2 2 2 2 2 2 2 2 2 2 2

Water Truck 1 1 1 1 1 1 1 1 1 1 1

Compactor 1 1 1 1 1 1 1 1 1 1 1

Dozer 3 4 4 4 3 3 3 3 3 3 3

Excavator 1 2 2 2 2 2 2 2 2 2 1

Haul Truck 5 13 13 13 13 13 13 13 13 13 10

Surface Equipment Purchases

Grader 2 - - - - - 1 - - - -

Water Truck 1 - - - - - - - - - -

Compactor 1 - - - - - - - - - -

Dozer 3 1 - - - - 2 1 - - -

Excavator 1 1 - - - - 1 - - - -

Haul Truck 5 8 - - - - - - - - -





Project No.: 162230


Table 16-7: Mining Personnel Requirements Mine Operations Personnel Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10

Operators for Major Equipment

Jumbo - 2 boom 14 17 15 13 11 9 8 7 6 5 8

Longhole drill 0 14 14 14 15 15 15 15 16 16 13

Bolter 21 26 22 19 16 14 12 11 9 8 10

Emulsion Truck 9 13 11 10 9 8 8 7 6 6 7

Scissor Lift 14 17 15 13 11 9 8 7 6 5 8

LHD - Ejector - 7 m3 9 20 19 18 17 16 16 15 15 14 14

Trucks 9 34 32 31 31 30 30 30 29 29 26

Operators for Support Equipment

Low profile U/G Motor Grader 5 5 5 5 5 5 5 5 5 5 5

Service truck w/ scissor lift 9 9 9 9 9 9 9 9 9 9 9

Forklift/Cable reeler 3 3 3 3 3 3 3 3 3 3 3

Operators for Surface Equipment

Grader 9 10 9 9 9 9 9 9 9 9 9

Water Truck 5 5 5 5 5 5 5 5 5 5 5

Compactor 3 6 3 3 3 3 3 3 3 3 3

Dozer 14 23 23 18 14 14 14 14 14 14 14

Excavator 3 6 6 6 6 6 6 6 6 6 3

Haul Truck 23 59 59 59 59 59 59 59 59 59 50

Maintenance

Crane truck 3 3 3 3 3 3 3 3 3 3 3

Fuel/lube vehicle 5 9 9 9 9 9 9 9 9 9 7

Mechanics truck w/ flat deck 10 18 18 18 18 15 15 15 15 15 10

Supply truck w/ flat deck 10 18 18 18 18 15 15 15 15 15 10

Maintenance Shop 13 20 14 9 5 8 7 5 3 2 6

Mine Services

Services installations 5 5 5 5 5 5 5 5 5 5 3

Grouting/shotcrete 5 5 5 5 5 5 5 5 5 5 3

Construction 5 5 5 5 5 5 5 5 5 5 3

Level maintenance 2 2 2 2 2 2 2 2 2 2 2

Nipper 2 2 2 2 2 2 2 2 2 2 2

General Mine Administration, Technical and Services Mine Administration 19 19 19 19 19 19 19 19 19 19 15

Maintenance 11 11 11 11 11 11 11 11 11 11 8

Technical Services 23 23 23 23 23 23 23 23 23 23 20

Subtotal Mine Operations 263 407 384 364 348 336 331 324 317 312 279





Project No.: 162230


• The geotechnical investigation indicates that an extraction of 67.5% can be

achieved, resulting in an overall mining recovery of 58%. It was assumed that an

additional 2% of waste material would be added as external dilution and 2% of

mineralization losses were incurred due to operating conditions. An overall mining

recovery factor of 58% of the estimated Mineral Resources was considered in this

study; this accounts for 25.0 Mt of mineralized material as run-of-mine (ROM)

production to be processed inclusive of diluting material

• Considering waste inside stopes and external dilution, the overall Mineral

Resource estimate grades were diluted to 185 ppm of Ta2O5 and 1,591 ppm of

Nb2O5 for the ROM estimates

• Production was estimated at 2.7 Mt/a of mineralized material to be extracted over

10 years; the first year was considered as preproduction, leaving nine years of full-

scale production.

• Developments were modeled following a decreasing activity level from the

beginning to the end of the life of mine. A total of 92,500 m of development was

estimated.

• The deposit will be accessed through two main portals, Upper and Lower, and a

series of adits from the portals. Top access crosscuts will be driven from the main

ramps to each level on vertical intervals between 20 to 30 m. Stope access

crosscuts will be driven at level from west to east. Bottom access crosscuts will be

driven to function as mucking drifts. Underground mine services such as

ventilation and air heating, compressed air, water for drilling and power supply will

be provided to the mine via the adits.

• Radio remote-controlled load-haul-dump units (LHDs) will be used to extract the

mineralized material from the stope beyond the safety of the brow. The

mineralized material from stopes will be loaded directly to the haulage trucks that

will be spotted at the end of the crosscut. The trucks will drive down the ramps and

will exit the mine at the Lower Portal. The trucks will deliver the mined material to

a surface stockpile at the primary crusher close to the portal. Crushed material will

be transferred to the process plant by a belt conveyor.

• Waste from development will be initially utilized for construction of a structural shell

of the tailings co-disposal site on surface, which is located in an area east of the

processing plant site. This material will be transferred to the plant site in batches

on the conveyor system.

• Several areas of investigation are required to support a detailed mine plan. These

include underground geotechnical and geo-hydrological conditions for mine design,

the possible use of mining methods utilizing backfill, and handling systems for

mineralized material and waste rock. Opportunities include possible better ground





Project No.: 162230


conditions than assumed. With improved ground conditions the size of stopes and

production drifts could be increased.

• There is opportunity to increase the net present value (NPV) of the Project by

mining higher-grade zones early in the mine life providing that a practical mining

sequence can be implemented and the overall recovery of the Mineral Resources

is not negatively affected.

• There is opportunity to increase the net present value (NPV) of the Project by

optimising the mine layout to minimize development costs.





Project No.: 162230


17.0 RECOVERY METHODS

17.1 Plant Design

The design for the process facilities considered a nominal processing capacity of

7,500 t/d. Where data were not available at the time of flowsheet development, AMEC

developed criteria for sizing and selection of equipment based on comparable industry

applications, benchmarking, and the use of modern modelling and simulation

techniques.

The mineral processing and the refining are based on conventional technology and

industry-proven equipment.

Run-of-mine (ROM) mineralized material from the underground will be crushed and

conveyed to the concentrator where the mineralization will be ground to liberate the

mineral values from the host rock and then separated by flotation. The bulk tantalum–

niobium concentrate produced will be filtered, dried and introduced into the refining

plant. There the concentrate will undergo a thermal reduction which will remove most

of the gangue material and create a smaller, higher purity material for chlorine

processing. The distillation of the anhydrous metal chloride products will produce

high-purity Nb and Ta chlorides.

Tantalum chloride is the precursor to capacitor grade Ta powder, so would be

marketed in this form. Niobium chloride can be sold as a chemical precursor. Both Ta

and Nb chloride products can be readily converted and marketed as high-purity

technical grade Ta2O5 and Nb2O5 oxides respectively. The simplified flowsheet is

shown in Figure 17-1.

17.2 Comminution (Crushing, Storage and Grinding)

The primary crushing station will be a fixed jaw crusher. Mine haul trucks will dump

ROM mineralized material into the ROM surface stockpile located close to the Lower

Portal. Mineralized material will be fed to the crusher using an apron feeder. Crushed

mineralized material will fall onto a conveyor and be fed to a fine crushing circuit which

will further reduce the material to –8 mm. The material will be stored in a fine ore silo.

It will be withdrawn by feeder into a rod mill. The discharge from the rod mill will flow

into the cyclone feed pumpbox. The cyclone feed pump will transfer the material to the

cyclone circuit which will produce finished product in the overflow (a P80 size of

100 µm). The cyclone underflow will report to a ball mill for additional grinding. The

discharge from the ball mill will join the rod mill discharge as feed to the cyclone

pumpbox.





Project No.: 162230


Figure 17-1: Concentration and Refining of Blue River Mineralization

17.3 De-sliming and Flotation

After grinding, the flotation feed will be first de-slimed using high-frequency fine

screens or cyclones or a combination of the two. The fine product from the first de-

sliming step will be sent to tailings for disposal while the coarse product will undergo a

second step of de-sliming. The fine product of this second de-sliming step will also go

to tailings for disposal while the coarse product will be sent to flotation.





Project No.: 162230


The flotation area will consist of four steps; pyrrhotite flotation, carbonate flotation,

magnetite separation, and finally pyrochlore flotation. Each of the flotation steps will

include a conditioning step prior to the onset of flotation.

Pyrrhotite flotation will occur at a pH of 9.7 and will be performed with a promoter

(copper sulphate), a collector (Xanthate) and a frother (MIBC) to float off the contained

iron sulphides. Carbonate flotation will also occur at a pH of 9.7 and will involve a

dispersant (sodium silicate), a carbonate collector (oleic acid) and a frother (MIBC). In

both cases, the pyrrhotite and carbonate concentrates will join the de-sliming fines in

the tailings filtration system.

After carbonate flotation, the carbonate tails will be processed through a magnetite

separator to pull off any magnetite. The water will be exchanged at this point to allow

a higher level of control in the pyrochlore flotation.

Flotation of Nb-Ta-bearing minerals will occur at a pH of 7.0 employing a pH

modifier/promoter (fluorosilicic acid), a collector (a Duomac-T equivalent) and a frother

(MIBC) as required. The pyrochlore tails will pass to the tailings filtration system. The

pyrochlore rougher concentrate will be reground and cleaned in five stages with the

same reagents. The mass of material will be reduced substantially, to less than 1% of

the feed into the plant. All cleaner tails will be sent directly to the tailings filtration

system.

17.4 Filtration

After de-sliming, magnetic separation, and flotation, the combined tailings will be

pumped to two separate tailings thickeners for water recovery. After thickening, the

material will be pumped into one of four tailings pressure filters. These filters will

reduce the moisture to a level for disposal to drystacked tails.

Should a paste backfill option be considered during future studies, material after

filtration could be sent to a paste silo feed thickener. In this latter case, the material

could be withdrawn for use as required and transported to the underground mine portal

where it would be mixed with cement prior to use underground as backfill.

The pyrochlore concentrate product will be a much smaller mass and is first sent to a

small concentrate thickener and filtered.





Project No.: 162230


17.5 Concentrate Pre-Treatment

There are two options to pre-treat the flotation concentrate. If the concentrate grade is

between 10% and 30% Ta and Nb, then it is possible to perform a pre-leach with

strong acid to dissolve some of the gangue after which the material would be sent for

filtration. If the concentrate is more than 30% the non-acidified material would be sent

directly to filtration. After filtration, the concentrate would be dried and send to the

concentrate receiving bin.

Material would be recovered from the concentrate bin to a blender where flux (calcium

fluoride, quicklime), iron oxide, scrap aluminum and fuse mix would be added. After

blending the material is charged to burn pits. The aluminothermic reduction occurs in

these pits. After smelting and solidification, the alloy ingot (with adhering slag) is

removed from the burn pit by crane. After further cooling, the alloy ingot and adhering

slag would be broken, crushed and separated through the use of magnetic separation.

The slag material would be disposed of to tailings while the alloy material containing

the tantalum and niobium would be charged to the chlorination system.

17.6 Chlorination and Distillation

The material, which is charged into the chlorination system, will be set into fixed

charge pots where chlorine will be added and the mixture will be heated to 350º C.

The mixed chloride product from chlorination of the ferroalloy smelting product will then

be distilled to achieve high purity Ta and Nb chlorides. These distillation products will

be captured by separate condensers. The fumes will undergo hydrolysis to recover

chlorine.

17.7 Product/Materials Handling

A short conveyor is planned to transport materials from the portal to the plant.

17.8 Energy, Water and Process Materials Requirements

Power for the proposed operation may be sourced from B.C. Hydro. Power needs will

be investigated during more detailed Project studies.


In the opinion of the QPs, metallurgical programs completed on the Blue River Project

have met their objective of identifying a processing method allowing for the extraction





Project No.: 162230


of tantalum and niobium mineralization that has reasonable prospects of being

economic. Additional work is required to confirm the process, produce the target

flotation concentrate grade and examine the response of the process to variability of

mineralization within the deposit and process conditions. The following interpretations

apply to the plant design and metallurgical testwork results:

• Tantalum and niobium occur as ferrocolumbite and pyrochlore, which are

amenable to conventional flotation and proven refining processes with estimated

recoveries of 65% to 70%. For the purposes of the financial analysis in Section 22

of this Report, it was assumed that the process plant will have a 65% recovery for

Ta and 69% recovery for Nb in the flotation stage. The refining process will have a

97% recovery for both Ta and Nb

• Optimization of the supply and pricing of reagents for the refining process may

support lower operating cost assumptions.

• Metallurgical testing has not yet attempted to demonstrate that a 30% combined

oxided concentrate grade as a feed for the refining stage is achievable







Project No.: 162230


18.0 PROJECT INFRASTRUCTURE

18.1 Site Layout

The overall Project site layout plan is included as Figure 18-1. The planned Upper and

Lower Portals will be located about 4 km from the plant site. At the front of the Upper

Portal (service portal) sufficient space will be provided to accommodate the required

facilities for operation.

18.2 Buildings

18.2.1 Mine Service Building

The plant service building will be a multi-purpose complex in a two story building east

of the process building. The first floor (18 m by 40 m) will be a maintenance bay for

minor repairs and maintenance of the mobile equipment and will have an office for the

maintenance foreman, a small parts area, a tool crib, and a storage area for safety

equipment.

Part of the first floor will contain the men’s and women’s dry. The second floor will be

the administration offices. The complex will be connected to the process building by a

covered walkway.

18.2.2 Truck shop

A 24 m by 36 m truck shop on the southwest end of the site will be operated by a

qualified contractor. The required equipment and tools for regular maintenance and

possible repair of the haul trucks are assumed to be supplied by the contractor.

18.2.3 Warehouse

To save cost and construction time, the warehouse will be a 24 m by 50 m Coverall-

type fabric building. One third or more of the building will be dedicated to warehouse

cold storage, and the remainder allocated to fire water and potable water tanks and a

fire pump skid. The building will be equipped with an interior liner as an insulation

layer to minimize the potential of freezing inside the warehouse. This will save the

cost of insulation and heat tracing of the tanks, pipes and all equipment and is easier

for operation and maintenance.





Project No.: 162230


Figure 18-1: Proposed Site Layout Plan





Project No.: 162230


Forklifts, pallet racking, bins, and carousels will be provided for handling materials.

Flammable products such as solvents and paints will be stored separately.

The Coverall-type building life expectancy is 15 years and it will withstand design snow

and wind loads.

18.2.4 Process Building

The process building will be a 22 m by 52 m steel structure building sitting on a

concrete foundation. The mill foundation will be on bed rock; a geotechnical

investigation is required prior to determination of the final location of the mill.

A 15 m diameter tailing thickener will be located to the south of the process building

with a walkway/pipe rack connection to the building.

18.2.5 Crushing and Screening Circuit

The preliminary site layout is designed to take advantage of the topography to

minimize earthworks. The secondary crusher will be located in the northeast corner of

the site at a higher elevation.

The conveyor span from the secondary crusher discharge will rise at 12º (an access

safety constraint) to reach the height of the screen above the fine ore bin.

Screen oversize will pass via a return conveyor to the tertiary crusher.

18.2.6 Portal Infrastructure

A yard will be constructed in front of the Upper Portal, which will accommodate an

electrical substation and generator set, and office building, provision for storage,

ventilation infrastructure, and heater, air compressor, diesel storage, a first aid rescue

vehicle bay and a water tank.

The buildings and related facilities will be pre-engineered as much as possible.

Drinking water will be provided from the plant site water treatment plant by containers.

A portable wash room will be provided for workers. Sewage will trucked to the site

wastewater treatment plant.

18.2.7 Explosives Storage

Ammonium nitrate, blended emulsion, and explosives will be delivered to site on

demand by contractors. A small storage magazine will be constructed at a distance of





Project No.: 162230


about 200 m from the Upper Portal. Room for explosive storage will be provided by

excavating into the rock. The walls and roof will be reinforced and a lockable door will

be provided, as per the requirements of the Quantity-Distance Principles User’s

Manual published by the Explosives Regulatory Division of NRCan. The magazine will

hold boosters, delays, detonating cords, detonating caps, and other explosive

accessories.

18.2.8 Aggregate Crushing and Concrete Batch Plants

A crushing and stockpiling facility will be required during construction to provide

crushed product for roads and surfacing. The mobile plant assembly will include a jaw

crusher, screening plant, closed-circuit secondary crushing unit, and washing plant.

Concrete supply from nearby towns is assumed adequate for construction purposes.

An on-site concrete batch plant is not proposed for the project. Availability of existing

concrete supply should be examined in the next phase of study.

18.3 Roads and Logistics

18.3.1 Access Road

The road access design includes a short new road with a 7.2 m wide gravel surface

from the existing road to plant site about 80 m in length and a 1.5 km new service road

from the existing road to the Upper Portal and upgrades to the current access road.

An existing 80 m-long bridge crossing over the Thompson River has a limited load

capacity and might not qualify for crossing heavy loads during construction or long-

term use during the life of the mine. Therefore a new bridge has been included in

capital cost estimate. Using the existing railway for shipment should be investigated in

next phase of study.

18.3.2 Haul Road

Dual-lane traffic requires a travel width (16.2 m) of not less than three times the width

of the widest haulage vehicle used on the road (assumed to be trucks of the size of

CAT 775F with 5.4 m of overall width).

Single-lane traffic requires a travel width (11 m) of not less than two times the width of

the widest haulage vehicle used on the road.

Shoulder barriers should be at least three-quarters of the height of the largest tire on

any vehicle hauling on the road wherever a drop-off greater than 3 m exists. The





Project No.: 162230


shoulder barriers are designed at 1.5:1 (H:V). The width of the barrier is excluded

from the travel width.

There is one main haul road, from the site to the waste rock stock pile, that will have a

length of about 8 km (refer to Figure 18-1). The road from the upper portal to the plant

will be a service road for personnel and supplies.

18.4 Co-Disposal Storage Facilities

The PEA design for tailings and waste management is to construct a co-disposal

drystack facility.

18.4.1 Drystack Considerations

Filtered tailings stacks are often referred to as “drystacks” and that nomenclature is

used in this chapter. However, these facilities are not “dry” per se as the tailings, while

placed in an unsaturated state, do have moisture contents that are typically 70% to

85% of saturation. Generally the tailings are filtered to within a few percent of optimum

standard Proctor moisture content, which is typically on the order of 15 % (wt water :

wt solids). The main distinguishing feature from other tailings deposits is that filtered

tailings stacks are a solid rather than the more typical slurry and need to be

transported using mechanical means versus hydraulic methods.

Tailings drystacks are particularly suited to locations with flat or gently sloping storage

sites and arid environments where it is relatively easy to place and compact filtered

tailings into a stable geometry. However, unlike conventional slurried tailings, tailings

drystacks can also be placed on relatively steep terrain in a similar manner to

conventional waste rock facilities. Placement practices and drystack design can be

adapted to different environments and drystacks have been successfully constructed

in cold climates.

The Blue River site is expected to have about 1.5 m of precipitation annually (KCB

2009a). The site experiences a cold winter and steep topography which can present

challenges for construction of a tailings drystack. AMEC notes, however, there are

operating drystacks in even wetter environments.

The general drystack concept for the Blue River site is to have an outer shell zone with

a general tailings placement area located upstream of the shell. The shell zone would

consist of well-compacted tailings placed only when it can be assured that such

compaction can be achieved. The waste rock could be placed inter-layered or mixed

with the tailings in the shell or could be used as armour on the face of the shell to

prevent erosion of the tailings surface. The outer shell would support a general tailings





Project No.: 162230


placement area where tailings could be placed in poor weather and with less ability to

achieve compaction during winter months. The same operating attempts at

compaction would be made in winter/wet weather conditions but it is more difficult to

get assured densities. As a consequence, having a general placement area for such

materials where structural integrity of the overall stack is not at jeopardy if lower

densities are achieved is sound tailings management. It would also be possible to

store waste rock in the general placement area where it could be encapsulated within

the tailings. In order to effectively co-dispose of waste rock in the general fill area,

waste rock and tailings disposal would have to be appropriately timed and managed.

Typically, slopes of tailings drystacks are designed to be 3H:1V or less to minimize

erosion of the tailings on the downstream slope. This is particularly a consideration for

closure.

Steeper slopes will be required at the Blue River site due to topographical constraints.

The steeper slopes are achievable but require a potentially wider shell zone of good

compaction, confirmed foundation conditions for the shell and the slopes will require

erosion protection.

18.4.2 Evaluation of Potential Sites

A series of tailings storage location screening assessments were carried out during

2008, 2009, and 2010 (KCB 2009a, 2009b, AMEC 2010a, 2010b, 2010c), focusing

mainly on conventional tailings storage. Some evaluation of potential waste rock

storage sites and a tailings drystack facility was undertaken in the 2009 studies

The initial 2009 study (KCB 2009a) focused on the area in the immediate vicinity of the

deposits and was based on a desk study using available topography and other

information. Five conventional storage sites, using containment dams, were assessed.

The potential tailings drystack locations identified by KCB in 2009 are located on valley

sidehills and in flatter areas adjacent to the North Thompson River The tailings storage

facilities (TSFs) were assumed to be constructed of local borrow with slopes of

2.5H:1V and a settled density of 1.3 t/m3 was assumed for the tailings.

The second study carried out in 2009 (KCB 2009b) reviewed a number of alternative

sites for TSF storage away from the proposed mine site. The main focus of the study

was an industrial land parcel near Valemount, BC, and consideration was given to

construction of a 40 Mt TSF, a plant, 30 kt of rock storage, and a rail-siding facility on

the property. Four other areas around Valemount were also considered as TSF

alternatives.





Project No.: 162230


None of the options were considered ideal and additional studies were carried out over

an increasingly large area and with varying constraints to identify potential tailings

storage areas (AMEC 2010a, b, and c). Generally topographic and climatic conditions

are challenging for surface storage of tailings and no ideal location has been identified.

Given the current stage of project development and the constraints imposed, the input

provided should be considered as conceptual. Tailings storage site locations

determined by Klohn Crippen Berger (KCB) in 2009 were visually identified by AMEC

staff during a reconnaissance flight in July 2010 and were visited on foot by AMEC

personnel.

However, the proposed tailings storage sites have not been visited by geotechnical

personnel, there have been no site investigations and there have been no specific

technical analyses in support of facility layout. The suggestions presented in the PEA

for tailings storage were based on engineering judgment, AMEC’s experience with

filtered tailings, and standard industry practice for similar facilities.

18.4.3 Site Selection

Additional review for the PEA indicated that a site identified by KCB in 2009, termed

WSF3 and shown in Figure 18-1 as the co-disposal site, could be utilized, since the

volume of tailings storage had been revised downward from about 30 Mt as

conceptualized in 2009 to 22 Mt in the PEA. The facility was moved uphill slightly

relative to the 2009 study to take advantage of locally flatter areas and reduce the

footprint of the facility in steeper areas.

Additional optimization of the facility layout could be carried out in consideration of

local topography and stability; however the volumes and areas would not be expected

to be significantly altered.

18.4.4 Facility Design

The facility was laid out with 2H:1V slopes. Attempts were made to use flatter slopes;

however, due to the topography, they were not feasible. If the project is advanced to a

later stage, the slopes could potentially be locally flattened in specific areas of flatter

topography. The facility will have the following dimensions:

• Total storage volume: 20.9 Mm3

• Total footprint area: 534,000 m2

• Total surface area of drystack: 559,000 m2

• Volume of sloped (shell) portion of facility (shell portion of facility): 13.2 Mm3





Project No.: 162230


• Footprint surface area under slope (shell): 385,000 m2

• Surface area of sloped (shell) portion of tailings: 416,000 m2.

Conceptually, the sloped portion of the drystack would form a structural shell,

supporting the general tailings placement area behind it. The shell zone would consist

of compacted filtered tailings possibly interlayered or mixed with waste rock. Stringent

placement control and compaction would be required in the shell. It is possible that

the limits of the shell zone could be optimized at a later stage. The general tailings

placement zone would allow for placement in wet and cold weather with less assured

compaction though identical operating practices to those required in the shell area

would be prescribed and followed.

Depending on the timing of any mine start-up and the tailings production schedule, it

may be necessary to construct a starter berm out of non-tailings material to provide

sufficient storage for the first winter of tailings placement. The starter berm could be

constructed of non acid generating waste rock, general rockfill or granular overburden

obtained from a local borrow source.

Based on the volumetric proportions of the shell and general placement areas and

assuming a six-month period available for shell construction, there may be a deficit

with respect to the amount of tailings and waste rock available to construct the shell

zone at certain points during operation. It is likely that the configuration of the facility

could be optimized during design to overcome this deficit. If supplemental fill material

were required in the shell, it could consist of general rock fill or granular overburden.

The option of storing some of the tailings produced during the winter for placement in

the drystack during the summer construction season could also be considered.

18.4.5 Co-Disposal Facility Geohazards Considerations

The area in the vicinity of the deposits has steep-sided valleys with glaciers located in

the upper portions of many of the catchments and visible avalanche tracks. The

WSF3 site is located on a side-hill near the mouth of the valley, away from the glaciers

located farther up the valley. It is also located on a spur minimizing the risk of

avalanches. A site-specific geohazards evaluation for the site has not been

completed, however, and should be carried out as the Project is advanced.

18.4.6 Co-Disposal Facility Stability Considerations

The 2H:1V slope required because the topography at the Blue River site is steep

compared to typical tailings drystacks. However, in such situations, flatter slopes are

often developed to minimize operating and closure erosion concerns. From a purely





Project No.: 162230


structural stability perspective, provided the shell zone can be developed of

appropriate size while meeting compaction criteria, the use of 2H:1V slopes is

acceptable. Additionally, the general fill zone behind the shell is intended to allow

placement of tailings during poor weather and will potentially have lower strength and

likely areas of wetter tailings. The shell zone is therefore necessary to support the

facility and will require significant compaction and placement control to provide

sufficient resistance.

The foundation is expected to consist of colluvium and/or silty to sandy tills which

should not be a concern with regards to overall stability, although this will require

confirmation as the project is advanced. A toe berm or shear key may be required for

stability of the drystack depending on specific foundation conditions. This should be

assessed in future design stages.

The Project is located in a moderately seismic area (KCB 2009a) and stability under

seismic loading will have to be considered during design of the facility. Because of

their unsaturated nature, the general placement tailings are considered unlikely to

liquefy and the shell tailings will not present a concern because the compacted nature

of the downstream shell will result in dilative behaviour under shear further improving

liquefaction resistance. Rain, ice and snow may potentially create isolated areas in the

general placement area that are saturated and poorly compacted, and therefore locally

susceptible to liquefaction. It is highly unlikely that the entire zone would liquefy and

over time the pore water in any saturated zones would seep out of the stack, leaving it

unsaturated and resistant to liquefaction in the long term. Although liquefaction is not

considered a concern, deformations due to increased loading during seismic events

will occur and should be quantified during detailed design phases.

For the Project base case it is assumed that the tailings facility does not require lining.

A requirement to line the facility would involve clearing and stripping the entire footprint

at start up, placing and compacting a bedding layer over the entire footprint, installing

the liner system over the entire footprint, and placing a protective cover layer over the

entire liner.

18.4.7 Co-Disposal Facility Surface Water Run-off Considerations

Two surface water management systems will likely be required for the tailings

drystack.

• The first system would divert non-contact (clean) water around the facility.

• The second would collect run-off water which had been in contact with the tailings

drystack area.





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Contact run-off water would have to be collected and potentially treated prior to

release, whereas the clean water would be diverted around the facility and into the

creeks downstream of the tailings facility.

The non-contact water diversion system would consist of a ditch located beyond the

final footprint of the drystack and would be a long-term structure. The contact water

collection system would consist of a ditch around the perimeter of the drystack

footprint and would be reconstructed annually, so as to be located slightly ahead of the

advancing drystack footprint. The perimeter ditch would require sediment-control

structures within it to help contain any tailings mobilized from the drystack. The

perimeter ditch would direct the contact run-off water to a collection and sedimentation

pond where eroded tailings could settle out and water treatment could be carried out if

required.

Located as it is on a side-slope and near the nose of the ridge, there is little catchment

area uphill of the drystack facility and it may be feasible to combine the two water

management systems, particularly upon closure. Combining the two systems would

increase the volume of water collected and potentially requiring treatment, however,

the increased water volume may also contribute to dilution of the contact water.

18.4.8 Co-Disposal Facility Closure Considerations

At mine closure, the surface of the drystack would be sloped/contoured for drainage,

covered with an appropriate material and vegetated to enhance erosion protection.

The shell face will require armouring to reduce erosion. The perimeter ditches will be

increased in size and lined with rip-rap following mine closure to minimize the amount

of run-off water that comes into contact with the stack, further reducing the potential for

erosion. Depending on the chemistry and flows, it may be possible to combine the

water diversion ditch with the closure perimeter ditch.

The Project base case is that the uranium and thorium levels are sufficiently low to not

be a concern. If the levels were found to be an issue, however, it would affect closure

requirements. In addition to the potential difficulty of treating any seepage, if radon

gas emission was above regulatory levels, the cover system would have to be

designed to contain radon gas. Specific details of the cover would be based on

applicable regulations but it may well require that the cover last for a very long period

of time necessitating rock armouring among other considerations.





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18.5 Avalanche Hazard

No allowance for avalanche protection for any infrastructure has been considered at

this stage, but considering the steep slope and possible heavy snow at the area an

avalanche study is recommended during more detailed studies.

18.6 Water Supply, Distribution, and Treatment Systems

The potable water system and layout for the process and administrative area is

designed to service buildings and a workforce of 120 persons. Potable water for the

mining section will be constructed as part of the portal and underground infrastructure.

Raw water will be provided from a well and a prefabricated water treatment module will

treat water to standard drinking water requirements. Treated water will be stored in a

potable water tank.

Fire protection water for the construction camp and later for all buildings will be

provided by a prefabricated diesel-driven fire and electric jockey pump on a skid that

will be located in the coverall building.

Both the potable and the dedicated fire water tanks will be located under the coverall

building adjacent to the fire pump skid and the water treatment module.

Potable and fire water will be distributed to the plant buildings through separate pipes.

All water mains will be buried to a depth in excess of 3 m, or will be insulated providing

an equivalent degree of cover to prevent freezing.

18.7 Waste Considerations

Waste-water treatment sludge will be trucked away to a nearby municipal facility or

approved landfill.

Waste lubrication and hydraulic oils from vehicle maintenance will be stored in

dedicated tanks and sent to a recycling facility offsite. Their disposal will be contracted

to an approved contractor.

A modular sewage treatment system will be installed as part of the initial construction

infrastructure. A small package treatment plant will provide treatment to the domestic

sewage at the site. Effluent from this plant will meet specified water discharge

guidelines prior to discharge into the environment.





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Buried gravity sewer lines and manholes will collect and direct sewage from the

service building and truck shop to an equalization tank adjacent to the wastewater

treatment module for treatment.

18.8 Accommodation

Contractors and employees will commute from the nearby towns, such as Blue River

and Valemount, during construction.

No on-site permanent accommodation will be provided for personnel. It is assumed

that the workforce, including management staff, will reside in the nearby communities

and will commute, via buses, on a daily basis. For safety reasons, no private vehicles

will be permitted on the site access road or at the site.

18.9 Power and Electrical

Power supplies in the region have been assumed to be sufficient for Project

requirements, and no allocation for additional power line construction has been

included.

The BC Hydro 136,000 volt supply line for the North Thompson valley passes through

the west side of the property adjacent to the rail line. The 20 megawatt Bone Creek

run-of-river hydroelectricity project, owned by Transalta Corp., was commissioned in

June 2011, and is adjacent to the Project area its powerhouse is located approximately

4.4 km south of the Lower Portal.

The mill will be the greatest consumer of power. The high voltage line from the grid

would go to a main substation close to the mill. From this sub-station, lower voltage

power will be distributed to the mill, offices, maintenance shop, and other

infrastructure, and to the proposed mine substation. The PEA assumes power will be

supplied to the mine via the portals. The sub-station for the main power distribution

system and the air compressors will be installed in facilities located adjacent to the

Upper Portal.

18.10 Fuel

Fuel will be delivered to the mine site using tanker trucks. The fuel storage tanks will

be single-walled within a lined containment berm. Tank design will comply with the

appropriate regulatory requirements.





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• Facilities to support mine operations will require construction

• Infrastructure envisaged includes a plant, plant service building, truckshop, potable

and process water systems, a sewerage system, co-disposal site, underground

mining operation, conveyor system, and various haul and access roads. The

planned Upper and Lower Portals will be located about 4 km from the plant site.

The co-disposal facility will be about 8 km from the plant site

• The infrastructure contemplated in the PEA is appropriate to planned throughput

rate, treatment plant, and mining method

• No on-site permanent accommodation will be provided for personnel. It is

assumed that the workforce, including management staff, will reside in the nearby

communities and will commute, via buses, on a daily basis

• Geohazards are present in the area, and will require careful consideration in future

studies

• Water management studies, in particular for the co-disposal site, will be required.





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19.0 MARKET STUDIES AND CONTRACTS Commerce has prepared assessments of the tantalum and niobium markets which

outline the supply and demand for tantalum and niobium. The tantalum assessment

was prepared by a tantalum market expert, although he is not independent of

Commerce. His analysis reflects the general consensus of other analysts regarding

the tantalum market expressed in publicly-available information.

The niobium assessment was prepared by an independent niobium expert and also

reflects the general consensus of analysts in publicly-available information for the

niobium market.





In the opinion of the QPs, the following conclusions can be drawn from the marketing

strategy used to support the PEA:

• Commerce has prepared assessments of the tantalum and niobium markets which

outline the supply and demand for tantalum and niobium

• As the Project is still at an early evaluation stage, Commerce has not initiated

requests from potential buyers for expression of interests from potential buyers of

the proposed Blue River products and has not negotiated any purchase or off-take

agreements

• Contracts that will be negotiated are expected to be within industry norms.





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20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND

SOCIAL OR COMMUNITY IMPACT

20.1 Environmental Assessment for Mining Projects

The Blue River Project will require approval under the Federal and Provincial


and authorizations for construction and mine operation. This section discusses the

environmental assessment and permitting process as it stands today, and describes

the principal licences and permits which would be required for the Blue River Project.

The British Columbia Environmental Assessment Office (BCEAO) and the Canadian

Environmental Assessment Agency (CEAA) would both conduct an environmental

review of the Upper Fir Project, as defined respectively under the BC Environmental

Management Act and the Canadian Environmental Assessment Act.

Overall the environmental review of a project is a process that will take up to

18 months to complete. The process would include the development of several

important documents by Commerce, including the Project Description, Assessment

Information Requirements and an Environmental Impact Assessment application,

followed by the review of these documents by the public, interested stakeholders, First

Nations and regulators.

Both the Provincial and Federal processes have defined timelines for project review

though these timelines are not currently harmonized and past attempts to do so have

not been overly successful. The Federal timeline is the longer of the two at 365 days

to review the Environmental Assessment application and make a decision, while the

provincial application review stage is 180 days plus 45 days for decision. Additionally,

the Federal clock is stopped each time CEAA submits comments to the proponent for

review, to respond to, or revise; this can possibly extend the time line.

There is also a need for additional time on the front end of the process to develop the

Project Description. This document must first be accepted by the regulators, and then

will be used to develop Assessment Information Requirements or Terms of Reference

which must be put out for public review prior to acceptance. These Information

Requirements will define the content required for the Environmental Assessment. The

Federal government allows 90 days for this, while the Provincial government has no

timeline on this process.

The environmental review and assessment process results in a decision with respect

to whether or not the Project should be issued an Environmental Assessment





Project No.: 162230


Certificate by the Provincial government, as well as receive Federal Ministerial

approval based on the recommendations put forth to the Minister of Environment in a

Comprehensive Study Report prepared by the Major Projects Management Office.

Both are required for a project to proceed to permitting and development.

20.2 Project Studies


at the Blue River Project site since the summer season of 2006 with the selection of

local and regional studies areas for each biophysical discipline.

Field studies completed by specialist consultants independent of Commerce

Resources include:

• Site hydrology (2006–present)

• Snow course depths (2007, 2008)

• Fisheries and aquatics (2006–2008)

• Soils, flora and fauna assessments (2006, 2007), including studies of rare,

threatened and endangered plants (2007), breeding birds (2007) and terrestrial

ecosystem mapping (2006, 2007), wildlife studies and habitat suitability mapping

(2006–2008);

• Geochemistry, mineralized material and waste rock characterization with baseline

ABA and metals analyses (2007, 2008)

• Surface water and sediment quality (2006–present)

• Groundwater (2007–2009)

• Terrain stability assessment for roads (2007–present).

Kinetic test work for ARD/ML was initiated in June 2010 and is ongoing; results of this

work will give an indication of the type of management strategies required for handling

PAG waste rock.

Additional environmental baseline programs are expected to continue, as required

through 2011.

Monitoring of meteorology, air quality, hydrology, and water quality will continue

throughout the construction, operation, closure and post-closure phases.





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It is anticipated that with this work and the results of the PEA in hand, Commerce will

have the necessary information to start development of a draft Project Description

which is a prerequisite to entering the Environmental Assessment process. These

data will provide a strong base to initiate meetings with the BCEAO and CEAA, as well

as with key provincial regulators such as Ministries of Energy and Mines, Forests

Lands and Natural Resource Operations, and Environment, to discuss specific project

requirements under the Provincial and Federal environmental assessment processes.

Summaries will be prepared of baseline data collected and work plans as appropriate

for submission, review and input by Federal and Provincial regulators.

20.3 Environmental Setting and Review of Environmental Baseline

Characterization of existing environmental conditions, which began in 2006, is an

important component of the risk management and permitting process for the Blue

River Project.

The Blue River Project area is located within the B.C. Ministry of Environment

Thompson-Nicola Region (Region 3), the Ministry of Forests and Range Headwaters

Forest District and Fisheries and Oceans Canada Sub-district 29J (Clearwater). It falls

towards the northern end of the Kamloops Land Resource Management Plan (LRMP)

which was approved by the province in 1995. This LRMP is the first plan of its kind in

British Columbia in that it is a locally-developed plan that is designed to guide land and

resource management decisions in a way to balance community needs, environmental

concerns and economic values. This LRMP is termed a sub-regional integrated land

use plan in that it establishes the framework for land use and resource management

objectives and strategies. The Plan requires that more detailed operational plans

which are subsequently developed be consistent with the management strategies and

objectives defined in the LRMP.

Following implementation of the BC Mountain Caribou Recovery Strategy, site-specific

objectives and strategies for caribou management were developed (first in 2006, then

updated in 2009) which included objectives pertaining to mineral exploration. The Blue

River Project area falls within the Wells Gray-Thompson caribou planning unit (unit

4A). In this case, the relevant portion of the LRMP was repealed and replaced under a

Government Actions Regulation (GAR) by a caribou-management strategy, specifically

the identification of Ungulate Winter Range (UWR) zones. The GAR order states that

exploration and mine development activities within the UWR are considered by the

government to be an acceptable risk to caribou, and are allowed to proceed without

requiring an exemption from the Ministry of Environment.





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The bulk of the infrastructure proposed for the Blue River Project falls within an area

identified as a Modified Harvest Zone. In such zones, operational activity is expected

to be considerate of caribou habitat and disturbance in caribou areas. Commerce is

committed to seeking specialized professional advice to minimize or eliminate

disturbances to caribou using best management or other practices. Wildlife studies to

date suggest that caribou are recorded only rarely in the area of the proposed Project.

The larger Blue River Project area lies on the eastern side of the south-flowing North

Thompson River where claims held encompass portions of the Bone, Gum,

Moonbeam, Paradise Lake, Pyramid, and Serpentine Creek watersheds. The overall

relief is moderate to steep, with an average elevation of 1,625 m, a maximum elevation

of 3,225 m, and a minimum elevation of 580 m. Some small glaciers exist in the

easternmost part of the Project area, and moderate to steep forested slopes rise

above the North Thompson River valley. The North Thompson River drainage

continues south to join the South Thompson River at Kamloops, BC.

Proposed Project infrastructure is located along a western-facing slope immediately

above a gravelly part of the North Thompson River valley, and includes portions of the

Bone and Gum Creek watersheds as well as residual areas draining directly into the

North Thompson River. The tree line is located at approximately 2,000 m elevation,

and the Upper Fir deposit centre is located at 1,180 m elevation along a network of

previously-constructed logging and skid roads. Much of the area of the proposed

Project infrastructure, including the Upper Fir deposit had been logged prior to the

commencement of mineral exploration. Naturally-occurring outcrop is generally poor,

with limited exposure of underlying country rocks along road cuts and locally in

streams.

The area has a continental climate which is subject to frequent modification by

maritime air masses from the Pacific Ocean. The area is part of a "wet belt" which

occupies part of eastern British Columbia. Heavy snow falls occur almost every

winter, in which temperatures stay close to the freezing point when maritime air

dominates. The most severe cold spells may send thermometer readings below -

40°C/F. Rain is frequent in other seasons. Summer days are warm or occasionally

hot, with thunderstorms often spawning over the nearby mountains. The optimal field

exploration season is mid-June through mid-September.

Streams within the Project Area are generally characterized by a snowmelt-dominated

peak rising in April or May and peaking sometime between June and July. Rain-on-

snow events occasionally occur in this region and these can enhance both winter flows

and spring peaks. In addition, late fall rainstorms are common, recharging soil

moisture heading into winter and producing short-duration peak flows. Low flows





Project No.: 162230


occur generally from the end of November to March, and in hot summer months, with

the lowest flows commonly occurring in January or February. No wetlands have been

identified within areas of proposed Project infrastructure.

Surficial materials were typed for texture, drainage, moisture and nutrient regime, and

parent material. This not only supports ecosystem classification, but also provides

baseline soils data for eventual environmental impact assessment and reclamation

planning. Materials range from colluvial veneers to fluvial plains, glaciofluvial terraces,

rock, and morainal blankets. On the Upper Fir slope, morainal materials are most

common, while fluvial and glaciofluvial deposits are limited to lower elevations near the

valley bottom. Silts and sands are the most prevalent soil textures, although mixed

fragments, rubble, and gravel also occur. Overall the deposits are relatively shallow,

although blankets, which have more than 1 m of surficial materials (e.g., glacial till),

are much more common than veneers, which have less than 1 m of surficial materials.

Soil moisture and nutrient regime are generally average over the majority of the site

and have formed in place within morainal and glaciofluvial landforms.

Brunisols are the most dominant soil order found, while Podzols are secondary,

becoming more prominent in areas with increasing rainfall and elevations above

1,500 m. Soil quality and quantity appear to be adequate to support soil salvage and

reclamation activities in the area of the proposed development. Soil fertility suggests

normal levels of soil nutrients as compared to other mine sites in BC. Soils are

moderately acidic (i.e., pH 4-5.5) and considered normal for mesic, conifer-dominated,

forested vegetation in similar areas. Soils are very rapidly to imperfectly-drained with

soil moisture regimes ranging from sub-xeric to hygric and soil nutrient regimes limited

to moderate and rich.

Terrestrial Ecosystem Mapping (TEM) was completed to describe terrestrial

ecosystems according to the bioterrain base and standards established by British

Columbia’s Resource Information Standards Committee. The Blue River Project falls

within the Cariboo Mountain Ecosection of the Northern Columbia Mountains. Two

biogeoclimatic zones are found in the project study area. These are the Interior Cedar

Hemlock (ICH) zone, which occurs at lower elevations, and the Engelmann Spruce

Subalpine Fir (ESSF) zone which occurs at higher elevations above the ICH zone.

Biogeoclimatic zones, subzones and variants within the Study Area were classified

using the Ministry of Forests Biogeoclimatic Ecosystem system. The following

subzones/variants are present within the larger project study area:

• Wells Gray Wet Cool Interior Cedar – Hemlock Variant (ICHwk1)

• Mica Very Wet Cool Interior Cedar – Hemlock Variant (ICHvk1)





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• Northern Monashee Wet Cold Engelmann Spruce – Subalpine Fir Variant

(ESSFwc2)

• Wet Cold Engelmann Spruce –Subalpine Fir Woodland Subzone (ESSFwcw)

• Wet Cold Engelmann Spruce – Subalpine Fir Parkland Subzone.

Much of the area is forested, although avalanche chutes punctuate the landscape at

seemingly regular intervals and forest harvesting has been extensive at lower to mid

elevations.

In terms of vegetation, the area directly covered by the Blue River Project is relatively

small and generally characterized by common plant communities associated with five

biogeoclimatic subzones (ICHwk1, ICHvk1, ESSFwc2, ESSFwcw and ESSFwcp).

Based on an assessment of biogeoclimatic units in the Study Area and the BC

Conservation Data Centre species at risk list for the Headwaters Forest District (BC

Conservation Data Centre 2006), at least 37 ranked plants may occur within the larger

study area. However, a field study of rare vascular plants identified only four

populations of two Provincially-listed rare plants (Galium trifidum ssp. trifidum and

Carex paysonis) occurring outside the current project envelope. No Federally-listed

plant species or plant communities were identified.

The region encompassing the proposed Project is likely home to many terrestrial

wildlife species including black and grizzly bears, deer, moose and mountain goats;

birds are likely to include osprey, eagle, woodpecker and raven, migratory songbirds,

raptors; and numerous small mammals.

Wildlife species of concern, whose confirmed distribution intersects that of the Blue

River Project, include the blue-listed grizzly bear (Ursus arctos) and red-listed

mountain caribou (Rangifer tarandus caribou). Mountain caribou presence on the

larger area of the mining claims making up the property and in the general area has

been confirmed through ongoing government radio-telemetry studies. Mountain

caribou are a Federally- and Provincially-listed species and are of considerable

concern to the public. They have the greatest potential to interact with the Project

property in early winter.

Due to deep snow pack in the region, and considerable management and local

interest, other species of concern include mountain goat (Oreamnos americanus) and

moose (Alces alces). Mountain goat winter range exists throughout the area, with

occupied ranges within 2 km of the area of potential Project development. Moose are

the most heavily-hunted ungulate in the area. Moose winter range occurs throughout

the North Thompson valley, with valley wetlands and early seral stage habitats of

prime importance.





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Habitat suitability was assessed for wildlife focal species selected based on their at-

risk status under the British Columbia Conservation Data Centre (CDC) and the

Committee on the Status of Endangered Wildlife in Canada (COSEWIC) systems, and

their level of local concern. Species include mountain caribou (Rangifer tarandus

tarandus; southern populations), moose (Alces alces), grizzly bear (Ursus arctos),

mountain goat (Oreamnos americanus), and marten (Martes americana). Specific

habitat notes were also recorded for other species such as deer (Odocoileus sp.), elk

(Cervus canadensis), wolves (Canis lupis), and black bear (Ursus americanus).

Habitat ratings suggest moderate and high suitability for mountain caribou during the

early winter season in the ICH. Moderate ratings were also assigned to marten in

many sites in the ICH. No sites were rated as high or moderate suitability for mountain

goat in the immediate areas of Project infrastructure.

Regional and site specific fisheries studies show that bull trout and mountain whitefish

are utilizing lower Gum Creek for rearing. Bull trout consisted of both juveniles and

young-of-the-year suggesting that it is being used for spawning by this species. The

lower reach of Bone Creek is being utilized by coho salmon, parr, and torrent sculpin.

Benthic invertebrate data were also collected. Habitat available for fish within Gum

and Bone creeks is limited to their lower most reaches, near their mouths. Gum Creek

fish habitat use is limited to the lower most portion of the creek, from the mouth to

600 m upstream before a falls/gradient barrier (>20%) and fish distribution in Bone

Creek is limited to the section from its confluence with the North Thompson River to

approximately 2,100 m upstream before an impassable water fall.

Water quality studies were conducted at various sites within the project area from 2006

to the present, with the objective of providing a long-term record of the relative

chemical stability of the project area. Samples were analyzed for physical variables,

anions, nutrients, total organic carbon and total and dissolved metals. Data for each

site were compared to the Canadian Council of Ministers of the Environment (CCME)

and BC water quality guidelines (BCWQG).

Total suspended solids (TSS) and turbidity values tend to be the highest at sample

sites in the North Thompson River and Bone Creek, the latter the result of small scale

debris flows upstream caused by larger rain events. Metals that exceeded the

applicable aquatic life protection guidelines included total and dissolved aluminum,

total and dissolved cadmium, total chromium, total cobalt, total and dissolved copper,

total and dissolved iron, and total lead, total manganese, total selenium, total silver,

total thallium and total zinc. Of these, dissolved cadmium, total cobalt, dissolved

copper, total and dissolved iron, total manganese, total selenium, total silver, total

thallium and total zinc did not exceed the applicable guidelines in most years. In most

years, concentrations of total chromium and total copper tend to be naturally elevated





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in the North Thompson River and Bone Creek compared to other sites. As the Project

is at the exploration stage, these values reflect natural background values. While

elevated metal concentrations were noted in the sediment samples from selected

streams, chromium was the only metal to exceed CCME Sediment Guidelines in one

sample collected from Bone Creek.

Samples from surface rocks and drill core were collected and tested as part of the

Upper Fir acid rock drainage/metal leaching (ARD/ML) characterization program.

Laboratory tests included static acid-base accounting, total inorganic carbon and a

standard multi-element ICP suite on material solids.

A subset of the samples were submitted for additional testing including solids trace

element/rare earth element (REE) chemistry, short-term leach extraction tests, acid

buffering characteristic curves (ABCC) and mineralogical analysis including

petrography and Rietveld XRD. All carbonatite samples tested were classified as non-

potentially acid generating (non-PAG). Paste pHs for nearly all host rock samples

were near-neutral to alkaline indicating currently available buffering capacity in the

samples at the time of testing.

Most country rock in the Upper Fir deposit was characterized by generally low to

moderate sulphide content (<1% sulphur) and low to moderate neutralization potential

predominantly provided by slower reacting silicate minerals. However, a minor

proportion of country rock (~10% of samples) was associated with elevated sulphide

content (>1% sulphur). These rock units showed a range in acid generation potential

classifications, and in particular, a significant proportion of gneiss (~52% of gneiss

samples) was considered potentially acid generating (PAG). The majority of

amphibolite (~85%) and pegmatite (~65%) samples were classified as non-acid

generating (non-PAG), with a minor proportion classifying as PAG, typically associated

with higher sulphide samples. Fenite material was considered to be non-PAG.

Based on this initial characterization program, though some proportion of waste rock

appears to be PAG, it would appear that the Blue River Mineral Resource has an

overall low potential for acid rock drainage/metal leaching (ARD/ML) generation,

especially if waste segregation strategies can be incorporated into proposed mining

methods. Kinetic test work on two composite samples is ongoing and results will be

incorporated into planning for additional sampling as well as modeling of PAG. No

work has yet been completed on the ARD/ML potential of tailings.

Hydrogeologic investigations show that groundwater elevation in bedrock roughly

mimics topography in the Project area, so flow in the vicinity of the Upper Fir Deposit is

generally from east to west. Groundwater depth in boreholes was observed to range





Project No.: 162230


from 15 m to 130 m in the area of the deposit, and is near the surface at lower

elevations. Most groundwater flow through bedrock occurs through fractures, and the

bulk hydraulic conductivity of bedrock was estimated to range from 10-8 m/s to

10-6 m/s. Due to fracture-control, flow through bedrock is likely complex, and may not

be well connected to the surface.

Of the six groundwater samples collected in 2009, the British Columbia Ministry of the

Environment Water Quality Criteria for Freshwater Aquatic Life (BCMOE) and the

Canadian Water Quality Guidelines for the Protection of Aquatic Life (CCREM) criteria

for aquatic life were exceeded in five samples for fluoride, one sample for aluminum,

three samples for chromium, one sample for copper, and three samples for zinc. No

other parameters exceeded these criteria. However, as the boreholes used in this and

past hydrogeology studies of the Project site are deep exploration holes, and were not

developed or purged prior to sampling, sample chemistry may not be accurately

characterize existing groundwater conditions.

Current air quality at the site is considered excellent with limited influence from road

traffic and forestry activities. No site specific data has been collected to date.

An initial review of environmental conditions and planned project features indicates

that proactive design and mitigation can be successful in addressing environmental

impacts associated with developing, constructing, operating and closing the proposed

Blue River Project. As with other projects in the many BC mines located in

mountainous terrains, water management will be a key issue.

20.4 Closure Considerations

Commerce has engaged in progressive reclamation activities during exploration since

geological work and drilling began to focus on the area of, and around, the Upper Fir

resource.

Conceptual closure planning for the Blue River Project involves staged reclamation

and closure over the life of the exploration, development, construction and operation of

the mine. This will include appropriate contouring and revegetation of any waste

dumps, the Upper and Lower Portals, the drystack tailings storage facility, closure of

exploration roads, trails and platforms, closure of mine roads and removal of all mine

facilities, as well as post-closure management and monitoring plans for a defined

period of time.

It is expected that the initial design of the drystack tailings storage facility, water

storage, diversion and water management structures, waste rock dumps, and other





Project No.: 162230


mine and plant facilities will be integrated into closure designs for each component as

well as for the mine as a whole.

20.5 Current Environmental Liabilities

Current environmental liabilities are believed to be restricted to exploration drilling

programs. Existing disturbances due to exploration include drill pads, trails and

accessways, which are remediated in an ongoing program of progressive closure once

Commerce establishes exploration has been completed in a particular area.

Under the existing exploration permit, a reclamation bond is in place which will cover

the cost of any outstanding reclamation from these activities.

20.6 Closure Plan

For the purposes of the PEA, a closure estimate of $10 million was incorporated in the

financial analysis. The figure was obtained by benchmarking to similar-size mines with

the same level of complexity.

20.7 Permitting

Following environmental assessment approval, permits needed for construction and

mine operations can be issued. In BC there is an option to apply for concurrent

permitting. This allows a review for permit applications to be processed at the same

time as the environmental assessment is being conducted, resulting in the permits

required for construction being issued shortly after a positive environmental review

decision.

Table 20-1 and Table 20-2 provide a listing of possible federal and provincial permits

that will be required for construction, mine operations, closure and post-closure.

This listing cannot be considered comprehensive due to the complexity of government

regulatory processes, which evolve over time, and the large number of minor permits,

licences, approvals, consents, and authorizations, and potential amendments that will

be required throughout the life of the mine. The permit requirements will be reviewed

and updated as the Project advances.

20.8 Considerations of Social and Community Impacts

Socioeconomic and cultural heritage studies have not yet been initiated for the Blue

River Project. Basic community profiling has been completed of the individual





Project No.: 162230


communities nearby as well as relevant regional government and planning

organizations. This work shows that as with other areas of rural BC, there is a large

dependence on primary resource industries, and overall the population of the area is in

decline.

Table 20-1: Provincial Permits, Approvals, Licences and Authorizations Provincial Permits Description ACT

Notice of Work Approval for exploration and site programs to be conducted to gather geological and other site information

Mines Act

Mines Act permit Approval to construct, operate and reclaim mine and its infrastructure

Mines Act

Mining Lease Land occupancy for mine (sub-surface rights)

Mineral Tenure Act

Surface Lease Surface land occupancy for mine and site infrastructure

Land Act

Licence of Occupation Land occupancy for other features (e.g. borrow pits)

Land Act

Statutory Right of Way Land occupancy for linear features Land Act Waste Discharge Permit – Water Approval to discharge mine effluent and

sewage into the environment Environmental Management Act

Waste Discharge Permit – Air Approval to discharge air emissions into the environment

Environmental Management Act

Occupant Licence to Cut Approval to remove timber (mine, infrastructure, borrow areas)

Forest Act

Road Use permits Approval to use existing forestry roads Forest and Range Practices Act

Special Use permit Approval to construct new roads Forest Practices Code of BC

Water Licence Approval to construct, maintain and decommission water works

Water Act

Section 9 Approval Approval for changes in and about a stream

Water Act

Section 8 Approval Approval for short term use of surface water

Water Act

Authorization for Public Highway Use Approval to use public highways Transportation Act Exemption Permit Approval to haul concentrate (if required) Transportation Act Construction Permit To construct a potable water system Drinking Water

Protection Act

Table 20-2: Federal Permits, Approvals, Licences and Authorizations Federal Permits Description ACT

Navigable Waters Approval Approval to build bridges across streams Navigable Waters Protection Act

Section 35(2) Authorization Allows harmful alteration or disruption of fish habitat (HADD) (e.g. bridge upgrade)

Fisheries Act

Explosives Magazine Licence Approval to store explosives Explosives Act Radio Licences Approval to operate radios Radio Communications

Act





Project No.: 162230


The Blue River Project is located in the North Thompson River valley within the

Thompson-Nicola Regional District (TNRD). TNRD functions as a partnership of 11

member municipalities (Ashcroft, Barriere, Cache Creek, Chase, Clearwater, Clinton,

Kamloops, Logan Lake, Lytton, Merritt and Sun Peaks) as well as 10 electoral areas

whose voices at the Board table are representative of many small unincorporated

communities, member municipalities and electoral areas.

The area has a population of over 122,286 (2006 census) and a total area of

45,279 km2. The Regional District is active in providing over 115 services including

planning and building inspection, emergency preparedness and 911 services,

recreation, utilities, TV rebroadcasting, river buoys, transit, tourism, economic

development as well as environmental health services which include waste reduction,

mosquito and weed control.

The closest town to the Project is the small community of Blue River located about

20 km south of the project. Blue River is an unincorporated village, located at the

confluence of the Blue and North Thompson Rivers along the Yellowhead Highway about

halfway between Kamloops and Jasper, Alberta. It currently has a declining population

of about 260 residents, with a local economy supported by logging, tourism, and

transportation industries. Accommodation is available for exploration crews by way of

hotels and rental housing. Commerce maintains an active presence in the town with a

field office open during the exploration season.

The Project is about 90 km south of the village of Valemount, BC, a rural community of

about 1,150 situated between the Rocky, Monashee, and Cariboo Mountains. It is the

nearest community to the west of Jasper National Park, and is also the nearest

community to Mount Robson Provincial Park.

Outdoor recreation is popular in summer and winter; hiking, skiing, snowmobiling, and

horseback riding are common activities. Economic activities include logging, railway,

transport and tourism. Valemount is considered a fully-serviced village, boasting high-

speed wireless internet, train, bus and highway service. The town serves as a supply

centre for another 700 people who live in the Regional District of Fraser-Fort George,

from Albreda to Small River. Today Valemount’s economy is based on logging and a

rapidly-growing tourism industry.

20.8.1 First Nations

The Blue River Project lies on lands which comprise part of the traditional territory of

the Simpcw First Nation. Simpcw First Nation is a member of Secwepemc (Shuswap)

Nation Tribal Council (SNTC), a political organization, which works on matters of

common concern, including the development of self-government and the settlement of





Project No.: 162230


the aboriginal land title questions. The SNTC is involved with natural resource

management within the Secwepemc Nations territory and the creation of economic

development opportunities for Secwepemc communities. Commerce is very aware of

its responsibility to appropriately engage local and regional First Nations early in the

planning and development stages of the project.

On behalf of Commerce, members of the Simpcw First Nation completed

Archaeological Overview Assessments (AOA) over all areas of proposed disturbance

related to Commerce exploration activities, as well as over key areas of potential

project infrastructure. No concerns were noted by the archaeology field technicians

and exploration activities were approved to proceed by the Simpcw archaeologist with

no further recommendations for work necessary in the areas surveyed.

Traditional Knowledge/Traditional Use (TK/TU) studies, as well as a detailed

archaeological impact assessment will need to be undertaken and will also involve

Simpcw First Nation participation. Such studies may identify areas and seasons

where Simpcw have engaged in traditional activities such as hunting, fishing, gathering

and spiritual ceremonies, and the outcomes will be used to inform the overall design

and operation of the Project.

First Nations engagement, with respect to exploration activities, began in May 2007,

and will be continuing for the duration of the project. Engagement activities have

included presentations and discussions with Chief and Council, one-on-one meetings

and a site visit by elders.

On 25 October 2010, Simpcw First Nation and Commerce signed a confidential

Exploration Agreement with respect to exploration activities on the Blue River project,

which formalized a process for ongoing discussion regarding all exploration activities,

recognizes the traditional cultural, heritage, and environmental interest of the Simpcw,

and ensures that benefits from the project are realized by Simpcw First Nation.

Commerce has also committed to involve the Simpcw in environmental plans to gain

from their knowledge of the region, as well as to keep them informed of project goals.

20.8.2 Local Communities

The Blue River Project is only at the exploration and early evaluation stage; however,

to introduce Commerce and its Project to its communities, Commerce has hosted one

community meeting in each of Blue River and Valemount, and has made presentations

to the Valemount Council. The Valemount Mayor and Council have toured the

property and continue to receive regular updates on the project. Periodic Community

Newsletters provide updates on the Blue River Project; these are distributed in the





Project No.: 162230


town of Blue River and Valemount, and available on Commerce’s website. In the

summer of 2010, Commerce hosted a community barbeque in Blue River to thank its

neighbours and the local people for their assistance and support over the past

exploration seasons. As the project moves forward, open houses/information sessions

and meetings will take place in other local communities such as Barriere, Clearwater

and Chu Chua.

Public engagement to date has included meetings with local councils (e.g., Valemount,

Barriere) and informal discussions with local land-owners and Vavenby.



• The Blue River Project will require approval under the Federal and Provincial


and authorizations for construction and mine operation

• Overall the environmental review of a project is a process that will take up to

18 months to complete. The process would include the development of several

important documents by Commerce, including the Project Description, Assessment

Information Requirements and an Environmental Impact Assessment application,

followed by the review of these documents by the public, interested stakeholders,

First Nations and regulators

• The environmental review and assessment process results in a decision with

respect to whether or not the Project should be issued an Environmental

Assessment Certificate by the provincial government, as well as receive federal

Ministerial approval based on the recommendations put forth to the Minister of

Environment in a Comprehensive Study Report prepared by the Major Projects

Management Office. Both are required for a project to proceed to permitting and

development

• Environmental monitoring, baseline studies and site investigations have been

ongoing at the Blue River Project site since the summer season of 2006 with the

selection of local and regional studies areas for each biophysical discipline. Field

studies completed by specialist consultants independent of Commerce include: site

hydrology (2006–present); snow course depths (2007, 2008); fisheries and

aquatics (2006–2008); soils, flora and fauna assessments (2006, 2007), including

studies of rare, threatened and endangered plants (2007), breeding birds (2007)

and terrestrial ecosystem mapping (2006, 2007); wildlife studies and habitat

suitability mapping (2006–2008); geochemistry, mineralized material and waste





Project No.: 162230


rock characterization with baseline ABA and metals analyses (2007, 2008), surface

water and sediment quality (2006–present), groundwater (2007–2009) and terrain

stability assessment for roads (2007–present)

• Kinetic test work for ARD/ML was initiated in June 2010 and is ongoing; results of

this work will give an indication of the type of management strategies required for

handling PAG waste rock. Additional environmental baseline programs are

expected to continue, as required through 2011

• A preliminary list of the Federal and Provincial permits required for operation of a

mine has been developed. This listing cannot be considered comprehensive due

to the complexity of government regulatory processes, which evolve over time, and

the large number of minor permits, licences, approvals, consents, and

authorizations, and potential amendments that will be required throughout the life

of the mine. The permit requirements will be reviewed and updated as the Project

advances

• Socioeconomic and cultural heritage studies have not yet been initiated for the

Blue River Project. Basic community profiling has been completed of the individual

communities nearby, as well as relevant regional government and planning

organizations. This work shows that as with other areas of rural BC, there is a

large dependence on primary resource industries, and overall the population of the

area is in decline

• The Blue River Project lies on lands which comprise part of the traditional territory

of the Simpcw First Nation

• First Nations engagement, with respect to exploration activities, began in May

2007, and will be continuing for the duration of the project. Engagement activities

have included presentations and discussions with Chief and Council, one-on-one

meetings and a site visit by elders. On behalf of Commerce, members of the

Simpcw First Nation completed Archaeological Overview Assessments (AOA) over

all areas of proposed disturbance related to Commerce exploration activities, as

well as over key areas of potential project infrastructure. Traditional

Knowledge/Traditional Use (TK/TU) studies, as well as a detailed archaeological

impact assessment will need to be undertaken and will also involve Simpcw First

Nation participation

• On 25 October 2010, Simpcw First Nation and Commerce signed an Exploration

Agreement with respect to exploration activities on the Blue River project, which,

amongst other aspects, formalized a process for ongoing discussion regarding all

exploration activities, recognizes the traditional cultural, heritage, and

environmental interest of the Simpcw, and ensures that benefits from the project

are realized by Simpcw First Nation. Commerce has also committed to involve the





Project No.: 162230


Simpcw in environmental plans to gain from their knowledge of the region, as well

as to keep them informed of project goals

• Public engagement to date has included meetings with local councils (e.g.,

Valemount, Barriere) and informal discussions with local land-owners.





Project No.: 162230


21.0 CAPITAL AND OPERATING COSTS

21.1 Capital Cost Estimates

21.1.1 Basis of Estimate

Capital costs used in this PEA were derived from a variety of sources including but not

limited to comparative analysis of other operations, derivation from first principles,

equipment quotes and factoring from other costs contained within this study.

The accuracy of the estimates contained within this study vary due to the different

methods of derivation used to estimate the costs; however, in general the capital costs

are expected to be within a +40%/-10% range as per the AACE Class 5 (scoping level)

definition.

The estimate scope is limited to the battery limits of the plant and mine sites with no

allowance for off-site facilities.

All costs are expressed in first quarter (Q1) 2011 Canadian (CAD) dollars. No

allowance has been included for escalation, interest or financing fees, taxes or duties.

The capital costs are divided into five areas:

• Infrastructure costs

• Process costs

• Mining costs

• Material handling costs

• Contingency costs

• Indirect costs

Escalation is excluded from the estimate.

21.1.2 Infrastructure

Blue River initial direct civil infrastructure capital costs amount to $30M. This area

covers the infrastructure and facilities required to support the mine/mill operations

including site preparation, civil work, services, roads, explosive facilities and electrical

substation.





Project No.: 162230


The planned surface conveyor system is included as a material handling item, and

does not appear in the civil infrastructure total.

Power supplies in the region have been assumed to be sufficient for Project

requirements, and no allocation for additional power line construction has been

included.

21.1.3 Material Handling

Blue River initial direct material handling capital costs amount to $8M. This area

covers belt conveyors and transfer stations and is based on in-house AMEC data and

benchmarking against comparable projects. The cost estimate for the primary crusher,

bin and structure is included in the process plant capital estimates.

21.1.4 Process Plant

Blue River initial direct process plant capital costs amount to $116M. This area covers

all the process equipment and structures from mills to tailing filters, as well as pre-

treatment and refinery facilities.

21.1.5 Mining

Blue River initial direct mining capital costs amount to $89M. Mining direct capital costs

include pre-production mining, capital development costs, mine mobile equipment, and

mine infrastructure.

Development to be completed prior to the commencement of production at full rate

was classified as pre-production development. This development was assumed to be

undertaken by Owner’s mining crews with unit costs rates as shown in the operating

cost section.

Development associated to semi-permanent excavations, when used for more than

two years, was treated as capital development. Based on preliminary designs, an

estimate of 20% of all development was treated as capital development.

The equipment hours required for each unit of activity and daily service equipment

requirements were estimated. The required equipment operating hours for each

equipment type were aggregated. Assuming typical yearly operating hours for each

type of equipment, AMEC has estimated minimum equipment fleets to forecast capital

expenditure.





Project No.: 162230


AMEC used a database of budget costs for mine equipment. Where necessary, the

budget costs were factored to reflect Q1 2011 costs.

AMEC used a database of costs estimates for mining infrastructure and fixed service

equipment. Where necessary, the budget costs were factored to reflect Q1 2011

costs.

Initial direct mining capital requirements include major equipment for underground

operations (drilling, loading and hauling), support equipment for underground

operations, equipment for surface road maintenance and hauling and handling of

tailings from the process plant to the drystack (co-disposal) area.

Capital requirements were allocated to Year -1, but during more detailed studies,

consideration should be given to allocating the capital requirements over more than

one year, as it is likely that payments for equipment will be required prior to the

equipment being delivered to site.

In a similar manner, the development metreage allocated to the pre-production year

should be re-evaluated during more detailed studies, and a formal pre-production

development schedule with achievable monthly development metreage targets should

be developed.

21.1.6 Contingency Costs

Blue River contingency costs amount to $44 M. Contingency accounts for unforeseen

costs within the project scope. Contingency costs were calculated using a factor of

25% of civil infrastructure, material handling, process plant, and mine infrastructure

direct capital costs. A contingency factor of 5% was applied to the mine mobile

equipment direct capital costs. No contingency was calculated for pre-production

mining and capital development costs. The contingency factors are considered

appropriate for the level of engineering work performed in the preparation of this

Report. Input variables used in calculating the contingency are a result of information

gathered from previous projects and industry standards.

21.1.7 Indirect Costs

Blue River indirect costs amount to $92 M. Indirect costs covers temporary

construction facilities and services, construction equipment, freight, vendor’s

representatives, start-up and commissioning, engineering, procurement and contract

management (EPCM), working capital, warehousing spares, and first fill. Indirect costs

were calculated using a factor of 30% of civil infrastructure, material handling, process

plant, and mine infrastructure direct capital and contingency costs. An indirect costs





Project No.: 162230


factor of 5% was applied to the mine mobile equipment direct capital and contingency

costs. No indirect cost was calculated for pre-production mining and capital

development costs.

21.1.8 Sustaining Capital

The sustaining capital costs of Blue River total to $116 M. The primary sustaining

capital components of the proposed mine are:

• Underground mine development ($34 M)

• Fleet replacement ($73 M)

For underground development, the cost of development of the entire mine life was

estimated, and then factors were applied to distribute this cost over the life-of-mine,

with costs decreasing as time increased. A unit cost in $/m was then applied against

the annual metres.

The second portion of the initial truck purchases in the first production year was

categorized as sustaining capital.

Equipment fleet replacement costs were based on actual requirements as the useful

life of each unit was reached. Major mobile equipment replacements were considered

in Years 5–6 of operation, smaller mobile equipment was considered to be replaced in

four-year intervals.

21.1.9 Mine Closure

A total of $10 M was estimated for mine closure and was benchmarked to similar-size

mines with the same level of complexity.

21.1.10 Capital Cost Estimate Summary

Contractor-mining is not envisaged for the pre-production development period. Owner

mining is envisaged as costs associated with this option are reduced.

The estimate covers the direct field costs of executing the project, plus the Owner’s

indirect costs associated with design, construction, and commissioning. The

preproduction costs are capitalized that include all the expenditures before Year 1 of

production.

The estimate is summarized in Table 21-1 and Figure 21-1.





Project No.: 162230


Table 21-1: Summary of Estimated Capital Costs

Calendar year Total

(x $1,000) 2011

(x $1,000) 2012

(x $1,000)

Project year 1 2

Production year -2 -1

Capital expenditure

Initial Capital Infrastructure 29,500 10,300 19,200

Process Initial Capital 116,200 40,700 75,600

Mining Initial Capital 89,400 89,400

Material Handling 8,000 8,000

Contingency 43,600 12,800 30,900

Indirect/Owner Costs 92,300 29,600 62,600

Total 379,000 93,400 285,600 Note: Summation discrepancy due to rounding.

Figure 21-1: Distribution of Capital Costs

21.2 Operating Cost Estimates

The operating cost estimate excludes:

• Contingency

• Allowance for escalation

• Sales tax

• Product shipping





Project No.: 162230


• Import duties. This would only be applicable if any equipment is required to be

sourced outside of Canada.

21.2.1 Basis of Estimate

The operating costs for the Blue River project are based on an Owner-operated mining

fleet and process facility and have been prepared in first quarter 2011 Canadian

dollars. Owner-operation was selected as this method minimizes operating costs.

AMEC recommends that this selection be reviewed during more detailed studies. It

may be preferable to have contract labour for the pre-production development at a

minimum.

21.2.2 Mine Operating Costs

Over the life of mine, mining production costs total $529 M, and the unit cost was

$21.16/t mined.

The mine operating cost estimate incorporates costs for operating and maintenance

staff, underground mine development costs, waste management, and haulage costs.

All mining costs are based on production occurring between Years 1 and 10. Pre-

production/development costs have been capitalized and include the costs for the

mining infrastructure.

Operating costs were developed from models estimating the unit costs of the main

mining activities and overhead requirements to support the activities. The models

were previously developed for analogous mining operations and were adapted to suit

the Blue River Project.

Costs associated with activities related to mine services were factored from similar

operations and applied as fixed costs per year. Mine services costs included

consideration of items such as ventilation, heating, dewatering, compressed air,

maintenance costs, and support services.

AMEC estimated the cost to maintain the surface haulage roads and the operation of

the co-disposal site based on typical productivity of surface equipment such as dozers,

graders, excavators, compactors and water trucks. Unit hourly costs were related to

the volume of materials to be hauled and disposed.

Snow removal costs were included with surface road maintenance costs.

Operating costs for the tailings disposal facility were included in the Project sustaining

capital.





Project No.: 162230


Costs associated with Mine Administration and Technical Services for the mine were

calculated and applied as fixed costs per year.

21.2.3 Process Operating Costs

Over the life of mine, process operating costs total $339 M and the unit cost was

$13.54/t milled.

Processing cost include the costs of operating and maintaining the processing facility

from the crusher and mills to refinery and acid reagent systems. The processing

operating costs account for purchasing consumables, equipment maintenance, spare

parts, labour, and power consumption. All the workforce organization, salaries, hourly

workers, power consumption and rate are based on AMEC in-house data as well as

current market research.

21.2.4 Infrastructure Operating Costs

Based on benchmarking with other projects, a figure of $0.74/t of mineralized material

milled was used to estimate the conveying operating costs.

21.2.5 General and Administrative Operating Costs

Over the life of mine, general and administrative costs total $75 M.

The general and administrative (G&A) costs are the labour and overhead expenses for

cost centres that are not directly linked to the mining and process discipline. G&A for

each cost centre was estimated either from the first principles or input from AMEC

based on other operations.

Maintenance costs were assigned to G&A to cover maintenance cost not specific to

either mine or plant. Other incorporated costs include site service, safety and security,

warehouse, legal expenses, community and First Nations relations, insurance,

environment, and freight charges.

21.2.6 Operating Cost Summary


Operating costs include the three key areas of mining, process, and overall general

and administrative costs (G&A). The estimates are based on the staffing level,

consumables, and expenditures detailed as part of the underground mine plan and

process design.

Average operating costs are illustrated in Table 21-2. A percentage breakdown of life-

of-mine (LOM) operating costs is shown in Figure 21-2. Table 21-3 provides the





Project No.: 162230


operating costs by year, and Table 21-4 shows the unit operating cost per milled

tonne.

Table 21-2: Average Life-of-Mine Operating Cost Summary

Summary of Average Production Costs LOM Total (x $1,000)


Mining 528,900 21.16

Process 338,500 13.54

Material Handling 18,500 0.74

G&A 75,000 3.00

Sub-total 960,900 38.44

Figure 21-2: LOM Operating Cost Summary





Project No.: 162230


Table 21-3: Operating Costs by Year Calendar year $ 000 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

Project year 3 4 5 6 7 8 9 10 11 12

Production year 1 2 3 4 5 6 7 8 9 10

Mining Development 171,329 24,652 28,640 24,504 20,965 17,937 15,347 13,130 11,234 9,611 5,309

Production 84,237 8,653 8,364 8,663 8,919 9,139 9,326 9,486 9,624 9,741 2,322

Haulage 161,176 19,661 17,766 17,529 17,326 17,152 17,004 16,877 16,768 16,675 4,420

Service & Infrastructure 55,002 6,284 5,954 5,933 5,915 5,900 5,887 5,876 5,867 5,858 1,527

Mining G&A 57,193 6,177 6,177 6,177 6,177 6,177 6,177 6,177 6,177 6,177 1,600

Mining Production Costs 528,937 65,426 66,901 62,806 59,303 56,305 53,741 51,546 49,669 48,063 16,665

Unit Cost $/tonne 21.16 24.23 24.78 23.26 21.96 20.85 19.90 19.09 18.40 17.80 21.68

Table 21-4: Operating Costs Per Milled Tonne Calendar year 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

Project year 3 4 5 6 7 8 9 10 11 12

Production year 1 2 3 4 5 6 7 8 9 10

Mining (24) (25) (23) (22) (21) (20) (19) (18) (18) (6)

Process (14) (14) (14) (14) (14) (14) (14) (14) (14) (4)

Material Handling (1) (1) (1) (1) (1) (1) (1) (1) (1) (0)

G&A (3) (3) (3) (3) (3) (3) (3) (3) (3) (1)

Total Operating Cost/tonne (42) (42) (41) (39) (38) (37) (36) (36) (35) (10)





Project No.: 162230



In the opinion of the QPs:

• The total estimated capital cost to design and build the Blue River Project at an

assumed 7,500 t/d capacity is $379 M

• A total of $10 M was included in the capital cost estimate for mine closure and was

benchmarked to similar-size mines with the same level of complexity

• Operating costs over the life-of-mine are estimated at $38.44/t milled.





Project No.: 162230


22.0 ECONOMIC ANALYSIS The following section is partly based on 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 assessment based on these Mineral Resources will be

realized.

Approximately 15% of the Mineral Resources that were modified to produce the ROM

estimate and support the financial model have been classified as Inferred Mineral

Resources.

The results of the economic analyses discussed in this section represent forward-

looking information as defined under Canadian securities law. The results depend on

inputs that are subject to a number of known and unknown risks, uncertainties and

other factors that may cause actual results to differ materially from those presented

here. Information that is forward-looking includes:

• Mineral Resource estimates

• Assumed commodity prices and exchange rates

• The proposed mine production plan

• Projected recovery rates

• Infrastructure construction costs and schedules

• Assumptions that an EA will be approved by Provincial and Federal authorities.

22.1 Valuation Methodology

The project has been evaluated using a discounted cash flow (DCF) analysis. Cash

inflows consist of annual revenue projections for the mine and two years of

preproduction. Cash outflows such as capital, operating costs, and taxes are

subtracted from the inflows to arrive at the annual cash flow projections.

To reflect the time value of money, annual net cash flow (NCF) projections are

discounted back to the project valuation date using several discount rates. The

discount rate appropriate to a specific project depends on many factors, including the

type of commodity and the level of project risks, such as market risk, technical risk,

and political risk. The discounted present values of the cash flows are summed to

arrive at the project’s net present value (NPV).





Project No.: 162230


In addition to NPV, internal rate of return (IRR) and payback period are also

calculated. The IRR is defined as the discount rate that results in an NPV equal to

zero. Cash flows are taken to occur at the end of each period. Capital cost estimates

have been prepared for initial development and construction of the project and for

ongoing operations (sustaining capital).

The resulting net annual cash flows are discounted back to the date of valuation end-

of-year 2010 dollars and totalled to determine NPVs at the selected discount rates.

The IRR is calculated as the discount rate that yields a zero NPV. The payback period

is calculated as the time needed to recover the initial capital spent.

22.2 Financial Model Parameters

22.2.1 Mineral Resources and Mine Life

The model includes 36,349 kt of Indicated Mineral Resources as well as 6,385 kt of

Inferred Mineral Resources. For this study AMEC utilized average grades of

mineralized materials throughout the mine life at 195 ppm for Ta and 1,700 ppm for

Nb. The diluted grades as the result of the proposed mining method were assumed at

185 ppm Ta and 1,591 ppm Nb. After applying mine recovery and dilution factors, the

financial model assumes that the mine life is 10 years, assuming the plant will process

25 Mt at a 7.5 kt/d plant throughput rate (2.7 Mt/a).

22.2.2 Metallurgical Process

Recovery assumptions from the process plant include 65% recovery for Ta and 69%

recovery for Nb in the flotation stage. The refining process has an estimated 97%

recovery for both Ta and Nb.

22.2.3 Commodity Prices and Foreign Exchange

Publicly-available tantalum and niobium pricing information is very limited as the

markets tend to be based on long-term relationships between few buyers and sellers.

Slightly more information is available for niobium than for tantalum.

Given that there is currently a shortage of tantalum and the fact that historic prices do

not appear to sustain future production from any currently known projects, AMEC

considered the following question:

"What price is needed for a typical greenfield property to be brought into production

and give its owners a reasonable return on their investment?"





Project No.: 162230


Approximate calculations indicated that a tantalum price of US$317/kg of contained

metal in the oxide product was needed. This price is close to the prices for tantalum

reported on subscription news services during the last quarter of 2010.

The niobium price was set at US$46/kg of contained metal in the oxide product over

the life-of-mine.

An exchange rate of US$0.95 to C$1 is used for all years of the financial model.

22.2.4 Taxes

No taxes are incorporated in the model. AMEC is not an expert in taxation matters,

and appropriate taxation considerations should be reviewed with relevant auditors.

AMEC did review publicly-available taxation documents for taxes that could be levied.

Taxation considerations comprise Provincial and Federal corporate income taxes and

BC Mineral taxes. The following discussion outlines the main Federal and Provincial

taxation and considerations for mining ventures in BC:

• Federal taxes: Includes income tax, customs duties, fuel taxes, payroll taxes and

transaction taxes. The general rate of Federal income tax on active business

income earned by a corporation for 2011 is 16.5% and is legislated to decrease to

15% starting in 2012.

• Provincial income tax: The general rate of BC Provincial income tax on active

business income earned by a corporation in the Province is 10%.

• Provincial mineral taxes: The BC Mineral Tax provides for the Crown's financial

share of mineral production in two ways. The primary way is to receive 13% of a

producer’s profit that is in excess of a normal return on investment over the life of a

mine. This is referred to as Net Revenue Tax. To minimize any disincentive to

investment, the Province does not receive this share until the producer’s

investment and a reasonable return on it have been recovered. The second way is

to receive 2% of operating cash flow from production in each year. This is referred

to as Net Current Proceeds Tax. It is intended to provide compensation for

depletion of the resource when production yields less than a reasonable profit for

the producer. So that only one or the other share is paid, Net Current Proceeds

Tax is fully creditable against Net Revenue Tax.

22.2.5 Depreciation/Salvage Value

No depreciation is incorporated in the model.





Project No.: 162230


22.2.6 Financing

The project is assumed to be 100% Owner-financed.

22.2.7 Capital Costs

The total estimated capital cost to design and build the Blue River Project at 7,500 t/d

capacity is $379 million.

22.2.8 Operating Costs


22.2.9 Working Capital

No working capital is incorporated in the model. Working capital allowances required to

operate this Project are not expected to have a significant impact on the cash flow of

the mine.

22.2.10 Inflation

No inflation adjustments are incorporated in the model. Capital and operating costs are

based on first quarter 2011 Canadian dollars.

22.2.11 Royalty

No royalty is incorporated in the model.

22.3 Financial Results

A discount rate of 8% was applied as the base case. The Project base case returns

an NPV of $18.5 M before tax with a 9.1% IRR (internal rate of return), and a 6.3 year

payback period. Discount rates from 5% to 10% return NPVs $80 M to ($14) M.

Annualized cash flows are summarized in Table 22-1. The financial results are

summarized in Table 22-2.





Project No.: 162230


Table 22-1: Project Cashflow Model

Calendar year Units Life of Mine 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

Project year

1 2 3 4 5 6 7 8 9 10 11 12

Production year

-2 -1 1 2 3 4 5 6 7 8 9 10

Metal prices

Ta USD/kg

317 317 317 317 317 317 317 317 317 317 317 317

Nb USD/kg

46 46 46 46 46 46 46 46 46 46 46 46

Tantalum Revenue

Ta

Final Product Mt 2,403

260 260 260 260 260 260 260 260 260 67

Total in US$ USD000 761,702

82,264 82,264 82,264 82,264 82,264 82,264 82,264 82,264 82,264 21,328

Total in CAD$ CAD$000 801,792

86,593 86,593 86,593 86,593 86,593 86,593 86,593 86,593 86,593 22,450

Niobium Revenue

Nb

Final Product Mt 18,610

2,010 2,010 2,010 2,010 2,010 2,010 2,010 2,010 2,010 521

Total in US$ USD$000 856,039

92,452 92,452 92,452 92,452 92,452 92,452 92,452 92,452 92,452 23,969

Total in CAD$ CAD$000 901,094

97,318 97,318 97,318 97,318 97,318 97,318 97,318 97,318 97,318 25,231

Revenue CAD$000 1,702,885

183,912 183,912 183,912 183,912 183,912 183,912 183,912 183,912 183,912 47,681

Production costs

Mining CAD$000 528,937

65,426 66,901 62,806 59,303 56,305 53,741 51,546 49,669 48,063 15,177

Process CAD$000 338,500

36,558 36,558 36,558 36,558 36,558 36,558 36,558 36,558 36,558 9,478

G&A CAD$000 74,998

8,100 8,100 8,100 8,100 8,100 8,100 8,100 8,100 8,100 2,098

Material Handling CAD$000 18,516

2,000 2,000 2,000 2,000 2,000 2,000 2,000 2,000 2,000 516

Total CAD$000 960,951

112,084 113,559 109,464 105,961 102,963 100,399 98,204 96,327 94,721 27,269

Earnings

Earnings before taxes, depreciation & amortization CAD$000 741,934

71,828 70,352 74,447 77,951 80,948 83,513 85,707 87,585 89,191 20,412

Deductible interest CAD$000

0 0 0 0 0 0 0 0 0 0 0

Earnings before taxes, depreciation & amortization CAD$000 741,934

71,828 70,352 74,447 77,951 80,948 83,513 85,707 87,585 89,191 20,412

Operating earnings CAD$000 741,934

71,828 70,352 74,447 77,951 80,948 83,513 85,707 87,585 89,191 20,412





Project No.: 162230


Calendar year Units Life of Mine 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

Project year

1 2 3 4 5 6 7 8 9 10 11 12

Production year

-2 -1 1 2 3 4 5 6 7 8 9 10

Initial and Sustaining Capital Expenditure

Infrastructure Initial Capital CAD$000 29,491 10,322 19,169

Process Initial Capital CAD$000 116,240 40,684 75,556

Mining Initial Capital CAD$000 89,420

89,420

Material Handling CAD$000 8,000

8,000

Contingency CAD$000 43,613 12,751 30,862

Indirects/Owner Costs CAD$000 92,268 29,627 62,641

Sustaining Capital CAD$000 116,271

46,754 4,674 4,183 4,757 24,857 20,875 3,225 3,636 1,995 1,317

Total 495,303 93,385 285,648 46,754 4,674 4,183 4,757 24,857 20,875 3,225 3,636 1,995 1,317

Closure and Salvage

Closure costs CAD$000 10,000

10,000

Total CAD$000 10,000

10,000

Net Project Cash Flow

Pre-tax CAD$000 236,631 (93,385) (285,648) 25,074 65,678 70,264 73,194 56,091 62,639 82,482 83,949 87,196 9,095





Project No.: 162230


Table 22-2: Summary Pre-Tax Financial Analysis

Summary of Cash Flow Pre-tax

Cumulative net cash flow

Undiscounted CAD$000 236,631

Net present value




Discounted at 8% (Project Base Case) CAD$000 18,487


Discounted at 10% CAD$000 (13,514)

Internal rate of return % 9.1

Payback period Years 6.3

22.4 Cash Costs

The cash cost value represents the cost incurred to produce 1 kg of primary product

after deducting the revenue from sales of secondary products. Since the price

analysis for the report was performed around Ta price variation, Ta is chosen as the

main product and Nb is treated as the secondary product for the assessment of cash

cost.

Cash costs are derived through the following formulae:

Production costs = (Mining + Process + G&A + Material Handling)

Cash cost = (Production costs - Revenue of Nb sales) ÷ tantalum production in kg

Using the Brook Hunt convention for reporting C1 cash costs2, after credit for Niobium

contribution, the tantalum cash cost is calculated to be approximately $24.91/kg

contained in oxide product (equivalent to US$23.66/kg contained in oxide product with

the PEA study exchange rate of US$0.95 to C$1).

This is shown in Table 22-3.

2 Brook Hunt, established in 1975, is a global group that specializes in in-depth market analysis across the mining and metals

industries. Brook Hunt has established a method of comparison of costs between projects, countries and commodities that is considered an industry standard. C1 cash costs are defined by Brook Hunt as: the costs of mining, milling and concentrating, on-site administration and general expenses, property and production royalties not related to revenues or profits, metal concentrate treatment charges, and freight and marketing costs less the net value of by-product credits.





Project No.: 162230


Table 22-3: Life of Mine Cash Cost Summary

Summary of Cash Costs LOM Total (x $1,000)


Cost per Kg Ta Payable ($/kg)

Cash costs

Mining 528,900 21.16 220.13

Process 338,500 13.54 140.87

G&A 75,000 3.00 31.21

Material Handling 18,500 0.74 7.71

Sub-total 960,900 38.44 399.92

Credits

Nb (901,100) (36.04) (375.01)

Sub-total (901,100) (36.04) (375.01)

Adjusted cash costs

Total 59,800 2.40 24.91 Note: The figures in this table do not include considerations of working capital or royalty payments

The cash cost for production of tantalum during the earlier years of the proposed

mining operation is $57/kg and decreases over the life of the mine. The major driver

behind the changing costs is the decrease in the mining costs over the life-of-mine. In

the last three years of operation, the revenue generated from niobium exceeds the

total operating costs (mining, processing and G&A). The mining cost for the entire

Project (i.e. mining cost of both tantalum and niobium) drops from an average of $24/t

in the first few years to $18/t in the last year of full production.

22.5 Sensitivity Analysis

The analysis shows that the Project is more sensitive to changes in operating

expenditures than capital expenditures. The project is most sensitive, in order, to

changes in exchange rate, operating expenditure, Nb price, Ta price and capital

expenditure. Since the sales currency is in United States dollars and operational costs

are in Canadian dollars, a rising United States dollar value versus the Canadian dollar

improves potential profitability. Project sensitivities are summarized in Table 22-4 for

the Project Base Case discount rate of 8%. Figure 22-1 shows a sensitivity graph for

the Project at the Base Case NPV at 8%. The graph assesses the relative impacts on

the Project to variations in exchange rates, capital expenditure, operating expenditure,

niobium price, and tantalum price.





Project No.: 162230


Table 22-4: Sensitivity Summary, Project Base Case NPV at 8%

SENSITIVITY OF NPV @ 8% Change in Factor

-30% -20% -10% 0% 10% 20% 30%

Fa

cto

r

Exchange rate 448.7 269.5 130.0 18.5 (72.8) (148.8) (213.2)

Capital expenditure 117.9 84.8 51.6 18.5 (14.6) (47.8) (80.9)

Operating expenditure 190.5 133.1 75.8 18.5 (38.8) (96.2) (153.5)

Nb price (140.9) (87.8) (34.6) 18.5 71.6 124.7 177.9

Ta price (123.3) (76.1) (28.8) 18.5 65.8 113.0 160.3

Figure 22-1: Sensitivity Summary NPV at 8%


In the opinion of the QPs:

• Based on the assumptions in this Report, the financial analysis for the Blue River

project, using a discount rate of 8%, returns an NPV of about $18.5 M and an IRR

of 9.1% before tax.

• The Project is most sensitive, in order, to changes in exchange rate, operating

costs, niobium price, and less sensitive to changes in tantalum price , and least

sensitive to changes in capital costs.





Project No.: 162230


23.0 ADJACENT PROPERTIES There are no adjacent properties that are relevant to the Report.





Project No.: 162230


24.0 OTHER RELEVANT DATA AND INFORMATION AMEC is not aware of any other relevant data or required information for inclusion to

make the report more understandable and not misleading.





Project No.: 162230


25.0 INTERPRETATION AND CONCLUSIONS Commerce has delineated a significant tantalum- and niobium-rich carbonatite deposit

near the town of Blue River in central eastern British Columbia. They hold a 100%

interest in the project.

Commerce has professionally executed an exploration program. The quantity and

quality of the lithological, geotechnical, and collar location, down-hole survey, and drill-

core sample data collected by Commerce in the exploration and delineation drill

programs meet and exceed industry standard practice.

The Blue River project has very good access and supporting infrastructure. The

deposit is amenable to conventional underground mining methods with estimated

mining recoveries of approximately 60%. It is expected that any future mining

operations will be able to be conducted year-round.

Process testwork has shown tantalum and niobium occurs in ferrocolumbite and

pyrochlore minerals and is amenable to conventional flotation and refining processes.

High-quality technical grade tantalum and niobium products proposed for production at

site are suitable for several markets.

The project has a 10 year mine life and capital costs, at $379 M, could be paid back in

6 years. The project is most sensitive to commodity price. A 20% to 30% increase in

the assumed tantalum price significantly improves the potential profitability of the

project. The tantalum price assumption used in this PEA is based on 4th quarter 2010

information. The tantalum price moved significantly higher through 2011. A higher

tantalum price could not only improve profitability but also increase the mine life.

Additional exploration potential could also provide additional mine life. A two or more

times capital payback is possible.

Commerce has been pro-active with regard to environmental and socioeconomic

issues. Environmental monitoring, baseline studies and site investigations have been

ongoing at the Blue River Project site since the summer season of 2006. Kinetic test

work for acid rock drainage and metals leaching was initiated in 2010. Additional

environmental baseline programs are expected to continue, as required through 2011.

First Nations engagement, with respect to exploration activities, began in 2007, and

will continue for the duration of the project. The Blue River Project lies on lands which

comprise part of the traditional territory of the Simpcw First Nation. First Nations

engagement, with respect to exploration activities, began in 2007. Public consultation

to date has included meetings with local councils and informal discussions with local

land-owners.





Project No.: 162230





The opportunities to improve the project NPV are:

• Optimization of the mine plan to mine higher-grade zones early in the mine life

providing that a practical mining sequence can be implemented and the overall

recovery of the Mineral Resources is not negatively affected

• Optimization of the mine layout to minimize development costs

• Advanced geotechnical studied to identify and understand ground conditions which

could allow an increase in the size of stopes and production drifts

• Optimization of the supply and pricing of reagents for the refining.

The risk factors are:

• The base case mineral resource estimate is supported by current tantalum and

niobium prices which are higher than historic average prices and may not reflect

long term prices. Market analysts are in general agreement that current political

and market conditions support the probability of sustained higher prices, but this

may not occur.

• Commerce has not initiated requests from potential buyers for expression of

interests from potential buyers of the proposed Blue River products and has not

negotiated any purchase or off-take agreements.



• Testwork to date has not considered factors such as water recycling. A water

treatment plant may be required and may result in increased capital costs.

• The financial analysis is partly based on 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 Assessment based on these Mineral

Resources will be realized.





Project No.: 162230


• The Blue River Project will require approval under the federal and provincial


and authorizations for construction and mine operation. Overall the environmental

review of a project is a process that can take up to 18 months to complete.

• Traditional Knowledge/Traditional Use (TK/TU) studies, as well as a detailed

archaeological impact assessment will need to be undertaken.

The projects warrant additional work to examine the opportunities and mitigate the

risks. On completion of this work Commerce may consider proceeding with a pre-

feasibility study.





Project No.: 162230


26.0 RECOMMENDATIONS A two-phase work program is recommended. The first phase comprises a Mineral

Resources update incorporating the results of the 2010 and 2011 drill programs. This

phase of work is estimated at $350,000 to $400,000. The second phase comprises

completion of mining and related studies that can be used to support a pre-feasibility

study. This phase is estimated at $2.6 M.

26.1 Phase 1

The Phase 1 work program will culminate in an update to the Mineral Resource

estimate. The mineral resource estimate update will include the results of the following

work that is currently in progress:

• Incorporation of results of the 2010 and 2011 drill program in terms of updating the

block model with new geological data, and incorporation available of analytical

data

• Confirmation of current interpretation of polyfolding in the deposit from relogging of

selected drill holes

• Modelling of thorium and uranium to confirm the elements occur in low

concentrations in the deposit

• Completion of a drill spacing study which will use a conceptual volume that would

be equivalent to the volume to be mined in the first three years of mine life. The

drill spacing study will examine the uncertainty of metal content in this conceptual

volume as a function of the selected economic cut-off (mineable zone impact)

using “current” and planned drill hole designs.

26.2 Phase 2

The Phase 2 work program will consist of completion of mining and related studies.

Phase 2 can be conducted concurrently with Phase 1 work.

The program should include:

• Re-assay program on selected drill core to remove limitations on classification

based on precision and accuracy concerns

• Geometallurgical, geotechnical and structural studies

• Metallurgical testing, pilot plant design, and studies





Project No.: 162230


• Proof-of-process metallurgical program to produce Ta and Nb oxide

• Additional marketing studies and collection of letters of interest to purchase the

technical grade tantalum and niobium products

• Advance the waste and tailings management design

• Environmental biophysical programs, metals uptake studies and foundation

testwork

• Project management, claims, community liaison and office based work

This program will provide additional data such that Commerce will be positioned to

determine the most appropriate approach for a pre-feasibility study.





Project No.: 162230


27.0 REFERENCES Aaquist, B., 1982a: Blue River Carbonatites, British Columbia: Final Report. B.C. Min.

Energy, Mines Petr. Res. Ass. Rept. 10 274, 30 p.

Aaquist, B., 1982b: Assessment Report Blue River Carbonatites, British Columbia:

B.C. Min. Energy Mines Petr. Res. Ass. Rept. 11 130, 15 p.

Aaquist, B., 1982c: Assessment Report on Verity First 1,2,3, Claims, Blue River

British Columbia: B.C. Min. Energy Mines Petr. Res. Ass. Rept. 10 955.

Birkett, T.C. and Simandl, G.J., 1999: Carbonatite-associated Deposits: Magmatic,

Replacement and Residual: in Selected British Columbia Mineral Deposit Profiles,

Volume 3, Industrial Minerals, G.J. Simandl, Z.D. Hora and D.V. Lefebure, Editors,

British Columbia Ministry of Energy and Mines.

Canadian Institute of Mining, Metallurgy and Petroleum (CIM), (2010). CIM Standards

for Mineral Resources and Mineral Reserves, Definitions and Guidelines: Canadian

Institute of Mining, Metallurgy and Petroleum, November 2010,

http://www.cim.org/committees/CIMDefStds_Dec11_05.pdf

Chong, A., and Postolski, T., 2011: NI 43-101 Technical Report, Blue River Ta-Nb

Project, Blue River, British Columbia. 145 p.

Chudy, T., 2008: Mineralogical Report on samples from the Upper Fir Carbonatite,

Blue River, British Columbia. PART A: Petrographic description; PART B: Mineral

Liberation Analysis, December 2008.

Chudy, T., 2010: The Niobium-Tantalum Mineralization In The Upper Fir Carbonatite:

A Summary Of Current Knowledge, 4 p.

Currie, K.L. 1976: The Alkaline Rocks of Canada: Geol. Surv. Can., Bull. 239, 228 p.

Dahrouge, J., 2001a: 2000 Geologic Mapping and Sampling on the Verity Property:

B.C. Min. Energy, Mines Petr. Res. Ass. Rept 26550, 7 p.

Dahrouge, J., 2001b: 2000 Geologic Mapping and Sampling on the Fir Property: B.C.

Min. Energy, Mines Petr. Res. Ass. Rept 26549, 7 p.

Dahrouge, J. and Reeder J., 2001: 2001 Geologic Mapping, Sampling and

Geophysical Surveys on the Mara Property: B.C. Min. Energy, Mines Petr. Res. Ass.

Rept. 26733, 14 p.





Project No.: 162230


Dahrouge, J. and Reeder J., 2002: 2001 Geologic Mapping, Sampling and

Geophysical Surveys on the Fir Property: B.C. Min. Energy, Mines Petr. Res. Ass.

Rept. 26781, 9 p.

Davis, C., 2006: 2005 Diamond Drilling and Exploration at the Blue River Property:

B.C. Min. Energy Mines Petr. Res. Ass. Rept, 10 p.

Diegel, S.G., Ghent, E.D., and Simony, P.S., 1989: Metamorphism and Structure of

the Mount Cheadle area, Monashee Mountains: in Current Research, Part E, Geol.

Surv. Can., Paper 89–1E, pp. 95–100.

Ghent, E.D., Simony, P.S., Mitchell, W., Perry, J., Robbins, D. and Wagner, J., 1977:

Structure and Metamorphism in the Southeast Canoe River area, British Columbia: in

Report of Activities, Part C, Geological Survey of Canada, Paper 77–1C, pp. 13–17.

Gervais, F., 2009: Personal Communication to John Gorham.

Gorham, J., 2007: Technical Report on the Upper Fir Ta-Nb Bearing Carbonatite 20

June 2007, 48 p. plus appendices.

Gorham, J., 2008: Report on 2007 Diamond Drilling and Exploration at the Blue River

Property 20 June 2008: 48 p. plus appendices and maps.

Gorham, J., Ulry, B. And Brown, J., 2009: 2008 Diamond Drilling and Exploration at

the Blue River Property, Kamloops Mining Division B.C Ministry of Energy, Mines and

Petroleum Resources, Assessment Report 31174, 79 p (plus appendices and maps)..

Klohn, Crippen Berger, 2009a: Blue River-Upper Fir Deposit Tailings and Waste Rock

Scoping Study: Prepared for Commerce Resources Corp.

Klohn, Crippen Berger, 2009b: Valemount Tailings Storage Options Scoping Study:

Prepared for Commerce Resources Corp.

Kraft, J., 2010: Structural geology of the Upper Fir carbonatite deposit, Blue River,

British Columbia: Confidential report for Dahrouge Geological Consulting Ltd. and

Commerce Resources Corp., 15 p.

Mariano, A.N., 1982: Petrology, Mineralogy and Geochemistry of the Blue River

Carbonatites: Confidential report, 130 p.

Mariano, A.N. 2000: Personal communication to J. Dahrouge.





Project No.: 162230


MESH Environmental Inc. 2008: Static Test Characterization Of Rock Units From The

Upper Fir Deposit, Blue River Tantalum-Niobium Project: Confidential report prepared

for Commerce Resources, August 2008.

MESH Environmental Inc. 2009: Static Test Characterization Of Rock Units From The

Upper Fir Deposit, Blue River Tantalum-Niobium Project. Phase 2 Static Test Report:

Confidential report prepared for Commerce Resources, April 2009.

McCrea, J. 2001: Summary Report on the Blue River Carbonatite Property, East-

Central British Columbia: Prepared for Commerce Resources Corp., 34 p.

McCrea, J. 2002: Fir Carbonatite Property, Resource Estimate: Prepared for


Mitchell, R.H., 2010b: Niobium Mineralization in Carbonatites: Paragenesis and

Origins: in International Workshop of Geology of Rare Metals, edited by Simandl, G.J.

and Lefebure, D.V., extended abstracts volume, November 9–10, 2010, Victoria,

Canada British Columbia Geological Survey, Open File 2010-10, pp 13–14.

Nicholas, D., 1992: SME Mining Engineering Handbook, 2nd Edition, Volume 2

Pell, J. 1987: Alkaline Ultrabasic Rocks in British Columbia: Carbonatites, Nepheline

Syenites, Kimberlites and Related Rocks: B.C. Min. Energy, Mines Petr. Res. Open

File 1987-17, 109 p.

Pell, J. 1994: Carbonatites, Nepheline Syenites, Kimberlites and Related Rocks in

British Columbia: B.C. Min. Energy, Mines, Petr. Res., Bulletin 88, 136 p.

Pell, J. and Hoy, T. 1989: Carbonatites in a Continental Margin Environment - the

Canadian Cordillera: in: Carbonatites: Genesis and Evolution (K. Bell, ed.). Unwin

Hyman, London, UK. pp. 200–220.

Raeside, R.P. and Simony, P.S. 1983: Stratigraphy and Deformational History of the

Scripp Nappe, Monashee Mountains, British Columbia: Canadian Journal of Earth

Sciences, 20, pp. 639–650.

Rukhlov, A and Gorham, J. 2007: 2006 Diamond Drilling and Exploration at the Blue

River Property: B.C., Min. Energy, Mines Petr. Res. Ass. Rept. 29024, 383 p. with

appendices.

Simonetti, A. 2008: Personal communication to John Gorham.





Project No.: 162230


Simony, P.S., Ghent, E.D., Craw, D, Mitchell, W., and Robbins, D.B. 1980: Structural

and Metamorphic Evolution of the Northeast Flank of the Shuswap Complex, Southern

Canoe River Area, British Columbia: Geological Society of America, Memoir 153, pp.

445-461.

Smith, M. and Dahrouge, J. 2002a: 2001 Diamond Drilling on the Fir Property: B.C.

Min. Energy, Mines Petr. Res. Ass. Rept. 26911, 13 p. with appendices.

Smith, M. and Dahrouge, J. 2003: 2002 Diamond Drilling and Exploration on the Blue

River Property: B.C. Min. Energy, Mines Petr. Res. Ass. Rept. 27131, 20 p. with

appendices.

Smith, T., 2008: Terrain stability study for the development of an environmental

baseline for the Fir Property near Blue River B.C: Confidential report prepared for


Stone, M., and Selway, J., 2010: Independent Technical Report, Blue River Property,

Blue River, British Columbia, Canada. 116 p.

White, G.P.E. 1985: Further Notes on Carbonatites in Central British Columbia: B.C.

Ministry of Energy, Mines and Petroleum Resources, Geological Fieldwork, 1984,

Paper 1985-1, pp. 95-100.

Woolley, A.R. and Kempe, D.R.C. 1989: Carbonatites: Nomenclature, Average

Chemical Compositions, and Element Distribution: in: Carbonatites, Genesis and

Evolution (K. Bell, ed.). Unwin Hyman, London. pp. 1–37.





Project No.: 162230 Appendix A

29 September 2011

Appendix A

List of Claims






29 September 2011

Tenure Number Claim Name Ownership

Tenure Type

Tenure Sub-Type

Map Number Issue Date Good To Date Area (ha)

374665 FIR 3 100% Mineral Claim 083D025 2000/feb/16 2019/mar/31 25.000

374670 FIR 8 100% Mineral Claim 083D035 2000/feb/16 2019/mar/31 25.000

380034 MARA 5 100% Mineral Claim 083D045 2000/aug/18 2019/mar/31 25.000

382164 FIR 11 100% Mineral Claim 083D035 2000/oct/28 2019/mar/31 500.000

506262 100% Mineral Claim 083D 2005/feb/08 2019/mar/31 98.623





506270 100% Mineral Claim 083D 2005/feb/08 2019/mar/31 1,225.766


































29 September 2011


Tenure Type

Tenure Sub-Type








530510 LIGHTNING 100% Mineral Claim 083D 2006/mar/24 2019/mar/31 494.525

530511 LIGHTNING 2 100% Mineral Claim 083D 2006/mar/24 2019/mar/31 395.741

530513 LIGHTNING 3 100% Mineral Claim 083D 2006/mar/24 2019/mar/31 217.556

537452 PYRAMID 1 100% Mineral Claim 083D 2006/jul/20 2019/mar/31 493.795



550560 MUD 10 100% Mineral Claim 083D 2007/jan/29 2019/mar/31 495.976




550568 MUD15 100% Mineral Claim 083D 2007/jan/29 2019/mar/31 178.524

550603 ARIANE1 100% Mineral Claim 083D 2007/jan/30 2019/mar/31 493.608












550621 4512124519227380 100% Mineral Claim 083D 2007/jan/30 2019/mar/31 474.593


550623 ARIANE 14 100% Mineral Claim 083D 2007/jan/30 2019/mar/31 494.343












29 September 2011


Tenure Type

Tenure Sub-Type


550636 ARIANE 20 100% Mineral Claim 830 2007/jan/30 2019/mar/31 493.708










550648 ARIANE 30 1 100% Mineral Claim 083D 2007/jan/30 2019/mar/31 494.394


550651 ARIANE 32 100% Mineral Claim 083D 2007/jan/3D 2019/mar/31 493.274































29 September 2011


Tenure Type

Tenure Sub-Type






550700 ARIANE 63 100% Mineral Claim 083D 2007/jan/3D 2019/mar/31 494.066
















550886 HELLROAR 100% Mineral Claim 083D 2007/feb/01 2019/mar/31 435.471

550887 HELLROARS 100% Mineral Claim 083D 2007/feb/01 2019/mar/31 475.246

550888 BAT OUT OF HELL 100% Mineral Claim 083D 2007/feb/01 2019/mar/31 475.396

550889 THE MONSTER IS LOOSE 100% Mineral Claim 083D 2007/feb/01 2019/mar/31 475.203

565127 PROSPER 1 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 475.099


565129 PROPSER 3 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 494.798








565139 PROSPER 11 ' 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 494.798









29 September 2011


Tenure Type

Tenure Sub-Type


565144 PROSPER 15 100% Mineral Claim 830 2007/aug/28 2019/mar/31 475.184

























565171 SHADOWI 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 495.650

565172 SHADOW 2 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 436.167

















29 September 2011


Tenure Type

Tenure Sub-Type


565184 SHADOW 13 100% Mineral Claim 083D 2007/auq/28 2019/mar/31 495.942



565187 FALKOR 1 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 456.085











565198 FALKOR 12 100% Mineral Claim 830 2007/au•/28 2019/mar/31 495.497

565199 MINI 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 39.701







565206 MINI 2 100% Mineral Claim 083D 2007/aug/28 2019/mar/31 39.694

565207 FALKOR 18 100% Mineral Claim 083D 2007/au /28 2019/mar/31 437.056




















29 September 2011


Tenure Type

Tenure Sub-Type





588427 WASTED 1 100% Mineral Claim 083D 2008/jul/18 2011/jul/31 494.294




589537 FELIX1 100% Mineral Claim 083D 2008/aug/05 2019/mar/31 496.396













798362 JOIN 100% Mineral Claim 083D 2010/jun/25 2011/jun/25 177.980

Note: Commerce has informed AMEC that the “Wasted” and “Join” claims (588427 to 588430 and 798362) have been renewed to March 31, 2021 and are in good standing. Email titled: “Re: Blue River 162230: Claims status request”; dated 10/28/2011 5:11 PM; from Dave Hodge, Commerce President and Director; to Albert Chong, AMEC.

Date post:	08-May-2015
Category:	Investor Relations
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Technical Report: Blue River Tantalum and Niobium Project (Preliminary Economic Assessment)

Investor Relations