Technical Report on the Dikulushi Underground Project ...

Mawson West Limited

Technical Report on the Dikulushi Underground Project

Democratic Republic of Congo – 12 December 2013.

Technical Report on the Dikulushi Underground Project Democratic Republic of Congo – 12 December 2013

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Doc Ref:

Print Date: 12 December 2013

Number of copies:

Optiro:

Mawson West Limited:

Perth Office

Level 4, 50 Colin Street

West Perth WA 6005

PO Box 1646

West Perth WA 6872

Australia

Tel: +61 8 9215 0000

Fax: +61 8 9215 0011

Optiro Pty Limited

ABN: 63 131 922 739

www.optiro.com

Principal Author: Andrew Law MMin. MBA

FAusIMM, FIQA, MAICD, AFAIM.

Signature:

Date: 12 December 2013

Principal Reviewer: Ian Glacken FAusIMM(CP), CEng

Contributing author:

Signature:

Date: 12 December 2013

Important Information:

This Report is provided in accordance with the proposal by Optiro Pty Ltd (“Optiro”) to Mawson West Limited and the

terms of Optiro’s Consulting Services Agreement (“the Agreement”). Optiro has consented to the use and publication of

this Report by Mawson West Limited for the purposes set out in Optiro’s proposal and in accordance with the Agreement.

Mawson West Limited may reproduce copies of this Report, in whole or in part, only for those purposes but may not and

must not allow any other person to publish, copy or reproduce this Report in whole or in part without Optiro’s prior

written consent.

Unless Optiro has provided its written consent to the publication of this Report by Mawson West Limited for the purposes

of a transaction, disclosure document or a product disclosure statement issued by Mawson West Limited pursuant to the

Corporations Act 2001 (Cth), Securities Act (Canada) or the rules of any relevant exchange, then Optiro accepts no

responsibility to any other person for the whole or any part of this Report and accepts no liability for any damage,

however caused, arising out of the reliance on or use of this Report by any person other than Mawson West Limited.

While Optiro has used its reasonable endeavours to verify the accuracy and completeness of information provided to it by

Mawson West Limited and on which it has relied in compiling the Report, it cannot provide any warranty as to the

accuracy or completeness of such information to any person.

http://www.optiro.com/


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Technical Report on the Dikulushi Underground Project, Democratic Republic of Congo

A technical report on the Underground Project

Prepared for

Mawson West Limited

Authors

Andrew Law Director –Mining, Optiro Pty Ltd MMin; MBA; FAusMM; FIQA; MAICD

Ian Glacken Director –Geology, Optiro Pty Ltd BSc (Hons) (Geology); MSc (Mining

Geology), MSc (Geostatistics),

FAusIMM(CP), CEng

Date of report: 12 December 2013


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TABLE OF CONTENTS

1. SUMMARY 13

1.1. LOCATION 13

1.2. OWNERSHIP 14

1.3. MINERALISATION 15

1.4. MINERAL RESOURCES & RESERVES 15

1.5. MINING 17

1.6. METALLURGICAL 18

1.7. ECONOMIC ANALYSIS 19

1.8. ENVIRONMENTAL 19

1.9. CONCLUSIONS AND RECOMMENDATION 19

2. INTRODUCTION 20

2.1. SCOPE OF THE REPORT 20

2.2. AUTHORS 20

2.3. PRINCIPAL SOURCES OF INFORMATION 21

2.4. SITE VISIT 22

2.5. INDEPENDENCE 23

2.6. ABBREVIATIONS AND TERMS 23

3. RELIANCE ON OTHER EXPERTS 30

4. PROPERTY DESCRIPTION AND LOCATION 31

4.1. DEMOGRAPHICS AND GEOGRAPHIC SETTING 31

4.2. PROJECT OWNERSHIP 31

4.3. PROPERTY LOCATION 31

4.4. THE PROPERTY TENEMENT AREA 32

4.5. ENVIRONMENTAL PERMITS 34

5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND

PHYSIOGRAPHY 36

5.1. ACCESS 36

5.2. SITE TOPOGRAPHY, ELEVATION AND VEGETATION 36

5.3. CLIMATE, PHYSIOGRAPHY, LOCAL RESOURCES AND INFRASTRUCTURE 37

5.4. SURFACE RIGHTS 37

5.5. SITE INFRASTRUCTURE 37

5.5.1. WATER SUPPLY 38

5.5.2. POWER SUPPLY 38

5.5.3. MINE PERSONNEL 39

5.5.4. TAILINGS STORAGE FACILITY 39

5.5.5. ADMINISTRATION AND PLANT SITE BUILDINGS 39

5.5.6. ACCOMMODATION 40

5.5.7. COMMUNICATIONS 41

5.5.8. MOBILE EQUIPMENT 41


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5.5.9. SECURITY 42

6. HISTORY 44

6.1. BELGIAN EXPLORATION 44

6.2. ANVIL MINING LTD 44

6.3. MAWSON WEST 45

6.4. RESOURCE HISTORY 45

6.5. PRODUCTION HISTORY 46

7. GEOLOGICAL SETTING AND MINERALISATION 47

7.1. REGIONAL SETTING 47

7.2. PROJECT GEOLOGY 50

8. DEPOSIT TYPES 52

9. EXPLORATION 53

9.1. BRGM 53

9.2. ANVIL MINING LTD EXP LORATION 53

9.3. MWL EXPLORATION 54

10. DRILLING 55

10.1. INTRODUCTION 55

10.2. ANVIL PROGRAMME 1997 55

10.3. ANVIL PROGRAMMES 200 2 & 2003 56


10.5. ANVIL PROGRAMME 2005 /6 56



10.8. MWL PROGRAMME 2010 57

10.9 . SURVEY CONTROL 58

10.10. DRILLING ORIENTATION 58

11. SAMPLE PREPARATION, ANALYSIS AND SECURITY 59

11.1. DIAMOND CORE SAMPLING 59

11.1.1. DIAMOND CORE RECOVERY 59

11.1.2. DIAMOND CORE LOGGING 59

11.2. RC SAMPLING AND LOGGING 60

11.3. SAMPLE QUALITY 60

11.4. SAMPLE PREPARATION A ND ANALYTICAL PROCED URES 60

11.4.1. ANALYSES 60

11.5. BULK DENSITY DETERMI NATIONS 61

11.6. SAMPLE QAQC 61

11.6.1. STANDARDS AND BLANKS 61

11.6.2. LABORATORY QAQC 63

11.7. SUMMARY STATEMENT 63


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12. DATA VERIFICATION 64

13. MINERAL PROCESSING AND METALLURGICAL TESTING 65

13.1. INTRODUCTION 65

13.2. ANVIL TESTWORK 65

13.2.1. EARLY TESTWORK 65

13.2.2. LATER TESTWORK 67

13.3. PLANT OPERATIONAL RESULTS 71

13.4. METALLURGICAL PROPERTIES OF THE CUT BACK ORE AND UNDERGROUND ORE 71

14. MINERAL RESOURCE ESTIMATES 75

14.1. DIKULUSHI MINERAL RESOURCE ESTIMATE 75

14.1.1. GEOLOGICAL AND MINERALISATION MODELS 77

14.1.2. DRILL DATA FOR MINERAL RESOURCE MODELLING 78

14.1.3. DATA VALIDATION 79

14.1.4. DATA PREPARATION FOR MODELLING 79

14.1.5. DATA COMPOSITING 80

14.1.6. STATISTICS 81

14.1.7. SPATIAL STATISTICS 81

14.1.8. BLOCK MODEL 84

14.1.9. DENSITY ESTIMATES IN THE BLOCK MODEL 85

14.1.10. DETERMINATION OF TOP CUTS 85

14.1.11. GRADE ESTIMATION 85

14.1.12. ORDINARY KRIGING INTERPOLATION 85

14.1.13. MODEL VALIDATION 86

14.1.14. MINERAL RESOURCE CLASSIFICATION 88

14.1.15. RESOURCE TABULATION AND INVENTORY 89

14.2. MINERAL RESOURCE ESTIMATE COMPARISONS 89

14.2.1. MINERAL RESOURCE STATEMENT AUGUST 2011 VERSUS OCTOBER 2007 89

14.2.2. DEPLETION OF AUGUST 2011 MINERAL RESOURCES BY AUGUST 2013 OPEN PIT CUT BACK 91

14.3. KAZUMBULA MINERAL RESOURCE ESTIMATE 94

14.3.1. GEOLOGICAL AND MINERALISATION MODELS 94

14.3.2. DRILL DATA FOR MINERAL RESOURCE MODELLING 95

14.3.3. DATA VALIDATION 96

14.3.4. DATA PREPARATION FOR MODELLING 97

14.3.5. STATISTICS 97

14.3.6. SPATIAL STATISTICS 97

14.3.7. BLOCK MODEL 97

14.3.8. DENSITY ESTIMATES IN THE BLOCK MODEL 99

14.3.9. GRADE ESTIMATION 99

14.3.10. MODEL VALIDATION 100

14.3.11. MINERAL RESOURCE CLASSIFICATION 100

15. MINERAL RESERVE ESTIMATES 101

15.1. DEPLETION OF THE OPEN PIT RESERVES 101

15.2. UNDERGROUND MINE DESIGN AND SCHEDULE BASIS 101


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15.2.1. EXISTING WORKINGS 101

15.2.2. BASIS OF THE UNDERGROUND DESIGN AND SCHEDULE 103

15.3. CUT-OFF GRADE CRITERIA 103

15.4. MINING RECOVERY AND DILUTION 106

15.5. UNDERGROUND MINERAL RESERVE TABULATION 107

16. MINING METHODS 108

16.1. HISTORICAL MINING 108

16.2. PROPOSED MINING METHOD – CUT AND FILL 108

16.2.1. OVERHAND CUT AND FILL 108

16.2.2. UNDERHAND CUT AND FILL 109

16.2.3. MINING OF WIDER SECTIONS OF THE OREBODY 110

16.2.4. PROPOSED MINING METHOD – EXTRACTION OF THE CROWN PILLAR 112

16.3. GEOTECHNICAL DESIGN PARAMETERS 113

16.3.1. STOPE LAYOUT AND SEQUENCE 113

16.3.2. DRILL AND BLAST 113

16.3.3. ORE EXTRACTION 113

16.3.4. BACKFILLING 114

16.3.5. ACCESSING THE OREBODY & REHABILITATION OF OLD WORKINGS 116

16.4. VENTILATION 117

16.4.1. PRIMARY VENTILATION 117

16.4.2. SECONDARY VENTILATION 119

16.5. DEWATERING 119

16.6. MINING EQUIPMENT 121

16.6.1. MINE DEVELOPMENT 122

16.6.2. MINING SCHEDULE 123

16.6.3. MINING SHIFTS 128

16.6.4. DEVELOPMENT / STOPING RATES 131

16.6.5. AIR LEG DEVELOPMENT RATES 132

16.7. GEOTECHNICAL 133

16.7.1. DATA 133

16.7.2. GEOTECHNICAL DOMAINS 133

16.7.3. POTENTIAL FAILURES 135

16.7.4. MAPPING, MONITORING AND ADDITIONAL DATA 135

16.8. GROUND SUPPORT REQUI REMENTS 135

16.8.1. SPLIT SETS 137

16.8.2. SOLID STEEL ROCKBOLTS 137

16.8.3. CABLE BOLTS 137

16.8.4. SHOTCRETE 137

16.8.5. GEOTECHNICAL FILL REVIEW 138

16.8.6. CRF MIXING 138

16.9. GROUND SUPPORT STAND ARDS 139

16.9.1. DECLINE SUPPORT STANDARD 140

16.9.2. ACCESS SUPPORT STANDARD 141

16.9.3. ORE DRIVE SUPPORT STANDARD WITH MESH 142

16.9.4. 3 WAY INTERSECTION SUPPORT STANDARD 143

16.9.5. 4-WAY INTERSECTION SUPPORT STANDARD 144


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16.10. WASTE DUMP DESIGN 145

16.11. SURFACE WATER MANAGE MENT 145

17. RECOVERY METHODS 146

17.1. PLANT FLOWSHEET 146

17.2. TAILINGS STORAGE FACILITIES (TSF) 147

17.3. PROCESSING STATISTICS 148

18. PROJECT INFRASTRUCTURE 151

18.1. SURFACE FACILITIES 151

18.2. POWER 152

18.3. PROCESS WATER SUPPLY 152

19. MARKET STUDIES AND CONTRACTS 154

19.1. MARKETS 154

19.2. CONTRACTS 154

20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT156

21. CAPITAL AND OPERATING COSTS 158

21.1. CAPITAL COST ESTIMATE 158

21.2. OPERATING COST ESTIMATE 159

21.2.1. MINING OPERATING COST 159

21.2.2. PROCESSING OPERATING COSTS 160

21.2.3. MANAGEMENT AND ADMINISTRATION COSTS 160

21.2.4. TRANSPORT AND SMELTING COSTS 161

21.3. METAL PRICES 161

22. ECONOMIC ANALYSIS 161

22.1. OPERATIONS SUMMARY 161

22.1.1. SENSITIVITY ANALYSIS 166

22.2. PAYBACK 166

22.3. MINE LIFE 167

22.4. TAXATION 167

23. ADJACENT PROPERTIES 168

24. OTHER RELEVANT DATA AND INFORMATION 169

25. INTERPRETATION AND CONCLUSIONS 170

26. RECOMMENDATIONS 171


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27. REFERENCES 172

28. CERTIFICATES 174

TABLES

Table 1.1 Dikulushi Mineral Resource statement as at August 2011, using a 1.0% copper cut-off grade 15

Table 1.2 Depleted Dikulushi Mineral Resource statement as at August 2013, using a 1.0% copper cut-

off grade 15

Table 1.3 Dikulushi Mineral Reserve statement as at August 2011, at a 1.0% copper cut-off grade 16

Table 1.4 Depleted Dikulushi Mineral Reserve statement as at August 2013, at a 1.0% copper cut-off

grade 16

Table 1.5 Dikulushi Underground Mineral Reserve statement as at September 2013 16

Table 1.6 The Kazumbula Mineral Resource statement as at November 2010 17

Table 2.1 Glossary of terms 24

Table 4.1 Mawson West Limited Dikulushi tenement schedule 33

Table 6.1 Historical work summary at the Dikulushi Project 44

Table 6.2 Mineral Resource estimate as completed by FinOre in July 2006 and published in December

2006; a cut-off grade of 1.5% copper was used 46

Table 6.3 Historical Anvil production for the Dikulushi mine 46

Table 6.4 Recent MWL production for the Dikulushi mine 46

Table 9.1 Historical drilling summary for the Dikulushi copper silver project 53

Table 10.1 MWL drilling at Kazumbula 57

Table 13.1 Details of Dikulushi drill core used in Mintek metallurgical testing 66

Table 13.2 Head grades of chalcocite composites 67

Table 13.3 Relative abundance of significant minerals 68

Table 13.4 Comminution testwork results 68

Table 13.5 Effect of grind size on flotation performance (high grade chalcocite) 68

Table 13.6 Effects of collector addition on flotation performance (high grade chalcocite) 69

Table 13.7 Effect of grind size on flotation performance (disseminated and low grade chalcocite) 69

Table 13.8 Effect of collector addition on flotation performance (disseminated and low grade chalcocite) 69

Table 13.9 Effect of grind size and Eh level on flotation performance (Pb/Zn rich chalcocite) 69

Table 13.10 Head grades of chalcocite composites 70

Table 13.11 Locked cycle flotation test results 70

Table 13.12 Dikulushi processing summary (February 2007 – April 2008) 72

Table 13.13 Dikulushi processing summary (June 2010 – July 2013) 72

Table 14.1 Dikulushi Mineral Resource statement as at August 2011 above a 1.0% copper cut-off grade 76

Table 14.2 Domain codes for Dikulushi modelling 80

Table 14.3 Summary statistics for copper % and silver g/t per domain 81

Table 14.4 Dikulushi variogram models with angle1 about axis 3 (Z), angle2 about axis 1 (X) and angle3

about axis 3 (Z) 83

Table 14.5 Dikulushi - top cuts per domain 85

Table 14.6 Mean statistics per domain comparing model estimates with data values 86

Table 14.7 Dikulushi Mineral Resource statement using a 1.0% copper cut-off grade as at August 2011 89

Table 14.8 Comparison of 2011 and 2007 Dikulushi Mineral Resource estimates 90


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Table 14.9 Comparison of August 2011 and August 2013 Dikulushi Mineral Resource estimates, showing

the Open pit cut back depletion 92

Table 14.10 MWL drilling at Kazumbula 96

Table 14.11 Summary statistics of the two metre composite data for Cu% and Ag g/t for the Kazumbula

deposit 97

Table 14.12 Density estimates for the Kazumbula deposit 99

Table 14.13 A table of mean statistics comparing model estimates with data values 100

Table 14.14 Kazumbula Mineral Resource statement as at November 2010. 100

Table 15.1 Dikulushi Mineral Reserve statement as at August 2011, using a 1.0% copper cut-off grade 101

Table 15.2 Depleted Dikulushi Mineral Reserve statement as at August 2013, using a 1.0% copper cut-off

grade 101

Table 15.3 Mining dilution table 106

Table 15.4 Dikulushi Mineral Reserve statement as at September 2013 107

Table 16.1 CRF Specifications 115

Table 16.2 CAF Specifications 116

Table 16.3 Dikulushi production mining equipment at site from previous mining activities 121

Table 16.4 Major mining fleet and equipment required for the extraction of the Dikulushi underground

Mineral Reserves 122

Table 16.5 Underground horizontal development design parameters 122

Table 16.6 Underground vertical development design parameters 123

Table 16.7 Mining dilution 126

Table 16.8 Underground mine production physicals 128

Table 16.9 Work shifts 129

Table 16.10 Operational Management Labour 129

Table 16.11 Technical Services labour 129

Table 16.12 Support functions labour 130

Table 16.13 Labour requirements: underground operations 130

Table 16.14 Underground Workshop personnel 131

Table 16.15 Jumbo/production drill rates by development type 131

Table 16.16 Jumbo/production drill rates by individual machine 132

Table 16.17 Jumbo/production drill rates by fleet 132

Table 16.18 Air Leg development 133

Table 17.1 Dikulushi processing summary relevant to ore to be mined in the pit cut back 149

Table 17.2 Processing statistics for the LG material completed by MWL – June 2010 to May 2011 149

Table 21.1 Dikulushi underground capital expenditure cost estimate. 158

Table 21.2 Major mining fleet and equipment required for the extraction of the Dikulushi underground

Mineral Reserves 159

Table 21.3 Mining overhead and fixed costs 160

Table 21.4 Mining variable costs 160

Table 21.5 Metal prices used in modelling 161

Table 22.1 Dikulushi mining and financial summary 163

Table 22.2 Sensitivity analysis on the cash flow forecast for underground mining and treatment at

Dikulushi 166

FIGURES

Figure 1.1 Locality plan of the Dikulushi Project 14

Figure 4.1 Exploration Licences of the Dikulushi copper silver project 32


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Figure 4.2 Dikulushi mine infrastructure within the PE 606 33

Figure 4.3 Dikulushi mine site aerial view 35

Figure 5.1 Barge on Lake Mweru 36

Figure 5.2 Composite Temperature and Rainfall Data recorded over the last 5 years. 37

Figure 5.3 Dikulushi airstrip and the G1 Charter plane provides safe staff transportation to and from site 38

Figure 5.4 Dikulushi Administration Centre 40

Figure 5.5 Dikulushi Camp Site 40

Figure 5.6 Dikulushi Clinic and communications centre 41

Figure 5.7 Dikulushi Workshop 42

Figure 5.8 Dikulushi Store and Fuel Farm 43

Figure 7.1 Regional Geology of Mawson’s convention area in the DRC 48

Figure 7.2 Stratigraphy of Dikulushi region with known styles of mineralisation 49

Figure 7.3 Local geology of the Dikulushi open pit 50

Figure 7.4 A typical vertical cross section through the Kazambula deposit, highlighting key geology

associated with mineralisation 51

Figure 11.1 GBM301-7 suggests accurate values around low value samples (~0.55% Cu) 62

Figure 11.2 The GBM301-8 is a high Cu value standard and suggests accurate results for high value

samples (~10% Cu) 62

Figure 11.3 The GBM398-4c is a low Cu value standard and suggests accurate results for low value

samples (~0.39% Cu) 62

Figure 11.4 Results for this blank demonstrate that contamination is well contained 63

Figure 13.1 Dikulushi Underground sources of ore - showing North-South section view at 50205E 73

Figure 14.1 An oblique southward looking 3D view of drillhole type and distribution at Dikulushi 76

Figure 14.2 A vertically oriented 3D view at Dikulushi, looking southwest, showing mineralisation lenses

and current drilling 77

Figure 14.3 A plan showing the distribution of drillhole types across Dikulushi; blasthole data from the pit

have been excluded 78

Figure 14.4 Quantile-Quantile (Q-Q) plot of Diamond (DD) drilled samples versus sludge drilled samples

within a common area 79

Figure 14.5 Cumulative distribution of sample lengths highlighting the dominant 1m sample length 80

Figure 14.6 Log histogram and probability plot for the main FW zone of mineralisation showing the results

of robust domaining 82

Figure 14.7 Variogram models for copper % across the FW zone of mineralisation 84

Figure 14.8 A plan view slice through the FW zone block model illustrating the good comparison between

model estimates and the nearby drillhole data 87

Figure 14.9 A statistical plot of estimates versus drillhole data grades for successive 30m increments in

elevation and the full strike length of the FW zone mineralisation 87

Figure 14.10 3D view of the Dikulushi model, looking south, and showing resource classification categories 88

Figure 14.11 A waterfall chart of cumulative Mineral Resource changes from 2007 to 2011 91

Figure 14.12 A waterfall chart of cumulative Mineral Resource changes from 2011 to 2013 93

Figure 14.13 Grade tonnage curves for the combined remaining Measured and Indicated Mineral

Resources 93

Figure 14.14 Kazumbula vertical section, looking north, highlighting the modelled mineralisation as per the

RC and diamond drilling 95

Figure 14.15 Plan showing the distribution of RC and diamond drillholes across the Kazumbula deposit. 96

Figure 14.16 Histogram and probability plots for the Kazumbula deposit two metre sample data. 98

Figure 14.17 Variogram modelling for Cu % in the plane of mineralisation. 99


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Figure 15.1 Existing workings, showing as built underground development (grey), and the as-built pit

(green) 102

Figure 15.2 Underground reserve, showing as-built underground development (grey), as-built pit (green),

and measured (purple) and indicated (red) Mineral Resources 103

Figure 15.3 Relationship between cut-off NSR and metal grades 106

Figure 16.1 Overhand cut and fill mining process 109

Figure 16.2 Underhand cut and fill mining process 110

Figure 16.3 Diagrammatic representation of sequential mining in wide orebody areas 111

Figure 16.4 Orebody access development 112

Figure 16.5 Pillar ratio diagram 112

Figure 16.6 LHD loader with ‘rammer-jammer attachment 114

Figure 16.7 Underground primary ventilation circuit (full) 118

Figure 16.8 Underground primary ventilation circuit required for the extraction of the measured and

indicated material only 118

Figure 16.9 Primary ventilation fan location 119

Figure 16.10 Existing underground dewatering infrastructure locations 120

Figure 16.11 Proposed underground dewatering infrastructure locations 120

Figure 16.12 Ore loss due to gaps left in the backfilling process 126

Figure 16.13 Ore level schedule, by quarter 128

Figure 16.14 Dikulushi orebody rock quality, Q (Turner, 2013) 134

Figure 16.15 Dikulushi footwall rock quality, Q (Turner, 2013) 134

Figure 16.16 Dikulushi hanging wall rock quality, Q (Turner, 2013) 135

Figure 16.17 Dikulushi rock reinforcement chart (Turner, 2013) 136

Figure 17.1 Dikulushi Plant flow diagram 147

Figure 18.1 On-site office facilities at Dikulushi 151

Figure 18.2 On-site Underground change room facilities at Dikulushi 151

Figure 18.3 Average water balance 153

Figure 20.1 Community Business making work clothes for the mine. 157


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1. SUMMARY

Mawson West Limited’s (MWL’s) Dikulushi Underground Project (the Project) is located in the

Katanga Province of the Democratic Republic of Congo (DRC). The Underground Mine comprises

Mineral Resources from the main Dikulushi deposit’s “Footwall” zone, which has a 230 m strike

length and true widths of up to 25 m. The open pit was recently completed as a cut back extension

of the old Dikulushi open pit mined by Anvil Mining Limited (Anvil) during its tenure of the Dikulushi

deposit. MWL has now completed an underground pre-feasibility study to re-enter and re-establish

the old underground workings and mine out the previously developed high grade Mineral Reserves

as a first stage. This study is the focus of this Technical Report, which also details information

regarding the associated Kazumbula project.

In addition to the mining of the remaining developed Mineral Reserves during the first stage, MWL

will continue to explore and evaluate depth extensions of the remaining underground Inferred

Mineral Resource. This will be done through additional underground drilling from within the re-

established Dikulushi underground workings and, once completed, will form the basis of further

underground feasibility study work based on the additional drilling and Mineral Resource evaluation

outcomes.

1.1. LOCATION

The Project is located at latitude 08°53’37.7 south and longitude 28°16’21.8 east in the south

eastern corner of the DRC, approximately 50 km north-northwest of the small town of Kilwa and

situated on the south western side of Lake Mweru (Figure 1.1).


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Figure 1.1 Locality plan of the Dikulushi Project

1.2. OWNERSHIP

The Dikulushi mine is governed by the Dikulushi Mining Convention signed on January 31, 1998 with

the Government of the DRC, and ratified by Presidential Decree issued on February 27, 1998.

The Dikulushi Mining Convention is a mining concession granted to Anvil Mining Congo SARL (AMC)

which sets out the regulatory and fiscal regime applicable to the tenements owned by AMC.

Mawson West Investments Ltd, a wholly owned subsidiary of Mawson West Limited, holds 90% of

the issued capital of AMC, with the remaining 10% being held by the Dikulushi–Kapulo Foundation

NPO.

Mining operations at Dikulushi are currently conducted under the Exploitation Permit 606 (PE)

issued by Ministerial Decree under the terms of the Dikulushi Mining Convention. This guarantees

the sole and exclusive rights to the benefit of the holding company for 20 years until 2022. The

Dikulushi deposit forms part of the PE.

This report presents technical information on the Dikulushi deposit, relating to the recently

completed open pit cut back and, more particularly, to the planned re-establishment of the

underground workings and trial stoping of the previously developed levels. Additionally, further

exploration and drilling of the Inferred Mineral Resource and currently unclassified material will also

be undertaken from the re-established underground workings.


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1.3. MINERALISATION

The Dikulushi copper deposit is interpreted to be a hypogene, fault-controlled orebody, comprising

disseminated, brecciated and massive chalcocite-bornite mineralisation with a supergene weathered

and oxidised zone of semi-massive malachite, azurite and nodular cuprite. Most of the oxidised

portion of the Dikulushi deposit has been mined out.

1.4. MINERAL RESOURCES & RESERVES

The current Mineral Resource of the Dikulushi orebody has been derived from a mineralisation

interpretation based upon copper drillhole grades. A block model estimate was completed in May

2009 by David Gray of Optiro and was depleted in August 2011 with updated surveyed volumes of

historical mining. The resulting Mineral Resource is stated for a 1.0% copper cut-off grade in Table

1.1.

Table 1.1 Dikulushi Mineral Resource statement as at August 2011, using a 1.0% copper cut-off grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured Mineral Resources 184 2.8 516 7.0 211

Indicated Mineral Resources 90 2.8 251 5.6 114

Total Measured and Indicated Mineral Resources 274 2.8 767 6.6 179

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Inferred Mineral Resources 136 2.8 380 6.8 91

The inferred silver grade was incorrectly reported previously at 91 g/t; the correct grade is 155g/t

The resulting estimates are supported by historical production and current processing grades.

The August 2011 Mineral Resource from Table 1.1 has now been depleted by the open pit cutback.

The remaining Mineral Resources are stated for a 1.0% copper cut-off grade in Table 1.2.

Table 1.2 Depleted Dikulushi Mineral Resource statement as at August 2013, using a 1.0% copper cut-off grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)




Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)


The inferred silver grade was incorrectly reported at 91 g/t in the August 2011 Mineral Resource table and should have

been 155g/t. This has now been corrected and adjusted accordingly in the depleted Mineral Resource.

The open pit Mineral Reserves, as published 16 September 2011 and revised 8 January 2013, are

shown in Table 1.3 and are stated for a 1.0% copper cut-off grade. Mineral Resources are reported

as inclusive of Mineral Reserves. The Mineral Reserve, as per the CIM definition, incorporated

mining losses and dilution material brought about by the mining operation.


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Table 1.3 Dikulushi Mineral Reserve statement as at August 2011, at a 1.0% copper cut-off grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000) Copper (%)

Silver

(g/t)

Proven 66.6 2.8 184.7 7.27% 207

Probable 127.8 2.8 354.3 5.51% 169

Total Proven and Probable Reserves 194.4 2.8 539.0 6.12% 182

The open pit Mineral Reserves have now been depleted with the mining of the open pit cut back as

this was completed during July 2013. The open pit Mineral Reserves were based on the open pit

reaching the 810 mRL. Mining ceased at the 825 mRL following some isolated sections of the pit wall

deteriorating beyond what was predicted. Table 1.4 shows the depleted Mineral Reserves post

cessation of mining of the open pit cut back.

Table 1.4 Depleted Dikulushi Mineral Reserve statement as at August 2013, at a 1.0% copper cut-off grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000) Copper (%)

Silver

(g/t)

Proven 1.0 2.8 2.7 6.8 186

Probable 29.9 2.8 83.7 5.5 188

Total Proven and Probable Reserves 30.9 2.8 86.4 5.5 188

The above remaining Mineral Reserves have subsequently been incorporated into the Underground

Mineral Reserves, which are presented in Table 1.5 below and are now based on an NSR value cut

off value of US$329/t, using a copper price of US$3.08/lb and a Silver price of US$20 per oz. Mineral

Resources are inclusive of Mineral Reserves. The Mineral Reserve, as per the CIM definition,

incorporates mining losses and dilution material expected to be incurred through the underground

mining operation.

Table 1.5 Dikulushi Underground Mineral Reserve statement as at September 2013

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000) Copper (%)

Silver

(g/t)

Proven 0 0 0 0 0

Probable 62 2.8 173 5.2 127

Total Proven and Probable Reserves 62 2.8 173 5.2 127

Notes:

1) The reporting cut-off grade is based on an NSR value of US$329/t, using a copper price of

US$3.08/lb and a Silver price of US$20 per oz.

2) The above Mineral Reserve does not include any Inferred Mineral Resources.

The Mineral Reserves detailed above are derived from the depleted Measured and Indicated Mineral

Resources that remain below the open pit floor at the 825 m RL, and which can be economically

extracted based on the modifying factors as compiled in the underground pre-feasibility study.

The Kazumbula orebody was originally drilled by Anvil. MWL has developed confidence in this

deposit’s grade and geological continuity by drilling additional reverse circulation (RC) and HQ3

diamond core during 2010. A litho-structural and grade based interpretation was completed by


P a g e | 17

MWL geological staff. The Mineral Resources for Kazumbula effectively use a 0.5% copper cut-off

grade for defining the mineralised volume and are shown in Table 1.6 below.

Table 1.6 The Kazumbula Mineral Resource statement as at November 2010

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)


1.5. MINING

With the completion of the open pit cut back, continuation of mining activities at the Dikulushi

project will now focus on re-establishing the underground mining operation at the deposit in order

to exploit the previously-developed workings from underground mining activities conducted by

previous owner of the mine, Anvil.

Re-commencement of the underground mining activities requires the rehabilitation and re-

establishment of the ventilation, electrical, air and water services. The mine is currently being

dewatered, which is nearing completion. In addition to the re-establishment of services, additional

ground support checks and repairs will be required to ensure that the existing underground

development is to the required standard in order to allow safe access to the underground mining

areas. Initial rehabilitation of the underground workings is expected to take approximately one

month, and will be tied in with the development of service infrastructure requirements for the

underground operations which includes escape way rises, return airway rises, and the installation of

the primary ventilation fans.

Production activities will commence in levels that were partially complete from previous

underground mining activities. Work on these levels will extract the remaining ore contained within.

Once the ore has been extracted from these levels, each of the drives will be backfilled using

cemented fill. Development activities will also commence in other parts of the underground mine,

establishing new levels for production.

Ore extraction from the underground will be completed using overhand and underhand cut and fill

mining practices, with the bulk of the ore being removed using the overhand mining method.

Extraction of the ore between the upper levels of the underground workings and the bottom of the

pit (the crown pillar) will be completed using a long hole stoping method.

In addition to the recommencement of underground mining activities, additional exploration drilling

is planned to be undertaken in order to upgrade the known Inferred Mineral Resource to a

Measured and/or Indicated classification and possibly extend the depth of the Mineral Resource.

The crown pillar extraction is planned to take place on a retreat method as a final operation prior to

closing the underground on completion of the extraction of the current Mineral Reserves. Should

the planned exploration drilling upgrade and extend the additional areas of the Mineral Resource

classification, and thus the Mineral Reserves, mining of the crown pillar ore tonnes will need to be

deferred.


P a g e | 18

The current underground life of mine is 19 months, including one month for the re-establishment of

the existing underground workings. The average monthly ore production rate over the life of mine is

approximately 7,200 t at an average Copper grade of 5.15% Cu and a Silver grade of 127 g/t Ag. The

production tonnes and grade quoted include the extraction of the crown pillar ore during the last 6

months of the mining schedule.

1.6. METALLURGICAL

Several metallurgical testwork programmes have been completed by Anvil on the Dikulushi ore and

are discussed in Chapter 13. These results are appropriate for deposits with similar styles of

mineralisation, such as Kazumbula, and have subsequently been compared against actual production

results during the period of operation by Anvil and more recently by MWL.

The most recent metallurgical testwork was managed by Sedgman Metals, a metallurgical consulting

company of Perth, Western Australia. Testwork was completed at AMDEL Laboratories in Perth.

Metallurgical testwork was carried out previously by Anvil on the main Dikulushi orebody.

Additional testwork was reported on in June 2004 by Independent Metallurgical Laboratories (IML),

which utilised samples provided from the mill feed and an open pit sample to perform a locked cycle

flotation test. Results indicated that from a feed grade of 8.76% copper and 306 g/t silver a recovery

of 91.1% copper and 89.7% silver could be achieved to produce a concentrate with grades of 42.1%

copper and 1,447 g/t silver. This sample contained 18% acid soluble copper in feed. Actual

production results during operations by Anvil were higher.

The float plant at Dikulushi operated from 2007 to 2008, was fed with high grade ore from the open

pit and underground mine, and yielded recoveries of 90.4% copper and 90.3% silver, producing a

concentrate with 55.5% copper and 1721 g/t silver.

The current plant, under MWL control over the past year, has been fed from the open pit cut back

and low grade stockpiles, with recoveries averaging 91.5% copper and 90.3% silver, producing a

concentrate with 56.5% copper and 1,515 g/t silver.

Plant operation under MWL over the past 6 months of production (Feb 2013 to July 2013) has seen

fresh ore feed from the open pit cut back, combined with improved operating management

practises, resulting in improved recoveries of 94.3% copper and 92.1% silver, producing a

concentrate with 61.4% copper and 1,768 g/t silver.

There has been no change in the material ore types since the previous open pit and underground

operations and it is therefore expected that the current recoveries being achieved for the fresh ore

from the open pit cut back feed will continue to be achieved with the re-establishment of the

underground operations.

The financial model uses 94% recovery for copper and 90% for silver, with a copper concentrate

grade of 60% copper.


P a g e | 19

1.7. ECONOMIC ANALYSIS

In summary, the underground operation will produce some 174, 000 tonnes of ore at a copper grade

of 5.15g/t and a silver grade of 127 g/t over a 19 month production period. This will produce some

8,100 tons of copper metal and 573,500 ounces of Silver for sale. Total operating costs are

estimated at $44.6M ($257 per ore tonne milled or $1.86 per lb of copper - net of the silver credit);

and the total net revenue is estimated at $57.7M, with net cashflow totalling $3M. The Internal

Rate of Return is 8% for the project, and the Capital cost for the project is $9.6M.

1.8. ENVIRONMENTAL

An Environmental Impact Assessment (EIA) for the Dikulushi project was lodged in 2003. In 2009, an

EIA for the underground Project was submitted to the DRC Government. Both of these reports were

compiled by African Mining Consultants of Kitwe, Zambia, an environmental company that was

licensed to work and report in the DRC. In 2011, an EIA for the cutback project was prepared by

EMIS sprl, a DRC environmental company licenced to work and report in the DRC. All three

environmental reports received DRC Government approval. A revised EIA, extending underground

mining beyond 2013, has been submitted to Government.

MWL has lodged $1.19M as an environmental bond. This financial guarantee is a contribution

towards environmental rehabilitation costs for the Dikulushi mine.

1.9. CONCLUSIONS AND RECOMMENDATION

The Project is at an advanced stage and Dikulushi may be described as a producing and developing

property. MWL has completed a pre-feasibility study in order to determine the economics of

continuing to mine the Dikulushi deposit via the previously established and developed underground

workings. Since this was previously an operating open pit and underground mine, the remaining ore

zones present the same risks as before, being somewhat mitigated for the mineralogy, metallurgical

properties and the processing aspects, which are well known. Risks associated with the mining

operations will remain, however, and constant recognition of changing conditions will need to be

ensured with appropriate changes made as mining progresses. Geotechnical knowledge will

increase with the physical mining activities and a better understanding of the underground ground

conditions will be established. There is likely to be continued resource development drilling

throughout the mining operations in order to locate and evaluate additional resources associated

with the same ore zone, either at depth or as lateral or parallel extensions. During the period

required to re-establish the underground workings and re-commence development and stoping

operations, MWL intends to continue processing the HG open pit cut back stockpile. In addition

MWL is currently in the process of defining additional deposits on the Dikulushi property and within

50 km of the Dikulushi plant.


P a g e | 20

2. INTRODUCTION

2.1. SCOPE OF THE REPORT

Mawson West Limited (MWL) commissioned Optiro Pty Ltd (Optiro) in May 2013 to review the

underground pre-feasibility study, generated by MWL, and to prepare an independent technical

report regarding the copper-silver Mineral Reserves at the Dikulushi underground deposit based on

the above study. This Technical Report has been written to comply with the reporting requirements

of the Canadian National Instrument 43-101 guidelines, “Standards of disclosure for Mineral

Properties” dated April 2011 (the Instrument) and with the “Australasian Code for Reporting of

Mineral Resources and Ore Reserves” of December 2004 (the JORC Code) as produced by the Joint

Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute

of Geoscientists and Minerals Council of Australia (JORC 2004).

The Technical Report has been written to provide the market with an update on the status of the

Mineral Resources and Reserves for the Dikulushi Open Pit cut back project (mining now complete

and processing due for completion late 2013) and to present the first stage of the underground

Project study, which is to re-enter and re-establish the old underground workings and to mine out

the previously developed high grade Mineral Reserves. This is the focus of this Technical Report.

For completeness, the Mineral Resource estimation of the related but separate Kazumbula project is

also described.

All monetary amounts expressed in this report are in United States of America dollars (US$) unless

otherwise stated.

2.2. AUTHORS

The key authors for compiling this report are:

Mr Andrew Law is the principal author and Qualified Person and takes overall responsibility

for this report. Mr Law is the Director - Mining at Optiro and is a professional Mining

Engineer. He has a HND Metalliferous Mining (1982) and an MBA from the University of

Western Australia. He has more than 30 years’ experience in the planning, development

and extraction of mineral reserves. Mr Law is a Fellow of the Australasian Institute of Mining

and Metallurgy (FAusIMM) and has the relevant qualifications, experience and

independence to be considered as a “Qualified Person” as defined in Canadian National

Instrument 43-101. Mr Law has visited the Dikulushi deposit (February 2012) and the

underground workings to the 830 m RL. Mr Law was a previous author for the Dikulushi

Open Pit Cut back NI 43-101 report generated by Optiro for MWL. Mr Law has reviewed all

sections of the “Pre-Feasibility” study generated by various other Qualified Persons, most of

whom were independent of Mawson West, and collated into a pre-feasibility study by MWL.

Mr Ian Glacken is a Qualified Person and takes responsibility for the Mineral Resources

estimation portion of this report. Mr Glacken, is a full time employee of Optiro, where he

holds the position of Geology Director, and is a professional Geologist. Mr Glacken has


P a g e | 21

degrees from Durham University (BSc (Hons) Geology, 1979), The Royal School of Mines

(MSc Mineral Exploration, 1981), Stanford University (MSc Geostatistics, 1996) and Deakin

University (Postgraduate Diploma of Computing, 1996). Mr Glacken is a Fellow of the

Australasian Institution of Mining and Metallurgy (Member Number 107194) and a

Chartered Professional Geoscientist of that Institution. He is also a Member of the Institute

of Mining, Metallurgy and Materials (UK) and a Chartered Engineer of that institution. Mr

Glacken has the relevant qualifications, experience and independence to be considered as a

“Qualified Person” as defined in Canadian National Instrument 43-101. Mr Glacken has not

visited the Dikulushi deposit but has reviewed and supervised Mineral Resource models on

the Dikulushi deposits. Optiro is an Australian-based mining and resources consulting and

advisory firm which provides a broad range of expert services and advice, locally and

internationally, to the minerals industry and financial institutions.

In September 2011 Optiro generated and supervised Mineral Resource and Reserves models for the

Dikulushi open pit cut back. In August 2013, Optiro depleted the Mineral Resource and

subsequently the open pit cut back Mineral Reserves. With the completion of the open pit cut back,

Optiro has now generated underground Mineral Reserves for the Dikulushi deposit based on the

depleted Mineral Resources as at August 2013. Optiro is an Australian based mining and resources

consulting and advisory firm which provides a broad range of expert services and advice, locally and

internationally, to the minerals industry and financial institutions.

The following authors contributed to the report:

Name Position NI 43-101 Contribution

Andrew Law Director-Mining, Optiro Pty Ltd Principal Qualified Person

Ian Glacken Director – Geology, Optiro Pty Ltd Qualified Person and contributing

author of sections 1, 7.8, 9, 10, 11, 12

& 14.

Mike Turner Turner Mining and Geotechnical Pty Ltd Geotechnical, QP and author of

geotechnical submission in section 16

Duncan Grant-Stuart Knight Piesold Consulting Engineer, QP and reviewer of tailings

storage facilities in section 17

Peter Hayward Sedgman Ltd Metallurgical, QP and input into

section 13 and 17.

2.3. PRINCIPAL SOURCES OF INFORMATION

The principal source of information used to prepare this report is the information prepared for the

development of the pre-feasibility study and the previously submitted NI 43-101 Technical Reports

covering Mineral Resources and Reserves at Dikulushi. This pre-feasibility information was provided

to Optiro by MWL. The Mineral Resource information has been sourced from the previously

submitted NI 43-101 Technical Report, by Optiro, on the Dikulushi Project, Democratic Republic of

Congo, 16 September 2011 and revised 8 January 2013. The Mineral Resource has recently

undergone a review and depletion process based on the recently completed open pit cut back.

In summary, the following are primary data sources:


P a g e | 22

the NI 43-101 Technical Report on the Dikulushi Project, Democratic Republic of Congo,

February 3, 2011 and subsequently revised March 7, 2011

historical and current production and processing data

the NI 43-101 Technical Report on the Dikulushi Open Pit Project, Democratic Republic of

Congo, issued 16 September, 2011 and Revised 8 January, 2013.

A pre-feasibility study for the underground prepared by Mawson West based on inputs from

various independent qualified persons.

Optiro has made all reasonable enquiries to establish the completeness and authenticity of the

information provided. In addition, a final draft of this report was provided to MWL along with a

written request to identify any material errors or omissions prior to lodgement. The following

professionals have been consulted for relevant detail contained in this report.

Name Company Pre-Feasibility Contribution

Greg Entwistle Mawson West Ltd Operational Management Review

Chris Marissen Mawson West Ltd Mining

Gary Brabham Mawson West Ltd Geological

Mike Turner Turner Mining and Geotechnical Pty Ltd Geotechnical

Duncan Grant-Stuart Knight Piesold Consulting Tailings storage facilities

Peter Shephard SRK Consulting Hydrology and Water Management

Peter Hayward Sedgman Ltd Metallurgical

Andries Strauss Knight Piesold Consulting Tailings storage facilities

Glen Zamudio Mawson West Ltd Commercial

2.4. SITE VISIT

Mr Andrew Law visited the Dikulushi Project in February 2012 and specifically visited the

underground decline and openings that were available at the time (approx. 830 mRL). He has now

reviewed all sections of the Pre-Feasibility study collated by MWL and generated by various

Qualified Persons, many of whom were independent of MWL.

Mr David Gray (a former employee of Optiro and a QP for previous Dikulushi Technical Reports)

completed a comprehensive site visit to the Dikulushi copper Project in November 2010. The

purpose of this visit was to:

verify the relative size, position and presence of copper mineralisation at the Dikulushi and

Kazumbula deposits

verify the presence and position of drillhole sampling for the respective resources and

reserves

inspect the drill core for mineralisation, geological relationships with mineralisation and

general sample quality

review the respective sampling methods and QAQC with onsite geologists

review and confirm sample and assay data as stored in the drillhole database

review historical and current production and processing data.

Mr David Gray did not take independent samples due to the operational nature of the respective

resources and the visible in-situ mineralisation which confirms drillhole sample results. Mr Ian


P a g e | 23

Glacken, Director-Geology at Optiro, has not visited the Dikulushi Operation, but has nonetheless

supervised and peer reviewed the Dikulushi Project work complied by Mr Gray since 2009, and now

accepts responsibility for the Mineral Resources estimation as stated in this report.

Site visits have been carried out by the following persons:

Name Company Section Date of Visits

Andrew Law Optiro Mineral Reserves February 2012

David Gray Optiro Resource NI 43-101 November 2010

Chris Marissen Mawson West Ltd Mining Various as employee of MWL

Gary Brabham Mawson West Ltd Geology Various as employee of MWL

Mike Turner Turner Mining and

Geotechnical Pty Ltd

Geotechnical December 2012

Duncan Grant-Stuart Knight Piesold Consulting Tailings storage facility July 2010

Peter Hayward Sedgman Ltd Metallurgical February 2012

Peter Shephard SRK Consulting Hydrology, Water Management Once during 2007

Glen Zamudio Mawson West Ltd Commercial Various as employee of MWL

2.5. INDEPENDENCE

Neither Mr Andrew Law or Mr Glacken, nor Optiro, have or have had any material interest in MWL

or its related entities or interests. This report has been prepared in return for fees based upon

agreed commercial rates and the payment of these fees is in no way contingent on the results of this

report.

2.6. ABBREVIATIONS AND TERMS

A listing of abbreviations and terms used in this report is provided in Table 2.1 below.


P a g e | 24

Table 2.1 Glossary of terms

/ Per

$ Dollars

% Percentage

2D Two dimensional

3D Three dimensional

A Ampere(s)

AC Alternating Current

ADT Articulated dump truck

Ag The chemical symbol for the element silver

allochthonous A term applied to the material forming rocks which have been transported to the

site of deposition

anticline A description of folding of rocks which has produced a convex shape

arenaceous A group of detrital sedimentary rocks, typically sandstones, in which the particles

range in size from 0.06 mm to 2 mm

argillaceous A group of detrital sedimentary rocks, typically clays, shales, mudstones and

siltstones, in which the particles range in size from less than 0.06 mm

As The chemical symbol for the element arsenic

ASCu Acid Soluble copper

arsenopyrite A mineral that is made up of arsenic, iron and sulphur

azurite A mineral that is made up of copper, up to 55% copper, with carbonate and water

BCM, bcm Bank Cubic Metres, a measure of volume applied to unbroken rock

bimodal Statistical term for two peaks in a graph of values

black copper

An impure form of copper produced by smelting oxidised copper ores or impure

scrap, usually in a blast furnace. The copper content varies widely, usually in the

range of approximately 60 to 85% by weight

BOCO Bottom of complete oxidation

bornite A mineral made up of copper, up to 63%, copper, iron and sulphur

boudinaged A minor structure arising from tensional forces, resulting in an appearance in cross-

section similar to that of a string of sausages

brecciated Describes rock made up of angularly broken or fractured rock generally indicating a

fault plane

BMWi Bond Mill Work index

°C Temperature measurement in degrees Celsius (also called Centigrade)

carbonates Rocks made up mainly of a metal, commonly calcium or magnesium or copper, zinc

and lead and carbon dioxide

carrollite A rare mineral that is made up of cobalt, copper and sulphur

CCD Counter Current Decantation

cell A term applied to the three dimensional volume used in the mathematical

modelling by computer techniques of ore bodies

chalcocite A mineral that is made up of copper, up to 80% copper and sulphur

chalcopyrite A mineral that is made up of copper, up to 35% copper, iron and sulphur

chrysocolla A mineral that is made up of copper, up to 36% copper, silica and water

clastic

Rocks formed from fragments of pre-existing rocks which have been produced by

the processes of weathering and erosion, and in general transported to a point of

deposition

cm Centimetre

CMN Calcaire a Minerais Noirs (limestone and dolomite with black oxides)

Co The chemical symbol for the element cobalt

conglomerate A sedimentary rock made up of various size particles from small pebbles to large

boulders and rounded other rock fragments cemented together


P a g e | 25

Cu The chemical symbol for the element copper

CuOx copper in the oxide form, generally soluble in dilute sulphuric acid

cuprous copper in ionic state of one missing electron

cut-off

The minimum concentration (grade) of the valuable component in a mass of rock

that will produce sufficient revenue to pay for the cost of mining, processing and

selling it

DC Direct Current

DCF Discounted Cash Flow

Datamine A proprietary computer program developed to model, view, report and analyse

geological and mining data

diagenetic Pertaining to the processes affecting a sediment while it is at or near the Earth’s

surface, i.e., at low temperature and pressure

dilution A term used to describe the waste or non economic materials included when

mining ore

disseminated Ore carrying fine particles, usually sulphides scattered throughout the rock

dolomite A mineral containing calcium, magnesium and carbonate

domain

A term used mainly in mineral resource estimation or geotechnical investigations to

describe regions of a geological model with similar physical or chemical

characteristics

DRC Democratic Republic of Congo

DStrat Dolomies Stratifies (stratified dolomite)

DTD Direct tailings disposal

DTM Digital Terrain Model

Dwi Drop Weight index

E Easting coordinate

EAF Electric Arc Furnace – a smelting facility

Écaille

A French term meaning ‘fragment’, used to describe the large blocks of prospective

Mines Series stratigraphy that appear to ‘float’ in a mega-breccia-type

arrangement

EGL Effective Grinding Length

EIA Environmental Impact Assessment

EMP Environmental Management Plan

EW Electrowinning

FC Congolese Francs

ferric Iron in an ionic state of three missing electrons

fluvial A geological process in, or pertaining to, rivers

fluvio A description applied to moving material by streams of water

flotation

A widely used process to concentrate valuable minerals after mining that treats

finely ground rock in a water based pulp with chemicals that allow them to float to

the surface where they are recovered in preference to waste or gangue minerals

which sink

framboidal Akin to the skin of a strawberry or raspberry

g Gram

GAC Gangue acid consumption

Gécamines La Générale des Carrierés et des Mines, Parastatal copper Mining Company of the

DRC

geostatistics A mathematical method based on geological spatial knowledge of grade

distributions used to estimate mineralisation grades

GRAT Grey Roches Argilo-Talcqueuse (a dolomitic and talcose argillaceous rock)

GST Goods and Services Tax

ha Hectares


P a g e | 26

HAZOP Hazard and Operability Study

HDPE High Density Polyethylene

HG High Grade

HLS Heavy Liquid Separation

HMS Heavy Media Separation. A process that uses high density fluids to separate

valuable minerals from waste or gangue by exploiting differences in specific gravity

HQ3 Diamond drill core with a diameter of 63.5 mm

hrs Hours

HT High tension

HV High voltage

ICP Inductively Coupled Plasma Mass Spectrometry

ICWi Impact Crushing Work index

ID2/IDS

Inverse Distance Squared (method of estimating grades by mathematically

weighting samples based on their distance away from the estimation point)

IT Information technology

JORC

An acronym for Joint Ore Reserve Committee, an Australian committee formed by

the Australian Stock Exchange and Australasian Institute of Mining and Metallurgy,

the purpose of which is to set the regulatory enforceable standards for the Code of

Practice for the reporting of Mineral Resources and Ore Reserves

kg Kilogram

kL Kilolitre

km Kilometre

kt Kilotonne

kV Kilovolt

kW Kilowatt

kWh Kilowatt hour

kriging

A geostatistical method (named after the South African, D. G. Krige) of estimating

the unknown grade of resource blocks from the grades of samples, taking

cognizance of the sample distribution

kurtosis Statistical term for peaked graph shape (peakedness)

L, l Litres

L/sec, L/s, l/sec, l/s Litres per second

lacustrine Sediment deposition in lakes

lb Pounds

LIDAR Light Detection and Ranging – a remote sensing system used to collect topographic

data

LOB Lower Orebody

Log Natural logarithm to the base 10

LOM Life of Mine

LV Low voltage

m Metre

mm Millimetre

m% Metre percentage (obtained by multiplying metres by % of assay value)

m3 Cubic metre

Ma Mega annum (Million years)

malachite A mineral containing copper, up to 57% Cu, carbonate and water

mamsl Metres above mean sea level

massive A term used to describe a large occurrence of a pure mineral species, often with no

structure

MAX Maximum

mbgl Metres below ground level


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mbs Metres below surface

MCC Motor Control Centre

MCK Mining Company of Katanga

Mg Milligrams

MIR Milling in raffinate

MIN Minimum

MINDIL A Whittle Four-X mine planning software term for mining dilution

mineralisation The presence of minerals of possible economic value or the description of the

process by which the concentration of valuable minerals occurs

mm Millimetre.

ML Millions of litres

MN Magnetic North.

MODFLOW A groundwater modelling program used to assess the impact on the regional

groundwater table of pumping and abstraction, and also contaminant flow

MPa Millions of Pascals

Mt Millions of tonnes

MVa Millions of Volt Amps

MW Millions of Watts

N Northing Coordinate

Neo-Proterozoic The term used in the geological time scale for the period from 545 million years

ago to 1000 million years ago

NI National Instrument

OC Organic Continuous

ore

A natural aggregate of one or more minerals which, at a specified time and place,

may be mined and sold at a profit or from which some part may be profitably

separated

orogeny Greek for ‘mountain generating’ - the process of mountain building. Orogenic

events occur as a result of plate tectonic processes

P80 80% of product passes

Pb The chemical symbol for the element lead

PBC Pinned Bed Clarifier

PDT Phase Disengagement Time

PE Permis d’Exploitation (Exploitation Permit or Licence)

PFDs Process Flow Diagrams

PFS Pre-feasibility Study

P&IDs Piping and Instrumentation Drawings

pH Concentration of hydrogen ion

PLC Programmable Logic Controller

PLS Pregnant Liquor Solution

ppm Parts per million (same as grams per tonne)

pseudomalachite Pseudomalachite or ‘false malachite’ – named because it is visually similar in

appearance to malachite

PVC Polyvinyl chloride

QAQC Quality Assurance and Quality Control

raffinate A liquid stream that remains after the extraction with the immisciable liquid to

remove solutes from the original liquor. From French: raffinere, to refine.

RAT Roches Argilo-Talcqueuse (a dolomitic/talcose argillaceous rock)

RC Reverse circulation (as in drilling)

recovery

A measure in percentage terms of the efficiency of a process, usually metallurgical,

in gathering the valuable minerals. The measure is made against the total amount

of valuable mineral present in the ore


P a g e | 28

reserve (Ore Reserve)

The term for the economic quantities and grade of valuable materials as strictly

applied in compliance with the definition in the Australian JORC Code and in the

Canadian National Instrument (NI) 43-101

resource (Mineral Resource)

The term for the estimate of the quantities and grade of valuable materials but

with no economic considerations as strictly applied in compliance with the

definition in the Australian JORC Code and in the Canadian National Instrument (NI)

43-101

RL Reduced Level (same as elevation coordinate)

Roan Supergroup Describes the stratigraphic succession of sedimentary rocks of Neo-Proterozoic age,

in the Katanga Province of the Democratic Republic of Congo

RMWi Rod Mill Work index

ROM Run-of-Mine (ore)

RSA Republic of South Africa

RSC Roches Silicieuses Cellulaires (siliceous rocks with cavities)

RSF Roches Siliceuses Feuilletees (foliated and silicified dolomitic shales)

S South Coordinate.

s, sec Second

SAG Semi-autogenous Grinding

sandstone A sedimentary rock consisting of sand size grains, generally the mineral quartz,

which is in a consolidated mass

SCADA Supervisory Control and Data Acquisition System

SD Shales Dolomitiques (dolomitic shales)

SEM Scanning Electron Microscopy

SG Specific Gravity

siltstone A sedimentary rock consisting of grains from 0.063 to 0.25 mm, generally the

mineral quartz and clay, which is in a consolidated mass

silica A compound of silicon and oxygen, generally occurring in the form of a mineral

called quartz

SMC SAG mill comminution

SNEL Société Nationale d’Electricité – the provider of electrical power in the DRC

SPLP Simulated Precipitation Leach Procedure

S/S, SS Stainless steel

storativity The volume of water an aquifer releases from or takes into storage per unit surface

area of the aquifer per unit change in head

stratiform

Describes a layered or tabular shaped body of mineralized rock within a

sedimentary rock and implies that the layering of the mineralisation is parallel to

the bedding planes in that sedimentary rock

strings A term used to a digital line drawn within a computer program that outlines or

describes a shape of an object or interpretation

supergene

Pertaining to that part of an ore deposit in which the mineralisation has been

increased as a result of the downward percolation of fluids carrying metal in

solution

SURPAC A proprietary computer program developed to model, view, analyse and report on

geological and mining data

SX Solvent Extraction

SX-EW Solvent Extraction and Electrowinning

t Metric tonne

TCu Total copper

termitaria Termite mounds

TN True North

TOFR Top of fresh rock

tpa Tonnes per annum


P a g e | 29

tpd Tonnes per day

tph Tonnes per hour

transmissivity The volume of water flowing through a defined cross-sectional area of an aquifer

TSF Tailings Storage Facility

TSS Total Suspended Solids

UCS Unconfined Compressive Strength

UTM Universal Transverse Mercator grid

V Volts

VAT Value Added Tax

VESDA Very Early Smoke Detection and Alarm

VSD Variable Speed Drive

%v/v Percent by volume

W Westing Coordinate

Whittle Four-X A mine planning software program used to optimise resource models, based on

economic and mining/processing parameters

WNW West North West

WRD Waste Rock Dump

%w/w Percent by weight

Zn The chemical symbol for the element zinc

μm Microns, micrometers


P a g e | 30

3. RELIANCE ON OTHER EXPERTS

This Technical Report has been prepared and approved under the supervision of Mr Andrew Law,

Director Mining, Optiro Pty Ltd. Mr Andrew Law, who is the principal author of the report, is an

independent Qualified Person as defined in National Instrument 43‐101.

In preparing this report, the Qualified Persons have relied upon and taken responsibility for the

information provided by MWL relating to mining, legal, environmental and financial information as

noted below:

Legal title to the tenements held by MWL in the DRC and MWL’s permits to mine, which is

relevant to Sections 4 and 20 of this report.

Environmental permit and bond information which is relevant to Sections 4 and 20 of this

report.

The nature and validity of any off-take agreements for concentrate held by MWL, which is

relevant to Section 19 of this report.

Financial and cash flow models were provided to Optiro by MWL which is relevant to Section

22 of this report.

Metallurgical balance and current production information leading to the assessed head

grade of the copper-silver concentrate produced from treatment of the mined ore, which is

relevant to Sections 13 and 17 of this report.

Mine design, geotechnical, hydrology, planning, scheduling and costing which is relevant to

Sections 15, 16, 21, and 22 of the report.

The Qualified Persons have made all reasonable inquiries to establish the completeness and

authenticity of the information provided. Drafts of this report were provided to MWL with a request

to identify any material errors or omissions prior to filing. Notwithstanding the reliance of the

Qualified Persons on MWL for the financial and cash flow models, metallurgical balance information

and mine design noted above, the Qualified Persons accept responsibility for all of the

scientific/technical information related to these matters.


P a g e | 31

4. PROPERTY DESCRIPTION AND LOCATION

4.1. DEMOGRAPHICS AND GEOGRAPHIC SETTING

The Democratic Republic of the Congo (DRC) is located in central Africa and straddles the equator.

The DRC has an east-west lateral extent of approximately 1,500 km and extends over a north-south

distance of some 1,800 km. The DRC is Africa’s second largest country, covering an area of

approximately 2.3 million km2 and shares land borders with Angola, Zambia, Rwanda, Tanzania,

Uganda, the Republic of the Congo, Sudan, Burundi and the Central Africa Republic. The capital city

is Kinshasa, which is located in the western portion of the country. The DRC’s main port is Matandi,

approximately 115km from the coast on the Congo River.

The DRC has a population in excess of 75 million of which approximately 50% are aged between 15

and 64 years old. There are over 200 African ethnic groups within the country’s borders, although

the Bantu and Hamitic groups account for approximately 45% of the population. The majority of the

population reside in rural areas with one-third living in urban centres.

Christianity is the dominant religion in the DRC, with approximately half of the population being of

the Roman Catholic faith, with a further 20% Protestant. The remaining population follow the

Kimbanguist (10%), Muslim (10%) and other (10%) faiths.

The national language is French, although Lingala, Kingwana, Kikongo and Tshiluba are widely

spoken.

4.2. PROJECT OWNERSHIP

The Dikulushi mine is governed by the “Dikulushi Mining Convention”, signed on the January 31,

1998 with the Government of the DRC, and ratified by Presidential Decree issued on February 27,

1998.

The Dikulushi Mining Convention is a mining concession granted to AMC. Mawson West

Investments Ltd a wholly owned subsidiary of MWL, holds 90% of the issued capital of AMC, the

remaining 10% is held by the Dikulushi – Kapulo Foundation (NPO).

For the purposes of this report, the Mawson West Limited ownership structure, referred to as MWL,

is used in this report for ease of reference.

4.3. PROPERTY LOCATION

The Project is located within the Katanga Province in the south-eastern DRC, some 400 km north of

Lubumbashi and 50 km north of the regional town of Kilwa. The Project is centred at approximately

S 08° 53’ E 28° 16’, some 25 km west of Lake Mweru near the DRC border with Zambia.

Figure 4.1 shows the property location of MWL’s holding within the DRC, which are effectively two

distinct properties – the Dikulushi property (shown in green in figure 4.1) and the Kapulo property

(shown in blue in figure 4.1). The focus of this report and the projects discussed herein, relate

specifically to the Dikulushi property only.


P a g e | 32

Figure 4.1 Exploration Licences of the Dikulushi copper silver project

4.4. THE PROPERTY TENEMENT AREA

MWL holds title to the Dikulushi mine and surrounding exploration tenements, as governed by the

Dikulushi Mining Convention. Under the Dikulushi Mining Convention the exploration tenements

known as “PR’s” were issued for an initial five year period and are renewable a further three times,

each time for a period of five years; that is a total of 20 years. The Dikulushi PR’s shown in Table 4.1

below and Figure 4.1 above, and these were first granted on the 22 May 2001 and currently have

“renewed” expiry dates April 2016. A further 5 year renewal period is available post this date. MWL

currently holds 18 Exploration Permits and three Exploitation Permits under the Dikulushi Mining

Convention, covering 7,283km².

Under the Dikulushi Mining Convention, MWL is guaranteed sole and exclusive rights for exploitation

for a period totalling 20 years from the date of the issue of the permit.


P a g e | 33

Mining operations at the Dikulushi mine are conducted under an Exploitation Permit PE 606, issued

on 29 December 2003 by Ministerial Decree. The exploitation permit recognised that AMC had

commenced mining operations form 31 January 2002. The PE covers an area of 40.77 km2 over the

Dikulushi mine area (Figure 4.1 and Figure 4.2).

Table 4.1 Mawson West Limited Dikulushi tenement schedule

Tenement Schedule

Project Group Entity Permit No. Area km² Type Granted Expiry

Dikulushi AMC PE606 40.77 Mining 29-Dec-03 30-Jan-22

Dikulushi AMC PR546 283.8 Exploration 23-May-11 22-May-16

Dikulushi AMC PR1693 398.6 Exploration 12-Apr-11 11-Apr-16











Total Area 4,495.1

Figure 4.2 Dikulushi mine infrastructure within the PE 606


P a g e | 34

4.5. ENVIRONMENTAL PERMITS

An EIA for the Dikulushi project was lodged in 2003. In 2009, an EIA for the underground project was

submitted to the DRC Government. Both of these reports were compiled by African Mining

Consultants of Kitwe, Zambia, an environmental company that was licensed to work and report in

the DRC. An EIA was lodged for the cutback project, prepared by EMIS sprl, a DRC environmental

company licences to work and report in the DRC. All three environmental reports received DRC

Government approval. A revised EIA, extending underground mining beyond 2013, has being

submitted to Government.

Each EIA includes commitments relating to mine decommissioning. Annual reporting of

environmental issues and measurements to relevant government bodies is a condition of the

operating license and EMP.

MWL have lodged $1.19M as an Environment Bond. The financial guarantee is a contribution

towards an estimate of the total costs of closure, rehabilitation and re-vegetation of the Dikulushi

mine. The development of the financial guarantee is conducted in compliance with:

Articles 410 of the Mining Regulations

Articles 124 and 125 of Appendix XI of the DRC Mining Regulations 2003; and

Appendix II of the Mining Regulations 2003 Regular environmental audits are carried to determine the mine’s compliance with its Environmental

Management Plan.

An environmental monitoring database is maintained at the mine, comprising the following:

wet/dry, min/max temperatures

rainfall

dust exposure

noise levels

ground and surface water quality

groundwater levels

Tailings Dam piezometer water levels

light levels.

A study into the acid rock drainage potential of the process plant tailings was conducted in 2005 and

they were classified as low risk. Ongoing testwork and monitoring continues to support this

conclusion.


P a g e | 35

Figure 4.3 Dikulushi mine site aerial view


P a g e | 36

5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE AND PHYSIOGRAPHY

5.1. ACCESS

Access to the Dikulushi Mine is by sealed road from Lubumbashi to Kasenga, along the Luapula River

by boat to Kilwa and then approximately 54 km by gravel road from Kilwa to Dikulushi. The total

travelling distance is approximately 500 km. The closest international airport is at Lubumbashi,

approximately 450 km to the south. A gravel airstrip is located at the Dikulushi mine and charter

flights using a G1 plane (as shown in Figure 5.3) from Lubumbashi can land directly at site. Supplies

for the project are typically trucked on sealed roads from South Africa via Botswana to Nchelenge

port on the Zambian side of Lake Mweru. Supplies are then transferred from Nchelenge to Kilwa on

the Congo side of Lake Mweru on a 340 t capacity barge (Figure 5.1) owned by AMC; the water

journey takes 5 hours. Access from Kilwa port to the mine is via a 54 km gravel road and takes

approximately 1 hour by light vehicle.

Figure 5.1 Barge on Lake Mweru

5.2. SITE TOPOGRAPHY, ELEVATION AND VEGETATION

The Dikulushi deposit is located on a plateau approximately 1000 m above sea level. The area

surrounding the Dikulushi site is almost entirely covered with woodland and forest, with some

swamps or wetland areas. The plateau rises into the Kundelungu ranges 60 km to the west of

Dikulushi and forms an escarpment 25 km to the east along the fault-bounded edge of Lake Mweru.

A minor ephemeral stream is located near the Dikulushi mine site. The Luapula River is the main

drainage into Lake Mweru and both form the international boundary between Zambia and the DRC.


P a g e | 37

5.3. CLIMATE, PHYSIOGRAPHY, LOCAL RESOURCES AND

INFRASTRUCTURE

The average annual rainfall, as indicated by mission records, is 1,260 mm, with a range of 800 mm to

2,200mm. An Oregon Scientific weather station was installed at Dikulushi in 2006. A composite

graph of the weather data collected at Dikulushi over the past 5 years is shown in Figure 5.2. The

wet season begins towards the end of October and finishes at the end of April, with 90% of the

annual rainfall occurring during this period. The average minimum recorded temperature is 15°C

and the average maximum temp is 29°C during the year.

Figure 5.2 Composite Temperature and Rainfall Data recorded over the last 5 years.

The wet season generally has minimal effect on mining or processing operations at Dikulushi.

5.4. SURFACE RIGHTS

The Dikulushi mine is based on Exploitation Licence (PE606) granted on 29 December 2003. The

lease is valid for 20 years and can be renewed for up to a further 20 years.

There are no competing mining rights (for example, small artisanal mining licenses) in the project

area.

5.5. SITE INFRASTRUCTURE

The development of the Dikulushi mine has required development of seven major locations:

1. the treatment plant area, which includes the mine administration building

2. the mine services area, including workshops, fuel farm and powerhouse

3. the explosives storage area

0

20

40

60

80

100

120

140

160

180

200

0

5

10

15

20

25

30

35

Jan Feb Mar Apr May Jun Jul Aug Sep Oct Nov Dec

Rai

nfa

ll, m

m

Te

mp

era

ture

, °C

Rainfall

Temperature (High)

Temperature (Low)


P a g e | 38

4. the staff village

5. the airstrip

6. the process water dam

7. tailings storage facility.

These items of infrastructure are depicted in Figure 4.2 and Figure 4.3. This infrastructure was in

place for the previous operations under Anvil and has been used for the open pit cut back operations

with it being well established and maintained, as well as being of sufficient size for the current and

future underground requirements.

Figure 5.3 Dikulushi airstrip and the G1 Charter plane provides safe staff transportation to and from site

5.5.1. WATER SUPPLY

Mine water is sourced from a raw water dam located adjacent to the Tailings Dam. Supernatant

tailings water is reclaimed via penstock arrangements for use in the processing plant. A water

supply flowchart and site wide water balance is provided in section 18 of this report.

Potable water is supplied from various bores on the property which are tested regularly.

5.5.2. POWER SUPPLY

The project is located in a remote area where there is no electrical utility grid. The mine power is

supplied by diesel generators. There is sufficient back-up capacity.

The existing power station at Dikulushi comprises the following generators: 4 x 1.2 MW FG Wilson (

being new units and installed during the 3rd quarter 2013 ) , 1 x 2.0 MW Caterpillar, 1 x 1.6 MW

Caterpillar, 1 x 0.8 MW Mirrlees for a current total capacity of 9.2 MW. The current power demand

for the plant and infrastructure is in the order of 2.0 MW. The 2.0 MW Caterpillar, 1.6 MW

Caterpillar currently require major overhauls which will be completed during 2014. The 1 x 0.8 MW

Mirrlees will be decommissioned during the 4th quarter of 2013. The new FG Wilson generating sets

were installed to supply power to the operations as well as dewatering of the underground and


P a g e | 39

normal underground operations. The 2.0 MW Caterpillar and the 1.6 MW Caterpillar will be used as

backup standby power.

MWL recognises that a consistent reliable fuel supply is crucial to the success of the Dikulushi

operation. The operation currently uses approximately 450,000l of diesel per month. This fuel is

supplied by three DRC based companies, two receive supplies from the port of Beira and the other

receives supplies from the port of Dar es Salaam. MWL has contacted a further supplier from Dar es

Salaam whom would be able to supply fuel to Dikulushi. MWL is regularly speaking to suppliers to

guarantee no interruptions in fuel supply. MWL believes that it has mitigated the risk of fuel supply

by having a number of suppliers whom source fuel from different ports and transport routes.

5.5.3. MINE PERSONNEL

As at June 2013, the Dikulushi mine employed 515 people, of which 39 were expatriates. The

requirements for the underground operations and other associated activities will require a total

workforce of 500-550 employees. Contractors will be used as required. This a change from the open

pit operations where the workforce was mainly contractor supplied.

5.5.4. TAILINGS STORAGE FACILITY

There are currently three tailing storage facilities (TSF) on site. The initial TSF designed for HMS

tailings, dormant since 2004, has had a section of the coarse portion reclaimed and retreated in

early start up operations by MWL. The second TSF is dormant whilst the third is in use to

accommodate the tailings resulting from the treatment of the current open pit operations. The third

TSF has been reviewed for extended use beyond its current life. This will be raised to accommodate

tailings resulting from the final open pit cut back mining operations and the planned underground

operations.

More detail on the TSF is covered in Section 17.

5.5.5. ADMINISTRATION AND PLANT SITE BUILDINGS

The infrastructure on site includes administration offices (Figure 5.4), a warehouse, mining

equipment and maintenance workshops, mechanical workshops and a service area with access pit

for inspection and repair of vehicles.

There is a fully equipped clinic on site (Figure 5.6) and a hospital at Kilwa, approximately 50 km from

the mine. An assay laboratory on site facilitates metallurgical, exploration and grade control

sampling assaying requirements.


P a g e | 40

Figure 5.4 Dikulushi Administration Centre

5.5.6. ACCOMMODATION

A staff village has been constructed 1.8 km from the process plant. A mess hall, fully equipped

kitchen, food storage and laundry facilities serve all employees. Recreational facilities are also

available to employees. Figure 5.5

Figure 5.5 Dikulushi Camp Site


P a g e | 41

5.5.7. COMMUNICATIONS

Mobile phone coverage is available through a dedicated mast located on top of the waste dump.

There are satellite systems for data transmission and VOIP telephone coverage. There is a base

station radio system, along with vehicle and hand-held radios. Figure 5.6

Figure 5.6 Dikulushi Clinic and communications centre

5.5.8. MOBILE EQUIPMENT

Sufficient mobile equipment for the efficient running of the operations is in place, comprising light

vehicles (including an ambulance), light trucks, forklifts, buses and generators. Figure 5.7


P a g e | 42

Figure 5.7 Dikulushi Workshop

5.5.9. SECURITY

Security is provided by a contractor. Appropriate secure facilities are provided for the storage of fuel

and explosives. Figure 5.8


P a g e | 43

Figure 5.8 Dikulushi Store and Fuel Farm


P a g e | 44

6. HISTORY

The development of exploration and mining at Dikulushi and surrounds can be broken down into the

following periods:

• Early History – Copper mineralisation was first reported in the early 20th Century by

Simkat. Other assessments were made in the 1950s and by the French Bureau de

Recherches Geologiques et Minieres (BRGM) during the 1970s.

• Recent History – Anvil, from 1996 to April 2008.

• Current – Mawson West Limited from June 2010 to present.

The history of the project is summarised in Table 6.1.

Table 6.1 Historical work summary at the Dikulushi Project

Year Supervision Work Completed

1900s Belgian explorers Rock chips

1974-1981 BRGM 48 Diamond drillholes

1996-1997 Anvil Dikulushi Mining Convention signed and some drilling

1998-2002

2002-2006

2006-2008

Anvil

Anvil

Anvil

Modified convention signed, more drilling and metallurgical testwork, feasibility study

Mining of open cut and regional soils and termite sampling and drilling

Underground development at Dikulushi and further drilling of targets defined from above

2008 Anvil Anvil closes down Dikulushi Mine.

2010

onwards

MWL

Restarted plant on LG stockpiles and then proceeded to process ore from the pit cut back

operation until July 2013 when the pit concluded operations due to reaching its design

limits.

6.1. BELGIAN EXPLORATION

Copper mineralisation in the area was initially evaluated by Belgian explorers (Simikat) from 1910

until 1923.

BRGM purchased an interest in the Dikulushi Deposits during the early 1970s and completed adit

sampling, diamond drilling, metallurgical testwork, soil geochemistry and geophysics. The projects

lay dormant until Anvil pegged the ground in the late 1990s and subsequently signed the JV

agreement with Mawson West.

6.2. ANVIL MINING LTD

An open pit mine was commissioned at Dikulushi in October 2002 by Anvil, with run-of-mine ore

delivered to an on-site heavy media separation (HMS) concentrator at the rate of 250,000 tonnes

per year. The copper-silver concentrate was subsequently transported by barge across nearby Lake

Mweru into Zambia and then by road to smelters in South Africa and Namibia.

During the first 15 months of operation, the geology within the open pit was extensively mapped

and, with results of the drilling, resulted in a re-interpretation of the mineralised envelope at

Dikulushi.

The DevMin consulting group was approached by Anvil to undertake the Open Pit Mine Plan study to

estimate the remaining open pit reserves and prepare an open pit life-of-mine schedule for the

Dikulushi Mine. This study was initiated in August 2003 when a preliminary pit optimisation was


P a g e | 45

undertaken using Anvil’s current in-house resource model. The initial study was largely an update of

previous work carried out by DevMin and showed that with current economic parameters, the

feasible depth for open pit mining at Dikulushi could potentially extend down to 150 m below

surface. An earlier work phase, carried out in 2002 prior to the mine being commissioned, had

stopped the pit at 130 m depth with remaining mineralisation proposed to be exploited by an

underground mine.

Further metallurgical work was carried out by Anvil in 2003, resulting in the installation of a ball mill

and flotation plant in mid-2004 increasing the throughput to 350,000 tonnes per year. A second ball

mill was installed in June 2005 with a further increase in throughput to 520,000 tonnes. This is the

current capacity of the treatment plant and the mining fleet was increased to provide the

appropriate plant feed tonnage.

With the completion of the planned open cut operations in 2006, the focus moved to developing an

underground operation. The ROM stockpile at the end of 2006 was considerable and it was

expected that this would see the mine through to the commencement of production from the

underground in early 2008.

Upon commencement of underground mining, and due to time constraints, Anvil adopted a sub-

level caving method for the underground. However with the realisation that such a method was

inappropriate for the type of orebody geometry at Dikulushi, the decision was made to move to an

Avoca method of mining. This necessitated a hiatus in ROM feed to the plant in April 2008 due to

the intensive nature of development required for a bottom-up mining method.

By the end of 2008, with the World financial markets in turmoil and the subsequent plunge in the

copper price, the decision was made to place the Dikulushi Mine on care and maintenance.

6.3. MAWSON WEST

MWL acquired the Dikulushi project from Anvil in April 2010. Plant refurbishment was started

immediately and completed in July 2010 at which point MWL started processing the LG stockpile

which continued into 2012.

MWL has continued with the cut back of the open pit until July 2013, where practicable completion

of the pit was reached. Ore is stock piled on the ROM pad to continue the milling operations until

late 2013. The next phase is the re-establishment of the underground operations below the current

open pit (825 mRL). Previously, the underground was developed by Anvil down to the 750 mRL.

6.4. RESOURCE HISTORY

During the BRGM tenure of the deposit (1974 – 1981) a resource of 1.65 Mt at a copper grade of

10.46% and a silver grade of 310g/t to a vertical depth of 220 m was estimated for the Dikulushi

deposit. Anvil published a Mineral Resource estimate for the Dikulushi deposit in December 2006.

The estimate was completed by FinOre Mining Consultants (FinOre) in July 2006 and used 3D

wireframe volumes to define the mineralisation. A 0.5% copper mineralisation cut-off was used to

guide the wireframe volume. Estimates were completed using Datamine software and the resulting

Mineral Resource is detailed in Table 6.2 below.


P a g e | 46

Table 6.2 Mineral Resource estimate as completed by FinOre in July 2006 and published in December 2006; a cut-off grade of 1.5% copper was used

Category Tonnes Cu(%) Cu Metal (Tonnes) Ag g/t

Ag Metal Mozs

Measured 410,000 9.10 37,300 288 3.8

Indicated 650,000 7.90 51,100 184 3.84

Measured and Indicated 1,060,000 8.30 88,400 224 7.64

Inferred 1,380,000 5.80 80,600 141 6.26

There are no other historically published Mineral Resources available for Kazumbula or any of the

surrounding exploration targets.

6.5. PRODUCTION HISTORY

Anvil mined the Dikulushi deposit between 2002 and 2008. Production statistics are presented in

Table 6.3 below.

Table 6.3 Historical Anvil production for the Dikulushi mine

Year Tonnes processed Grade Cu% Grade Ag g/t

Average recovery

(%)

Cu produced (tonnes)

Ag produced Mozs

2002 (HMS) 36,010 7.26 144 66.7 1340 0.08 2003 (HMS) 273,500 7.68 195 66.1 13,613 1.16

2004 (HMS/Float) 245,000 6.39 177 69 10,840 0.89 2005 (Float) 410,000 5.07 149 86 16,900 1.63

2006/07 (Float) 1,059,950 5.12 151 81 46,507 4.45 2008 (No Dec) 471,590 3.16 84 75 11,177 0.97

From 2010 MWL recommenced processing operations of low grade stockpiles whilst the cut back of

the open pit commenced mining in Jan 2012 and ceased July 2013, with ore processing due to be

completed late 2013. Production statistics for this period are presented in Table 6.4 below.

Table 6.4 Recent MWL production for the Dikulushi mine

Year Tonnes processed Grade Cu% Grade Ag g/t

Average recovery

(%)

Cu produced (tonnes)

Ag produced Mozs

2010/11 467,958 1.46 35.31 64.0 4251 0.36 2011/12 347,863 1.80 34.2 63.0 3,948 0.27

2012/13* 355,043 5.21 141.5 91.6 16,925 1.46

*Note: 2013 Production is to the end of July 2013 (13months).

[Note: These figures are available in table 17.2]


P a g e | 47

7. GEOLOGICAL SETTING AND MINERALISATION

7.1. REGIONAL SETTING

The Dikulushi and Kazumbula copper-silver deposits are located west of Lake Mweru (08°53’37”S;

28°16’21”E), near the eastern margin of the Katanga sedimentary basin, in an area known as the

Kundelungu Plateau. Deformation of the Katanga Supergroup sedimentary rocks is mild, and

comprises open upright folds developed in association with major north-northwest and subordinate

north-northeast-trending faults. An angular unconformity separates the undeformed, uppermost

subdivision of the Kundelungu Group (Plateau Series) from the remainder. Regionally, copper

mineralisation is known from at least three stratigraphic levels in the Kundelungu Group, but all

occur beneath this unconformity. These deposits are hosted by the Kalule Formation, and ore is best

developed in red sandstones and shales of the Mongwe member, above a conspicuous

reduction-oxidation (redox) and pH boundary with the grey carbonates of the underlying Kiaka

member. Mineralisation at Dikulushi was strongly fault-controlled (Haest et al., 2007), but across

the district many copper occurrences, including the nearby Kazumbula deposit, are associated with

the same stratigraphic position. Stratabound mineralisation is known from higher (Mwitapile; Sonta

member, Kiubo Formation) and lower in the stratigraphy (Lufukwe; Monwesi Formation; El Desouky

et al., 2007). In the greater Dikulushi district, prospective parts of the stratigraphy are exposed in

areas of low terrane from which the upper, un-mineralised sequences have been removed by

erosion. Mineralisation occurred during the waning stages of the Lufilian Orogeny (560 Ma) as

compression and ductile deformation gave way to extension and brittle deformation. Several

hundreds of millions of years later the Dikulushi deposit was chemically reworked and upgraded by

circulating groundwaters (Haest, 2009; Haest et al., 2010). Figure 7.1 depicts the regional geology of

the Dikulushi district and Figure 7.2 summarises the regional stratigraphic sequence.


P a g e | 48

Figure 7.1 Regional Geology of Mawson’s convention area in the DRC


P a g e | 49

Figure 7.2 Stratigraphy of Dikulushi region with known styles of mineralisation

Kapenga Schists

Super Group m-Approx

thickness

Up

per K

un

delu

ng

u -

Ks

50Kyafwama -

Kombo Sandstone

Argillaceous red sandstones

interbedded with argillates.

Minor Sandstone

Red sandstone

Sonta Sandstone

Kalule

Ks - 1

Sandstone - argillaceous and

weakly calcareous<50

Mongwa Schists <100

Fault controlled Cu-Ag

mineralisation at the

dolomite/ sandstone contact

Kanianga

Sandstone

Ks - 1.1

Basal dolomite (BD)

Interbedded white/pink

carbonate muds and arenites<50

Low

er

Ku

nd

elu

ngu

Ki

Roan RMwashya

R-4R - 4.2

Monwesi

Ki -0 1.1Likasi Ki-1 Diamictite 500

Fluvio glacial sandstones <50mStratiform copper Kinkumbi,

Lufukwe anticline

Diamictite <100

Zn-Pb mineralization at

contact with diamictite

Dolomitic shales and

sandstones?

Le Grand

Conglomerate

Monwesi

Le Petit

Conglomerate

Interbedded red argillaceous

sandstone and red argillates150Lufila schists

Kiubo Sandstone Red sandstone 50

Group Formation Unit

Kilunga Lupili

Arkose

Lithology Mineralization

Disseminated chalcocite and

malachite at base of Sonta

Sandstone

100Red sandstone and quartzite

Interbedded red argillaceous

sandstone and red argillates350

Pink Arkose <200 Plateaux

Ks-3

Kiubo

Ks - 2Ks - 2.2

Sampwe schistsInterbedded red argillaceous

sandstone and red argillates300

Ks - 1.3

Lubudi

Dolomites

Interbedded pink, cross

bedded dol-arenites,with olitic

caps

<50

Kiaka

Carbonates Ks -

1.2Pink intramicrite flakestones <50

4

Lusele Pink

Dolomites


P a g e | 50

7.2. PROJECT GEOLOGY

The Dikulushi and Kazumbula deposits are principally fault-hosted lodes of massive to semi-massive

copper sulphides with minor disseminated and stockwork mineralisation in the wallrocks. The

deposits are mostly hosted by red siltstones, with lesser ore hosted by underlying grey carbonate

rocks and fragmental rocks that mark the contact between the two. The mineralised Dikulushi Fault

trends east-northeast and breaches the northeastern nose of a north-northeast plunging elongate

anticline. This orebody therefore is suborthogonal to stratigraphy. Prior to mining at Dikulushi, it

was approximately 400 m long, 10 metres wide and extended from surface to a depth of at least 450

metres (Figure 7.3).

Figure 7.3 Local geology of the Dikulushi open pit


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The Kazumbula orebody (Figure 7.4) comprises a fault-controlled lode of disseminated and minor

fracture hosted copper sulphides and oxides. The orebody is hosted by siltstones and intercalated

granular lithic sandstones. These have pervasive grey-green hydrothermal illite alteration when they

occur in the immediate wallrocks to the orebody, but are red-brown elsewhere. The mineralised

Kazumbula Fault trends east-northeast, but mineralisation is located at the intersection of the fault

with other structures trending north-northeast and north-northwest. These structures are

considered to have been active in concert as parts of a conjugate strike-slip shear array. Gently

folded stratigraphy on the eastern limb of the Kabangu Antiform abuts the fault plane. The orebody

is sub-orthogonal to stratigraphy and appears to be restricted to preferred brittle and permeable

stratigraphy. As it is presently known, the deposit is approximately 180 m long, 12 m wide and

extends from surface to a depth of approximately 80 m (Zukowski et al., 2010) as confirmed by

drillhole intercepts.

Figure 7.4 A typical vertical cross section through the Kazambula deposit, highlighting key geology associated with mineralisation


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8. DEPOSIT TYPES

The Dikulushi and Kazumbula copper deposits are interpreted to be hypogene, fault-controlled

deposits, containing semi-massive and disseminated chalcopyrite-bornite mineralisation with

azurite, malachite and cuprite developed in the supergene zone. The mineralisation is litho-

structurally controlled and is hosted in shales and sandstones of the Kundelungu Group. The host

sedimentary rocks show varying degrees of brecciation, with the highest grade zones comprising

semi-massive bornite-chalcopyrite fault fill. The Dikulushi and Kazumbula deposits are not typical of

the stratiform copper deposits which are common within the central African Copperbelt.

Dikulushi ore comprises massive to semi-massive and fracture-disseminated copper sulphide

minerals that filled open spaces and cemented breccias along the fault zone. Early mineralisation

was polymetallic and contained Cu-Fe-Pb-Zn and Ag as chalcopyrite, bornite, galena and sphalerite.

Subsequent remobilisation dissolved most of the Fe, Pb and Zn and led to upgrading of the copper-

silver content of the ore. The central parts of the deposits commonly contain >10% Cu and >200

ppm Ag. As a result of this process, the deposits are now composed largely of the copper sulphide

mineral chalcocite. Silver occurs as atomic-level substitutions and as very fine grained inclusions in

chalcocite grains (Dewaele et al., 2006; Haest et al., 2009; Haest et al., 2010). Gangue minerals

typically associated with the mineralisation at Dikulushi are quartz, feldspar, pyrite and clays. Figure

7.3 depicts the mapped geology of the Dikulushi open pit.


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9. EXPLORATION

The Dikulushi areas copper deposits have been explored by three main groups over the last 40 years.

The three main periods of exploration are:

Simkat - BRGM (1910-1981)

Anvil (1997 to 2008)

MWL (2010 - current)

9.1. BRGM

The Dikulushi deposit was discovered by Simkat around 1910 but no records of work completed are

available. In the period from 1910 to 1936 the deposit was investigated by Simkat and Societe

Miniere du Lac Moero. BRGM then appraised the deposit from 1974 to 1981, during which time 48

diamond holes for 5,223 m were drilled and a resource of 1.65 Mt @ 10.46% Cu and 310g/t Ag to a

vertical depth of 220 m was estimated. In 1988 the BRGM withdrew from Zaire (previous name for

the DRC) and the deposit was left dormant until Anvil took control in 1996.

9.2. ANVIL MINING LTD EXPLORATION

In 1996 Anvil submitted a Mining Convention application to the Zaire government for the Dikulushi

deposit. The documents were signed in January 1997 and this document was then re-negotiated

with the new Kabila government in 1998, with a presidential decree issued the month after signing

ratifying the convention.

During the latter half of 1997, Anvil completed 26 reverse circulation drillholes and 18 diamond

drillholes which, together with the BRGM drilling, formed the basis for a pre-feasibility study

completed by Signet Engineering of Perth, Western Australia. Anvil subsequently carried out

additional drilling in 2000, 2003, 2004, 2005-6, 2007 and 2008, which is detailed in Table 9.1.

Table 9.1 Historical drilling summary for the Dikulushi copper silver project

Company Period Type No. Holes Metres Sequence

BRGM 1974-1981 DDH 48 5226 DIK1-47

Anvil

1997 DDH 18 2115 DDH1-14, DR15, 19, 20, 38

RC 26 2305 DRC15-40

2000 RC 22 786 DRC043-064

2003 DDH 4 885 DDH16-19

RC 21 1768 DRC065-085

2004 RC/DDH 14 414/3811 DDH020-035 (no 032)

2005-2006 DDH 14 5779 DDD38-052

2007 DDH 9

78

2061

4130.9

DDD053-061

UGD001 to UGD97

2008 DDH 2 1251 SUR001&005


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9.3. MWL EXPLORATION

At Dikulushi MWL has completed some geotechnical drilling for pit cut back and underground

development studies, but this has not been sampled.

MWL has also carried out drilling at the Kazumbula deposit. Details of this are provided in Section

10.8.


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10. DRILLING

10.1. INTRODUCTION

All the drilling prior to 2010 was completed by Anvil. This section describes the various Anvil

programmes and the programmes carried out by MWL in 2010 which were relevant to the

Kazumbula deposit.

10.2. ANVIL PROGRAMME 1997

In 1997 Anvil carried out an exploration programme at Dikulushi which aimed at confirming the

BRGM results and allowing estimation of Mineral Resources to the standard of the Australian JORC

Code. A total of 40 drillholes were completed using two drill rigs between September and October

1997. All holes were oriented at 60° towards an azimuth of 340° (grid north), and all holes were in

the area previously drilled by the BRGM. These resources and reserves were subsequently reviewed

in the light of the requirements of NI43-101, since Anvil was a listed company in Canada.

A total of 18 HQ and NQ diamond drill (DDH) holes (DDH1-14 and DDH15, 19, 20, 38), including 4

with reverse circulation (RC) pre-collars, were completed in the 1997 programme. Mineralisation

was intersected in 17 of the 18 holes. Downhole camera surveys were completed at least every 12

metres in the DDH holes and the results showed the DDH remained essentially straight with a

maximum deflection of 2° in declination and 4° in azimuth. A core orientation spear was also used

after every core run (usually 3 metres). Upon recovering the core it was oriented when possible, and

was then logged for geotechnical defects, core recovery and geology. Core recovery averaged about

90%, except for minor soil, sandy and cavernous zones. Recoveries in mineralised zones are

reported to have been about 90%.

A total of 22 RC drillholes (DRC16-18, 21-37, 39, and 40) were completed during the 1997

programme using a booster compressor, which was essential due to large water inflows and broken

ground. Mineralisation was intersected in 19 of the 22 RC holes. None of the wholly RC drilled holes

was downhole surveyed. Four RC pre-collared holes were surveyed throughout their length but due

to the presence of steel casing, only the cored section returned valid azimuth readings. The dip

variations in these holes were minor; however, the cored sections showed consistent anticlockwise

azimuth rotations of between 9 and 18°. Consequently, an average azimuth correction factor of 3.2°

anti-clockwise deviation per 20 metre downhole has been entered into the RC drillhole database to

compensate for this interpreted drillhole rotation.

Of the holes drilled during the 1997 programme, 11 DDH holes and 6 RC drillholes were specifically

collared to twin earlier BRGM DDH holes.

Anvil also cleaned out and re-sampled six trenches and four test pits that were originally dug and

sampled by the BRGM. A total of 90 channel samples and 18 rock-chip samples (pits) were taken

from 191 metres of trenching as well as 18 rock chip samples from pits. Orientation soil and stream

sampling and other regional reconnaissance sampling and exploration were also completed during

the 1997 programme.


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10.3. ANVIL PROGRAMMES 2002 & 2003

The drilling programmes carried out in 2002 and 2003 were partly undertaken in response to the

observation in Munro (1998) that as the “upper 30 metres of the deposit is poorly defined, infill

pattern diamond drilling is required”, along with the need to clarify the extent of the mineralisation

and to sterilise certain areas prior to the commencement of mining.

The diamond core drilling procedures used during the 2002 and 2003 programs were compatible

with those of the 1997 Anvil drilling program. A Stanley Drilling Longyear 38 Rig was used in the

2002-03 programmes. To avoid the need to reduce core size at depth to NQ, the first 60 m was

drilled using PQ diameter, followed by HQ. This practice ensures that a good volume of mineralised

intercept core is available for both assay and archive.

Core orientation procedures were routinely conducted, although drilling conditions at times limited

successful achievement of orientation results. Core recoveries across the mineralised horizons

typically exceeded 95%.

Regrettably, all archived core from the 1997 and 2002 campaigns was rendered useless after core

trays were overturned by army units during the latter stages of the civil war. Fortunately,

photographs of this core, taken for geotechnical purposes, are still available.


The objective of the 2004 programme was to extend the resource down to 300 m below surface and

to provide data for a pre-feasibility study on an underground mine. 14 holes were pre-collared with

RC (414 m) and drilled to a maximum ore intercept depth of 280 m below surface (3,811 m).

The diamond core drilling procedures used during the 2004 programme were largely compatible

with those of the previous two programmes, although core recoveries were not as good due to

technical problems with the rig. Stanley Drilling was the contractor for both the RC and diamond

drilling.

Downhole surveys were carried out at 50 m intervals using an Eastman camera. Core orientation

was attempted using the spear method, but poor ground conditions rendered the data to be of little

practical use.

10.5. ANVIL PROGRAMME 2005/6

The objective of the 2005/6 programme was to increase the confidence in the geological model and

upgrade the resource classification to a depth of 400 m below surface for a possible future

underground mining operation. Diamond drillholes from this programme were identified with the

prefix “DDD”.

The programme was drilled with one of Anvil’s own Boart Longyear LF90 rigs, managed by Wallis

Drilling. The holes were all drilled to 45 m with HQ (to which depth the holes were cased) and

drilling continued with NQ. Drilling procedures were upgraded following recommendations made by

Arnold (2004b), and included reducing the downhole survey interval (to approximately 30 m) and


P a g e | 57

establishing the daily maintenance of an up-to-date digital database of geological and geotechnical

logs, survey data, and QA/QC data.

Each hole was surveyed every 30-40 m, using a single shot Tropari tool. Drilling procedures were

similar to those of the 2004 programme.


A surface programme was drilled from the pit in 2007 to increase the confidence in the underground

resource. A total of 9 diamond holes were completed for 2061 m to better define the mineralisation

beneath the current pit at the time. The drillholes were all diamond HQ near surface and reduced to

NQ at depth for the deeper holes. A total of 78 UGD series underground holes were completed as

grade control diamond holes.


In late 2008 there was the recognition that another surface drilling programme was needed as there

was significant and increasing capital cost being sunk into the Dikulushi Mine. The objective of the

2008 programme was to better define the resource below the 630 mRL to assist with the planned

underground development. The orebody below the 630 mRL was at the bottom of the resource

model at the time.

A decision was made to mobilize a Titan drill rig to site in late October 2008 to commence drilling.

By the time the mine was placed on care and maintenance in early December 2008, only two out of

the 5 proposed holes had been completed. Both holes intersected the mineralisation but were not

sampled.

10.8. MWL PROGRAMME 2010

The Kazumbula deposit was drilled by Anvil during 2008. The Anvil drilling data assisted MWL with

drillhole planning and targeting of the Kazumbula deposit. MWL drilled RC and diamond holes

(Table 10.1) to define the near surface copper mineralisation during August and September 2010.

The drillhole spacing was approximately 15 m along drill lines spaced 20 m apart. Drillholes were

drilled at approximately 60 degrees to the south-southeast to maximise the angle of intersection

with the orebody.

Table 10.1 MWL drilling at Kazumbula

Prospect Type No Holes Metres Samples

Kazumbula RC 17 1676 1676

Kazumbula DDH/tail 10 674.4 674

The RC drilling was completed by Titan Drilling of Lubumbashi, utilizing a truck mounted RC rig. A

supervising geologist was on site at all times during the drilling and industry standard procedures

were followed during the RC drilling programme. The diamond drilling was contracted and

completed by Chantete Emerald, who completed five diamond holes (HQ3) from surface and four

diamond tails from RC pre-collars (HQ3).


P a g e | 58

10.9. SURVEY CONTROL

Drill hole collar locations have been located using a Leica total station by mine surveyors. Collar

surveys were stored with both local mine grid and UTM coordinates. A local Dikulushi Grid was

established and collar locations stored in both local mine grid and UTM co-ordinates. The

relationships between Dikulushi Mine Grid and magnetic (MN) and true north (TN) orientations are

as follows:

• Grid North = MN - 20.2˚

• Grid North = TN - 22.0˚

10.10. DRILLING ORIENTATION

Downhole surveys of drillholes were completed every 30 m to 50 m of advance in order to ensure

that each hole was not deviating too much from the planned dip and azimuth. The surveys were

measured using Eastman downhole electronic single-shot cameras that record the dip and azimuth

and results were then tabulated for each hole as a report, which was checked by the site geologists

for accuracy. The camera has a stated accuracy of 1 degree in dip and azimuth. No magnetic

minerals have been noted in the logging of the drill core at Dikulushi and thus the recorded azimuths

are regarded as reliable.

Holes at both Dikulushi and Kazumbula were generally drilled orthogonal to the mineralisation, and

thus significant true width conversions were not required. The average mineralisation true thickness

is significantly greater than the average (1 m) downhole drilling increment, so distorted intersections

of the mineralisation at Dikulushi and Kazumbula were not obtained.


P a g e | 59

11. SAMPLE PREPARATION, ANALYSIS AND SECURITY

All of the drilling data used in the Dikulushi resource estimate was collected by Anvil. A review of

Anvil’s procedures by Mawson personnel concluded that they were of an acceptable industry

standard. The sampling procedures described below have been used for all diamond core drilled at

the Dikulushi deposits. All sampling and logging data are stored in secure database systems and

have been subject to routine validations during capture and storing.

11.1. DIAMOND CORE SAMPLING

Drill core (HQ and NQ size) is sampled by splitting it in half with a water lubricated diamond blade

saw. MWL has ensured the diamond blade is cleaned frequently using a brick to prevent across

metre contamination (especially in zones of massive sulphide). All drill core was sampled on a metre

basis from the start to end of hole. Minor residual sample lengths (less than one metre) may occur

along mineralisation contacts or at the end of holes. Each metre sample of half drill core is collected

from a consistent side of the drill core tray and placed into a sequentially numbered sample bag.

Sample log books are used to record the drillhole number, sample number and the “from” and “to”

sample depths. MWL has additionally incorporated electronic data capture of the sampling into a

toughbook laptop computer.

Calico/sample bags were tied up and placed into larger labelled plastic bags for transportation.

Submission forms for the laboratory were completed and placed into a small plastic bag within the

large labelled plastic bag. The samples were appropriately packaged for transport to the respective

international or minesite laboratories. Transport documentation and customs clearance were

completed by the company representatives. Laboratory turn-around times varied from a few days

to typically three to four weeks for the international laboratories.

11.1.1. DIAMOND CORE RECOVERY

At the end of a core run, the drillers attach a water hose and pump the HQ/NQ core out of the barrel

into an angle iron ensuring minimal disturbance. The driller records the total depth and core run

length on a core block, also noting any core loss, and places this into the core tray at the end of the

run. The site geologist regularly checks the depths provided by the drillers with the core in the trays

during site visits. Diamond core recovery is good and was noted to be above 95%. Any handling

core breaks are marked with a cross. Once a core tray is full, the tray was labelled with from and to

depths and the hole number. The labelled core trays were moved to the core logging/storage area

at the Dikulushi mine site.

11.1.2. DIAMOND CORE LOGGING

The drill core was washed to remove any residual cuttings. Downhole metre marks were made by

the geologist on the consistent half of the core. Wet core was photographed for the more recent

drillholes, using a digital camera before logging. Labels showing hole number, tray number and from

and to depths were placed in the photo frame for each core tray photograph. Core recovery, RQD,

geology, alteration and mineralisation were logged onto standard paper logs and more recently by

MWL into a toughbook laptop computer using LogChief software. The logs were electronically


P a g e | 60

captured into an Access database at Dikulushi or electronic logs were emailed to Mawson West’s

Perth office and loaded into the central Datashed database.

11.2. RC SAMPLING AND LOGGING

RC drilling by Anvil was generally limited to depths of less than 120 metres below the pre-mining

surface. The portion of the Mineral Resource and Ore Reserve estimates that are the subject of this

report are informed entirely by diamond core drill holes.

11.3. SAMPLE QUALITY

Drill core samples that inform the portion of the Mineral Resource and Ore Reserve estimates that

are the subject of this report are of good quality with no major risks identified for use in the resource

estimate. Open pit blast hole grade control samples and underground sludge hole samples have not

been used to inform the Dikulushi resource and reserve estimates.

With respect to sample security, no sample preparation other than diamond core cutting was carried

out by MWL or Anvil employees.

Reference materials, including chips, core, pulps and residues are retained and stored at the

Dikulushi mine site. Assessment of the data indicates that the assay results are generally consistent

with the logged alteration and mineralisation tenor.

11.4. SAMPLE PREPARATION AND ANALYTICAL PROCEDURES

Preparation and assaying of samples from the Dikulushi Project has been carried out at three

independent laboratories:

Genalysis (RSA) and Genalysis (Western Australia)

ALS-Chemex (RSA) (from Jan 2008)

SGS (Dikulushi).

11.4.1. ANALYSES

Samples sent to Genalysis in Johannesburg, South Africa, were processed and analysed using the

following methods. All samples were weighed, then dried at 110° for 8 hours and then crushed to a

nominal 10 mm crush size in a conventional jaw crusher. The entire sample was then pulverised to a

nominal 85% passing 75µm in an LM-5 mixer-mill. A scoop of the pulverised sample was then

digested by the AX method which was a modified (higher precision) 4 acid digest for base metals.

Analysis technique was by AAS for Copper (0.01% detection limit) and Inductively Coupled Plasma

Mass Spectrometry (ICP-MS) for Ag (1ppm detection limit), As (10ppm detection limit), Co (1ppm

detection limit) and U (0.1ppm detection limit). Results were reported electronically via email and a

hard copy report was mailed to MWL and Anvil staff.

Samples sent to ALS Chemex in Johannesburg, South Africa, were prepared and analysed by the

following procedures. Samples were weighed and then dried for 8 hrs at 110° and then fine crushed

to 2 mm with a 250 g split of the sample taken for pulverising to 85% passing -75µm. The sample

was then digested in a four acid mixture (HF, HNO3, HClO4) and a HCL leach with analysis by AAS


P a g e | 61

(method AA62) for Copper (0.01 to 40% reporting range). Other elements were analysed by the ICP-

AES (optical emission method) with detection limits of Ag (0.5ppm), As (5ppm), Co (1ppm), U

(10ppm) and Co (1ppm).

Samples sent to the SGS laboratory at Dikulushi were prepared and analysed by the following

procedures. Received samples were sorted and dried at 105 degrees for a minimum of 8 hours and

then crushed to a nominal 10 mm crush size using a jaw crusher. Samples were split to 250g and

then pulverised to 90% passing -75µm. The sample was then digested in triple acid digest (A103

method;0.4g, Hydrochloric acid, Nitric acid and Perchloric acid) and finally analysed by AAS machine

with low detection limit (DL) for Copper of 0.001% and an upper detection limit (UL) of 5%, Silver (DL

was 5ppm and UL was 500ppm), Cobalt (DL was 20ppm and UL was 5%), Lead (DL was 10ppm and UL

2.5%), Arsenic (DL was 0.01% and UL was 5%), Zinc (DL was 10ppm and UL was 5%) and Fe (DL was

0.01% and UL was 100%). The sample batches included 2 standards, 2 blanks, 2 repeats and 1

replicate per 43 samples.

11.5. BULK DENSITY DETERMINATIONS

Samples were collected every metre from the massive sulphide zones and from a representative

selection from the transitional and primary zones and un-mineralised zones. Diamond core samples

were prepared by ‘squaring off’ the ends of approximately 10-20 cm billets of half core. A total of

1,294 specific gravity (SG) measurements were made of dried half core to obtain the dry weight at

Kazumbula. The same piece of core was then measured in water on a suspension cage below the

same electronic scale. The conventional formula for SG determination was used, i.e.

SG = Dry Sample Weight / (Dry Sample Weight – Wet Sample Weight)

11.6. SAMPLE QAQC

11.6.1. STANDARDS AND BLANKS

QAQC for exploration drilling samples includes use of standards, blanks and duplicates, together

with internal/laboratory batch control information. Results from submitted standards are shown in

Figure 11.1, Figure 11.2 and Figure 11.3. GBM398-4c (Figure 11.3) is a low copper value standard

and suggests accurate results for low value samples (~0.39% Cu).


P a g e | 62

Figure 11.1 GBM301-7 suggests accurate values around low value samples (~0.55% Cu)

Figure 11.2 The GBM301-8 is a high Cu value standard and suggests accurate results for high value samples (~10% Cu)

Figure 11.3 The GBM398-4c is a low Cu value standard and suggests accurate results for low value samples (~0.39% Cu)

Analytical results from sample blanks, Figure 11.4, suggest that contamination was kept to a

minimum.


P a g e | 63

Figure 11.4 Results for this blank demonstrate that contamination is well contained

Both blank and standard sample results indicate that minor sample mislabelling occurred.

11.6.2. LABORATORY QAQC

The respective laboratories perform internal QAQC checks as per the following description from ALS

Chemex (Johannesburg):

“The Laboratory Information Management System (LIMS) inserts quality control samples (reference

materials, blanks and duplicates) on each analytical run, based on the rack sizes associated with the

method. The rack size is the number of samples, including QC samples, included in a batch. The blank

is inserted at the beginning, standards are inserted at random intervals, and duplicates are analysed

at the end of the batch. Quality control samples are inserted based on the following basis:

Sample Count Methods QAQC Sample Allocation

40 Regular AAS, ICP-AES and ICP-MS methods 2 standards, 1 duplicate, 1 blank

The laboratory staff analyses quality control samples at least at the frequency specified above. If

necessary, laboratory staff may include additional quality control samples above the minimum

specifications.”

Failed batches are automatically repeated until acceptable results are achieved.

11.7. SUMMARY STATEMENT

Sampling of drillhole material and QAQC is comprehensive in its coverage of the mineralisation and

does not favour or misrepresent in-situ mineralisation. Sampling and sub-sampling procedures are

of good standard industry practice and have occurred in a safe and secure manner, with minimal

time lags between drillhole sampling and analysis. Sufficient drillhole material has been retained

should additional verification of results be required. Sample security, preparation and analytical

procedures are believed to be able to support representative sample assay results for estimation.

Submitted blanks did not raise any risks with regard to contamination.


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12. DATA VERIFICATION

The comprehensive program of multiple standard and blank inserted at regular but random intervals

has highlighted that the Dikulushi sampling is both accurate and precise. The accredited and

independent laboratories have evidence of good internal QAQC practices. These results, combined

with the good spatial distribution of QAQC sampling, support accurate, precise and uncontaminated

sample assay results and have been verified by the principal author and Qualified Person. According

to these results and the number of samples available for estimation, the Dikulushi and Kazumbula

drillhole databases provide satisfactory sample support and quality for estimating in situ

mineralisation.


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13. MINERAL PROCESSING AND METALLURGICAL TESTING

13.1. INTRODUCTION

Historically Anvil has completed a significant amount of testwork for Dikulushi, and a summary of

this work is presented below. Relevant operational data from the Dikulushi processing plant is also

tabulated.

As the underground ore will be mined from the same or close to the same areas as the ore

previously treated at or below the current open pit floor, it is not unreasonable to expect that it will

exhibit similar metallurgical characteristics during processing through the existing Dikulushi

Processing plant. It should be remembered that the underground was previously developed, mined

and ore processed through the current operating plant pre MWL ownership.

13.2. ANVIL TESTWORK

13.2.1. EARLY TESTWORK

The following information was supplied by Mawson West as background to the original design for

the process plant that was built at Dikulushi. Sedgman has not been able to review the original

testwork reports and as such cannot verify the information in this sub-section.

A significant amount of metallurgical testwork was undertaken by Anvil for the pre-feasibility phase

of their Dikulushi Project between February 1998 and April 1998 by the Minerals Engineering Group

of Mintek at their laboratories in Randburg, South Africa. Resource Management Group (RMG)

established and supervised the testwork on behalf of Anvil. Local coordination and support in South

Africa were provided by Fluor Daniel, Southern Africa. The Mintek data were used as the process

design basis for the pre-feasibility study completed by Signet Engineering in Perth in April 1998.

A previous testwork program was carried out by the Bureau de Recherches Géologiques et Minières

(BRGM), the results of which were available in Report no. 80 SGN 260 MIN, issued in April 1980. A

limited amount of preliminary testwork was initiated by Anvil and undertaken by Goldfields in

Johannesburg and was detailed in their report no. FL04\ks dated 4 November, 1996.

The metallurgical testwork program carried out by Mintek in 1998 was on various sulphide, oxide

and host rock samples from Dikulushi. The locations of these samples, their average grades and the

rock type classification are listed below in Table 13.1. Each composite comprised material from one

to three drillholes.


P a g e | 66

Table 13.1 Details of Dikulushi drill core used in Mintek metallurgical testing

% Copper Silver

Composite No. Drillholes Classifications Total Oxide g/t

1 DIK 15, 22 East-oxidised 9.5 1.5-2 360

2 DIK 28, 31 East-deeper 15.2 0.8 525

3 DIK 6, 11, 14 West-main 10.1 1.0 150

4 DIK 26 West-disseminated 2.8 0.3 60

5 DIK 5, 14 West-complex 9.0 1.3 50

6 DIK 12, 13, 23 East-transition 7.9 0.6 260

The sample nomenclature indicates that compositing was based upon special and oxidation

properties of the ore. Sedgman cannot comment on the representivity of these samples with

respect to the current study.

Physical tests were undertaken for typical composites of massive sulphide and light grey sandstone.

Flotation tests were carried out on primary, transition, oxidised and highly oxidised composites from

the east zone, and primary and complex sulphide composites from the west. These composites

represented an arbitrary sub-division of the ore body.

Head analyses revealed a relatively high total copper grade of 15.2% for the East Primary composite,

while the others were in the range of 8.2 - 11.4%, which was reasonably close to the target grade of

10% copper. Silver assays were variable, with a range of 138 - 562 g/t, the highest being for the East

Primary. Iron and sulphur levels were relatively low. Potential penalty elements identified were

lead and zinc in the West Complex, arsenic in the West Primary and West Complex, and fluorine in

all composites.

The previous testwork by BRGM in the 1980s indicated good flotation characteristics, with

recoveries ranging from 84 - 96% for copper, and 79 - 96% for silver. High grade concentrate grades

of 63 - 72% copper and 950 - 2,600 g/t silver were produced. BRGM found that sulphidation with

Na2S was required for oxidised material, though highly oxidised near surface ore was not tested.

Mineralogical examination revealed that the dominant copper sulphide mineral was chalcocite, in

both massive and disseminated forms. Some of the massive chalcocite was crystalline, and may

tend to slime during grinding. Complex sulphides in the west zone contained chalcopyrite, bornite

and sphalerite. Sphalerite is also common in other areas associated with chalcopyrite. Near surface

oxide contained malachite, azurite and chrysocolla. The latter did not float even when sulphidised.

Silver was assumed to be present mostly in solid solution in chalcocite, and occasionally as selenide.

Arsenic occurred as arsenopyrite and tennanite. Sandstone was the dominant host rock.

The physical tests revealed that the Dikulushi ore was of moderate hardness, with figures of 14.1 -

17.4 for the Rod Mill Work Index (RMWI), 10.5 - 12.5 for Ball Mill Work Index (BMWI) and 0.21 - 0.39

for Abrasion Index (AI) being reported. The higher indices generally related to the massive ore.

Flotation results at a grind size of 80% passing 75 microns were comparable to those in the BRGM

data, with recoveries of 71 - 97% for copper and 63 - 95% for silver. The lower figures were for near

surface highly oxidised material. The predicted concentrate grades were 48 - 70% copper, and 661 -


P a g e | 67

2,300 g/t silver. Detailed concentrate analyses revealed that fluorine was the only impurity over the

penalty threshold. Reagent usage appeared modest, except for the Na2S required for the oxidised

material, which required up to 3.2 kg per tonne of ore

13.2.2. LATER TESTWORK

Additional testwork was performed by Independent Metallurgical Laboratories (IML) in Perth during

2003. The related testwork reports have been reviewed and Sedgman has been able to verify the

information detailed in this sub-section.

Five separate copper ore composites from Dikulushi were used for the testwork:

high grade chalcocite

disseminated and low grade chalcocite

lead and zinc rich chalcocite

bornite

stockpiled dense media separation tailings.

The various chalcocite composite assays are detailed in Table 13.2.

Table 13.2 Head grades of chalcocite composites

Element Unit High

grade chalcocite

Disseminated

& low grade

chalcocite

Pb/Zn rich

chalcocite

Cu (Total) – Assay % 21.9 3.05 6.30

Cu (Total) – Calc. % 20.1 2.99 5.53

Cu (Total – Sequential.) – Calc. % 20.4 3.05 5.67

Cu (Acid Soluble) % 3.52 1.27 0.26

Cu (Cyanide Soluble) % 16.8 1.73 3.83

Cu (Residual) % 0.14 0.04 1.58

Ag ppm 624 75 23

Pb 39 ppm 21 ppm 1.58%

Zn 189 ppm 115 ppm 10.88%

The Sequential Diagnostic Leach Analysis identifies the oxide component as Acid Soluble copper, the

Secondary Sulphides (including Chalcocite and Covellite) report as Cyanide Soluble species and the

residual fraction relates to primary copper sulphides such as chalcopyrite.

Mineralogical examinations identified the abundance of various minerals as illustrated in Table 13.3.


P a g e | 68

Table 13.3 Relative abundance of significant minerals

Mineral

High

grade chalcocite

Disseminated

& low grade

chalcocite

Pb/Zn rich

chalcocite

+0.1mm -0.1mm +0.1mm -0.1mm +0.1mm -0.1mm

Chalcocite Dominant Dominant Dominant Dominant Minor Accessory

Malachite Major Major Major Major - -

Bornite Accessory - Accessory Trace Trace Accessory

Chalcopyrite - - Trace - Major Minor

Pyrite - Trace - - Major Minor

Sphalerite - - Accessory - Dominant Dominant Note. Dominant: >50%, Major: 20 - 50%, Minor: 10 – 20%, Accessory: 1 – 10%, Trace: <1%.

The comminution data derived for these composites relating to the Bond Ball Mill, Bond Rod Mill

and Abrasion Indices are summarised in Table 13.4.

Table 13.4 Comminution testwork results

Composite BRMWi (kWH/t) BBMWi (kWh/t) BAi

High Grade Chalcocite 15.7 12.4 0.1472

Disseminated & Low Grade Chalcocite 17.3 13.8 0.4224

Pb/Zn Rich Chalcocite 17.7 - 0.2360

A series of flotation tests was performed on the composites.

HIGH GRADE CHALCOCITE

There was minimal difference in rougher flotation performance between grind P80s of 75, 106 and

150 microns using a stainless steel mill. See Table 13.5.

Table 13.5 Effect of grind size on flotation performance (high grade chalcocite)

Grind P80 - mic

Cumulative Rougher Concentrates

Copper Silver

Assay (%) Distribution (%) Assay (ppm) Distribution (%)

75 52.0 97.8 1567 97.3

106 53.2 97.6 1632 97.3

150 54.9 97.7 1543 97.0

Using a grind P80 of 150 microns in each case, rougher flotation tests at potassium amyl xanthate

(collector) additions of 70, 105 and 140 g/t resulted in high copper grades and recoveries in each

case although flotation kinetics were significantly slower at the lower addition rate, see Table 13.6.


P a g e | 69

Table 13.6 Effects of collector addition on flotation performance (high grade chalcocite)

Collector Addition (PAX) – g/t


Copper Silver


70 55.3 95.8 1818 96.8

105 51.0 96.2 1644 96.8

140 54.9 97.7 1543 97.0

DISSEMINATED AND LOW GRADE CHALCOCITE

A set of flotation tests was conducted at various grind sizes. A grind P80 of 150 microns produced

similar results to the finer grind sizes, see Table 13.7.

Table 13.7 Effect of grind size on flotation performance (disseminated and low grade chalcocite)

Grind P80 - mic


Copper Silver


75 13.4 82.3 303 80.6

106 14.3 81.5 326 79.4

150 13.1 81.5 303 78.8

The effect of variation in collector dosing was investigated. Although a higher collector addition

produced better results, these tests were performed at the fine grind P80 of 75 microns and before

an optimised pulp Eh had been established. Consequently the testing was inconclusive, see Table

13.8.

Table 13.8 Effect of collector addition on flotation performance (disseminated and low grade chalcocite)

Collector Addition (PAX) – g/t


Copper Silver


100 22.4 75.8 521 75.3

165 25.5 78.0 615 79.5

PB/ZN RICH CHALCOCITE

Two sets of tests were performed to investigate the effect of grind size at different pulp Eh levels.

The results are shown in Table 13.9.

Table 13.9 Effect of grind size and Eh level on flotation performance (Pb/Zn rich chalcocite)

Grind P80 - mic Eh – mV

(Ag/AgCl/Sat KCl)


Copper Zinc

Assay (%) Distribution (%) Assay (%) Distribution (%)

75 150 11.9 98.0 23.8 88.6

106 150 13.1 94.8 27.3 82.0

106 70 12.0 97.6 23.2 90.7

150 70 12.6 98.1 25.3 90.4


P a g e | 70

The tests showed high copper recoveries but the copper grades were diluted by the amount of zinc

also reporting to concentrate.

A series of tests were performed to determine the effect of a range of Zinc Depressants – Sodium

Cyanide, Zinc Sulphate and Sodium Meta-bisulphite. The results were disappointing with only

sodium meta-bisulphite demonstrating any depression of zinc, but unfortunately it also depressed

copper.

A mineralogical examination of a first rougher concentrate showed that approximately 50% of the

sphalerite was locked with chalcopyrite and another 10-20% of the sphalerite was associated with

other sulphides.

MIXED CHALCOCITE COMPOSITE

A locked cycle flotation test was performed on a composite comprising 41.6% High Grade Massive

Chalcocite and 58.4% Disseminated and Low Grade Chalcocite which produced a calculated head

grade of 9.41% copper.

A combined rougher/cleaner copper concentrate grade of 54.6% was produced at an overall

recovery of 86.9%. The combined silver concentrate grade was 1,683 ppm at a recovery of 91.9%.

ROM LOCKED CYCLE TEST

In 2004 a locked cycle test was performed on a plant feed sample dated 25/11/2003 producing a

unit flash flotation cell, rougher and cleaner concentrate.

The head feed sequential analysis is shown in Table 13.10 and the test results in Table 13.11.

Table 13.10 Head grades of chalcocite composites

Element Unit 25/11/2003 Feed Sample

Cu (Total) - Assay % 9.18

Cu (Total) – Calc. % 9.23

Cu (Total – Sequential.) – Calc. % 9.31

Cu (Acid Soluble) % 1.98

Cu (Cyanide Soluble) % 7.32

Cu (Residual) % 0.01

Ag ppm Not Assayed

Pb ppm 41

Zn ppm 857

Table 13.11 Locked cycle flotation test results

Product Wt% Copper Silver

Assay (%) Distribution (%) Assay (%) Distribution (%)

Unit Cell Conc. 4.99 61.32 34.92 2400 39.21

Rougher Conc. 8.62 47.80 47.72 1600 45.14

Cleaner Conc. 5.32 14.25 8.47 305 5.31

Scavenger Tail 81.07 1.03 8.89 39 10.34

Calculated Head 100.00 8.79 100.00 306 100.00


P a g e | 71

The results indicate a combined concentrate grade of 42.1% copper at an overall recovery of 91.1%.

The combined silver concentrate grade was 1,447ppm at an overall recovery of 89.7%.

TESTWORK SUMMARY

Of the three chalcocite composites tested at IML in 2003 the high grade chalcocite composite was

the most relevant to the Dikulushi Open Pit Project. However it cannot be considered truly

representative as the head grade was far higher than the planned feed grade and operational data at

Dikulushi showed that there was a positive correlation between copper head grade and recovery.

Overall the testwork did demonstrate that provided the flotation conditions, including Redox

potential, was carefully controlled, chalcocite ore could be effectively recovered by flotation

producing fast kinetics, high concentrate grades and good recoveries.

13.3. PLANT OPERATIONAL RESULTS

MWL has indicated that the flotation plant at Dikulushi previously operated from 2004 to 2008 and

processed high grade ore from both the open pit and underground mine. According to Anvil

production data between September 2004 and April 2008, it achieved recoveries of 88.3% copper

and 88.5% silver, producing a concentrate containing 54.7% copper and 1,659 g/t silver. The plant

was shut down in November 2008 after treating low grade stockpile material during the last months

of operation.

In May 2010 the plant was refurbished and commenced production in June 2010 by treating low

grade stockpile ore and HMS tails. Over the past 3 years, recoveries vary between 60 - 70% for the

low grade stockpile material and the open pit cut back ROM feed grade recoveries have varied

between 75% - 95%. Concentrate grades in the last 12 months have averaged 56% copper and 1,515

g/t silver over the past year. Table 13.13 in the next section, shows the last 3 years production on a

month by month basis.

Sedgman has reviewed the production data as supplied by MWL for the periods Feb 2007 to Apr

2008 and June 2010 to May 2011, however, Sedgman has not reviewed the production data for the

period June 2011 to July 2013.

13.4. METALLURGICAL PROPERTIES OF THE CUT BACK ORE AND

UNDERGROUND ORE

The Dikulushi deposit was mined and processed by Anvil for several years and the high grade

chalcocite ore below the current pit floor has previously been processed in the mill during

underground mining operations. Anvil monthly production reports, for the previously mined

underground are tabled in table 13.12, where the mill feed was from the old open pit and the then

underground mining operations. In reviewing the historic operating data, it can be seen that the

copper recovery was approximately 90.4% over the period, with underground recoveries averaging

86.9 to 92.8%.


P a g e | 72

Table 13.12 Dikulushi processing summary (February 2007 – April 2008)

Month Blend %

ROM

RL mined (Ore Only)

Flotation Plant

Tonnes

Plant Feed Recovery Concentrate

Grade

Cu

(%)

Ag

(ppm)

Cu

(%)

Ag

(%)

Cu

(%)

Ag

(ppm)

Feb-07*

60 27,779 5.93 181 85.7 86.9 56.0 1730

Mar-07 100 860 pit stockpile 28,508 8.44 264 91.6 90.4 56.2 1734

Apr-07 100 860 pit stockpile 28,487 7.68 240 90.7 90.5 55.1 1722

May-07 100 850 pit stockpile 26,188 7.61 231 90.2 90.1 55.1 1670

Jun-07 100 850 pit stockpile 30,805 7.74 233 91.0 90.5 55.3 1654

July 07*

91.4 870 Dev 31,838 7.28 214 89.5 89.8 56.6 1668

Aug-07 100 stockpile 30,802 7.96 245 91.2 90.5 56.2 1717

Sep-07 100 850 Dev 25,934 7.97 258 91.3 91.4 54.9 1777

Oct-07 100 850 Dev & 890 Stoping 31,193 8.18 272 92.4 92.5 54.6 1821

Nov-07 100 870 Dev & 890 stoping 30,286 7.81 250 92.2 91.8 56.3 1793

Dec-07 100 870 Dev & 890 stoping 30,641 8.45 266 92.8 92.0 56.8 1772

Jan-08 100 830 Dev & 890 stoping 30,746 6.00 187 90.6 89.4 55.2 1694

Feb 08* 81.4 830 Dev & 870 Stoping 30,789 5.09 154 87.2 87.7 55.7 1687

Mar-08 100 830 Dev & 890/870

Stoping 37,998 5.50 170 88.2 88.5 54.5 1691

Apr-08 90 830 Dev & 870 Stoping 33,400 4.76 139 86.9 88.4 54.0 1601

Total 455,395 7.04 218 90.4 90.3 55.5 1721 * Low grade ore blended in with the development or stoping ore.

Table 13.3 below shows the current processing statistics for the MWL operations from June 2010

through to July 2013. During this period a combination of LG stockpile material from the old Anvil

open pit was processed in the early months, to satellite orebodies such as Boom Gate etc, through

to the current mill feed being exclusively from the open pit cut-back material. This process feed

material is set to continue until December 2013.

In reviewing the recent operating data, it can be seen that the copper recoveries realised over the

past 7 months, reflecting the fresh open pit cut-back material feed, was approximately 94.3% over

the period.

Table 13.13 Dikulushi processing summary (June 2010 – July 2013)

Jun-

10 Jul-10

Aug-

10

Sep-

10 Oct-10

Nov-

10

Dec-

10

Jan-11 Feb-

11

Mar-

11

Apr-

11

May-

11 YTD

Ore Processed tonnes 5,387 36,157 43,882 40,839 27,450 49,029 41,111 49,650 42,839 46,054 40,855 44,705 467,958

Mill Feed Grade Cu % 1.28 1.45 1.04 1.27 3.78 1.52 1.17 1.33 1.32 1.28 1.40 1.32 1.46

Mill Feed Grade Ag g/t 35.87 40.4 27.63 31.72 77.17 41.2 28.5 32.6 29.2 27.8 34.6 33.31 35.31

Tails Grade Cu Cu % 0.34 0.39 0.35 0.46 1.64 0.63 0.52 0.46 0.43 0.46 0.44 0.52 0.54

Tails Grade Ag Ag g/t 10.1 11.1 10.5 10.9 23.0 13.6 10.70 11.3 7.95 7.85 8.7 9.1 10.95

Conc Tonnes dmt 128 896 719 783 1,380 1,066 684 1001 890 893 906 865 10,211

Conc Grade Cu Cu % 38.7 43.5 42.7 43.0 44.1 41.5 39.74 40.1 41.6 39.35 40.2 41.7 41.66

Conc Grade Ag Ag g/t 1,067 1,138 1,107 1,119 1,139 1,188 1139 1070 1033 941 1092 993 1089

Cu metal in Conc dmt 51.45 389.6 306.9 336.7 608.5 442.7 272 400 366 351 365 361 4,251

Ag metal in Conc oz 4,384 32,778 25,581 28,177 50,534 40,726 25,057 32,737 29,385 27,279 31,904 27,559 356,101

Recovery Cu % 74.62 74.31 67.25 64.92 58.64 59.41 56.54 64.13 66.91 62.66 67.46 61.37 64.05

Recovery Ag % 70.57 69.88 65.62 67.65 74.20 62.68 66.45 62.70 73.09 63.34 71.25 61.3 66.68


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Figure 13.1shows a cross section of the Mineral Resource as it relates to the previously mined parts

of the orebody via the Anvil open pit, Anvil underground and MWL open pit cut-back.

Figure 13.1 Dikulushi Underground sources of ore - showing North-South section view at 50205E

The planned underground ore production is to be from the previously developed and mined levels at

the 810 mRL down to the fully developed 770 mRL and a minor amount of ore from the partially

Jun-11 Jul-11 Aug-11 Sep-11 Oct-11 Nov-11 Dec-11 Jan-12 Feb-12 Mar-12 Apr-12 May-12 Jun-12 YTD-12

Ore Processed tonnes 41,684 44,113 31,913 35,566 32,799 34,417 30,084 14,008 23,756 15,009 13,061 14,096 17,355 347,863

Reconciliated Mill Feed Grade Cu % 1.55 1.44 2.12 1.70 1.77 1.36 1.43 1.96 1.95 2.06 2.24 2.73 2.90 1.80

Reconciliated Mill Feed Grade Ag g/t 28.14 29.65 37.08 33.20 32.12 27.30 29.86 41.37 37.87 38.07 40.57 49.30 50.06 34.15

Tails Grade Cu Cu % 0.62 0.59 0.92 0.68 0.73 0.54 0.53 0.71 0.93 0.74 0.69 0.76 0.74 0.69

Tails Grade Ag Ag g/t 7.40 8.28 9.02 7.77 8.90 7.70 7.85 14.25 16.50 13.10 12.60 14.76 13.84 10.75

Concentrate Tonnes Produced dmt 991 924 944 926 863 723 699 495 780 633 621 804 1,032 10,437

Concentrate Grade Cu Cu % 39.54 41.36 41.36 39.96 40.13 39.62 39.28 36.00 32.02 32.07 33.46 35.24 37.15 37.83

Concentrate Grade Ag Ag g/t 803.56 997.08 911.84 883.23 865.56 900.22 936.86 781.60 667.08 604.74 600.72 620.20 622.99 790.95

Cu metal in Concentrate dmt 392 382 391 370 346 287 275 178 250 203 208 283 383 3,948

Ag metal in Concentrate oz 25,608 29,629 27,683 26,290 24,007 20,934 21,064 12,440 16,737 12,314 11,998 16,036 20,667 265,405

Recovery Cu % 60.7 60.2 57.7 61.0 59.8 61.3 63.6 65.0 53.9 65.6 70.9 73.7 76.0 63.0

Recovery Ag % 67.9 70.5 72.8 69.3 70.9 69.3 72.9 66.8 57.9 67.0 70.4 71.8 74.0 69.5

Jul-12 Aug-12 Sep-12 Oct-12 Nov-12 Dec-12 Jan-13 Feb-13 Mar-13 Apr-13 May-13 Jun-13 Jul-13 YTD-13




Tails Grade Cu Cu % 0.65 0.59 0.55 0.45 0.42 0.44 0.40 0.43 0.36 0.44 0.51 0.61 0.46 0.48


Concentrate Tonnes Produced dmt 1,448 1,334 1,476 1,200 1,138 2,711 2,685 2,939 2,936 3,087 2,470 2,798 3,723 29,944


Concentrate Grade Ag Ag g/t 620 677 622 725 1,733 1,462 1,532 1,787 1,620 1,499 1,903 1,952 1,864 1,515

Cu metal in Concentrate dmt 505 487 554 486 696 1,584 1,596 1,804 1,829 1,936 1,552 1,701 2,195 16,925

Ag metal in Concentrate oz 28,876 29,014 29,496 27,950 63,410 127,429 132,235 168,837 152,974 148,824 151,060 175,621 223,058 1,458,783

Recovery Cu % 76.9 75.2 77.4 78.7 92.4 92.6 93.3 93.8 93.4 93.9 95.8 95.4 94.2 91.5

Recovery Ag % 76.9 75.4 76.0 83.2 87.1 90.2 90.5 94.2 90.3 88.7 93.9 94.0 92.1 90.3


P a g e | 74

developed 750 mRL. Thus large portions of the underground ore reserve ore tons have been

previously treated in the processing plant, as can be seen in the production processing summary

table 13.12; and the more recent production processing data in table 13.13, and hence significant

variations in ore quality is not expected from the mining of the underground Mineral Reserves.


P a g e | 75

14. MINERAL RESOURCE ESTIMATES

14.1. DIKULUSHI MINERAL RESOURCE ESTIMATE

The Dikulushi Mineral Resource estimate was prepared in May 2009 by Mr. David Gray, Qualified

Person and principal author of the technical report which was originally submitted in February 2011.

The May 2009 Mineral Resource was subsequently updated in August 2011 using the latest available

survey data of the historical volumes mined by Anvil and updated pre-feasibility study cut-off grades

for the proposed cut back of the open pit.

A previous (October 2007) Mineral Resource estimate for Dikulushi was generated for the purposes

of evaluating underground Mineral Resources. The geological interpretation of copper-silver

mineralisation beneath the open pit was largely based on the diamond drillhole database and

enabled the main Footwall zone of mineralisation to be extended to the Kiaka Carbonates.

Between the October 2007 estimate and the May 2009 estimate, an additional 23,610 m of

underground, infill and extensional drilling was completed (Figure 14.1) across the Dikulushi ore

body and can be broken down by sampling type:

802 m were derived from underground channel sampling

3,747 m from underground grade control diamond drilling

4,789 m from RC drilling

14,272 m from surface diamond drilling.

The May 2009 estimate was based on all available data as at the end of November 2008, with no

outstanding core logging, sampling or assay results remaining. Since that time there has been no

additional data that impacts on the estimate of resources remaining below the base of the open pit

cut back that was completed in July 2013.

Dikulushi mineralisation (Figure 14.2, showing footwall mineralisation in green and hanging wall

mineralisation in orange) is characterised by a hydrothermal copper-silver vein system hosted by

Proterozoic sediments of the Upper Kundelungu Group, and has two distinct ore zones. A dominant

“Footwall” zone is intersected over a 230 m strike length with thicknesses of up to 25 m, which

decreases with increasing depth. This zone comprises semi-massive chalcocite and/or bornite veins,

strikes east-northeast and dips southeast at approximately 65°. Exhibiting good strike continuity, it

can be traced to depths of approximately 500 m below surface. A secondary “Hanging Wall” zone is

observed within 50 m of the Footwall zone, and comprises discontinuous, steeply dipping, chalcocite

veins, veinlets and disseminations. These dip at varying angles to the Footwall zone and may

occasionally intersect it. Apart from minor other occurrences, the Hanging Wall zone is largely

absent below the base of the open pit.

Grade interpolation was undertaken for total copper (%) and silver grade (g/t). Wireframes were

created for the domains and defined zones of similar weathering, faulting, stratigraphy and copper

grade. Sample copper and silver analytical results were composited to one metre interval lengths

per domain. Variography displayed reasonable continuity with low nugget values.


P a g e | 76

Figure 14.1 An oblique southward looking 3D view of drillhole type and distribution at Dikulushi

The resulting Mineral Resource statement was depleted for open pit and underground material

mined as surveyed from mined volumes and since the previous October 2007 estimate through to

November 2008. The estimate is representative of all data acquired. Mineral Resources have been

classified into Measured, Indicated and Inferred categories for the fresh sulphide mineralisation

located below the November 2008 pit surface, as per Table 14.1.

Table 14.1 Dikulushi Mineral Resource statement as at August 2011 above a 1.0% copper cut-off grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured 184 2.8 516 7.0 211

Indicated 90 2.8 251 5.6 114

Measured & Indicated 274 2.8 767 6.6 179

Inferred 136 2.8 380 6.8 91

It should be noted that this model has now been further depleted with the recently completed open

pit cut back; the depletion tables are presented and discussed in Section 14.2.


P a g e | 77

14.1.1. GEOLOGICAL AND MINERALISATION MODELS

Lithology and lode profiles were developed using five metre spaced north-south cross sections. The

ore body was modelled as a Footwall fault zone with sporadic mineralisation intersected within 50 m

of the overlying hanging wall. Two Hangingwall domains, as observed in the pit, were delineated

and modelled. The open pit has mined most of the weathered material and has exposed weathering

to depths of 35 m; the impacts of weathering were therefore not considered in the 2009 estimate.

Wireframes representing the boundaries relevant to the mineralisation were constructed in three

dimensions (3D) using north-south vertical cross sections. Mineralisation outlines were guided by

geological continuity between drillholes and a mineralisation threshold between 0.3% and 0.7%

copper.

Both blasthole and underground channel data (Figure 14.2) supported depth extensions of the

Footwall Fault zone.

Figure 14.2 A vertically oriented 3D view at Dikulushi, looking southwest, showing mineralisation lenses and current drilling


P a g e | 78

14.1.2. DRILL DATA FOR MINERAL RESOURCE MODELLING

Drill data was stored in Dikulushi’s on-site Access database. While some risk exists regarding the

reliability of manually handled data in an Access database, the drillhole de-surveying process

revealed only minor location errors, which were immediately corrected.

A plan view of drillhole data by type is presented in Figure 14.3. A total of 567 holes were available

for geological modelling, comprising 22,129 m surface diamond, 4,951 m surface reverse circulation,

1,285 m channel, 4,131 m underground grade control diamond and 1,369 m sludge samples. This

translates as a net increase of 23,610 m from the previous resource estimate.

Diamond drilling was undertaken along north-south oriented lines spaced 20 - 25m apart, with holes

at 25 m intervals along each line. To maximise the true widths of the intersections, most drilling was

angled at 50 to 60 degrees to the south. As the risk of undetected changes to orebody orientation

increases with depth, additional infill drilling will naturally assist in improving the confidence in

deposit geometry. In 2008, a total of 4 surface exploration drillholes were drilled to both infill and

extend Footwall zone mineralisation. While the deposit remains open at depth, this recent drilling

has led to only minor east-west extension.

Figure 14.3 A plan showing the distribution of drillhole types across Dikulushi; blasthole data from the pit have been excluded

Since twin-hole drilling was not completed, drilling and sampling methods were compared for

potential bias across a similar volume of the FW zone mineralisation using quantile-quantile (Q-Q)

plots. Diamond core was accepted as generally providing the most representative sample. This

comparison emphasises the difference in copper values between diamond and sludge hole samples

(Figure 14.4), with the latter decreasing as the former increases. As a direct result, sludge hole data

was not used in this estimate.


P a g e | 79

Figure 14.4 Quantile-Quantile (Q-Q) plot of Diamond (DD) drilled samples versus sludge drilled samples within a common area

14.1.3. DATA VALIDATION

A series of data validations were completed prior to de-surveying the drillhole data into a three

dimensional format. These included:

verification of collar coordinates with existing topography and underground development

wireframes, with virtually no problems observed

visualisation of downhole survey data to identify improperly recorded downhole survey

values, with all minor discrepancies corrected

dataset examination for sample overlaps and/or gaps in downhole survey, sampling and

geological logging data, with none observed

database interrogation for negative values representing codes such as ‘insufficient sample’,

with all such samples set to absent

examination for negative assays reflecting ‘below detection’ range; these values were all re-

set to 0.01%

testing for absent or duplicate samples, with none recorded.

14.1.4. DATA PREPARATION FOR MODELLING

The de-surveyed 3D assay drillhole file was coded and selected within the mineralisation and

lithological 3D wireframes. Each sample interval was coded with a mineralisation zone and

weathering profile, providing mineralised domain codes for estimation (Table 14.2). The coded

drillhole data was exported for subsequent geostatistical analysis and grade interpolation.


P a g e | 80

Table 14.2 Domain codes for Dikulushi modelling

Field name Domain Code

OREZONE Oxidised FW zone 50 Fresh FW zone 100 Shallow HW zone A 200 Shallow HW zone B 300 Internal FW zone waste 400 WEATH Soil to 5m 0.1 Oxidised to 35m 0.2 Transitional to 75m depth 0.3 Fresh rock 0.4 Air 0 MINED Not mined 0 Open pit mined 1 Mined underground 2 Open pit reserves 3

14.1.5. DATA COMPOSITING

To determine the most common sample length, the distribution of raw sample lengths was plotted.

Approximately 45% of the data had a sample length within a few centimetres of 1 m (Figure 14.5).

All data was composited to 1 m sample lengths, ensuring that intervals provided good resolution

across domain boundaries. The total raw sample length is identical to the composited total sample

length.

Figure 14.5 Cumulative distribution of sample lengths highlighting the dominant 1m sample length


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14.1.6. STATISTICS

Statistical analyses of the data, including spatial statistics, were carried out using Supervisor

software. The statistical analysis of composite copper grades was undertaken within each of the

final domains and the summary results are presented in Table 14.3.

Statistics for copper and silver were investigated by domain with histograms and probability plots.

The objective of the domain selections was to reduce internal variability and domain mixing, thereby

assisting with spatial analysis and providing a more robust estimate.

The selected domains appear to be well defined, with a minimal degree of mixing as depicted in

Figure 14.6 for Dikulushi’s principal Footwall zone.

Table 14.3 Summary statistics for copper % and silver g/t per domain

Waste

domain (0)

Oxide FW zone

domain (50)

Fresh FW zone

domain (100)

HW zone A

(200)

HW zone B

(300)

Internal FW

waste zone

(400)

Cu (%) Cu (%) Cu (%) Cu (%) Cu (%) Cu (%)

Samples 10429 1284 17145 956 204 1629

Min 0.01 0.01 0.01 0.01 0.02 0.01

Max 5.00 63.80 74.34 11.00 23.00 17.00

Mean 0.20 7.14 6.06 2.29 3.03 0.50

Std Dev 0.48 9.69 8.47 1.76 4.27 1.69

CV 2.37 1.36 1.40 0.77 1.41 3.41

Variance 0.23 93.92 71.81 3.09 18.23 2.84

Skewness 5.59 2.42 2.84 1.60 2.68 6.62

Log variance 1.61 2.08 2.75 1.21 1.87 2.02

Geometric mean 0.07 3.08 2.32 1.54 1.36 0.12

Ag (g/t) Ag (g/t) Ag (g/t) Ag (g/t) Ag (g/t) Ag (g/t)

Samples 5456 1108 16221 849 179 709

Min 1.00 1.00 1.00 1.00 4.00 1.00

Max 325.00 2615.00 1800.00 325.00 730.00 470.00

Mean 14.22 214.27 251.69 58.70 101.73 27.19

Std Dev 31.09 340.51 305.00 53.95 145.82 59.91

CV 2.19 1.59 1.21 0.92 1.43 2.20

Variance 966.27 115947.00 93023.50 2910.94 21262.30 3588.70

Skewness 6.19 2.70 2.02 2.11 2.70 4.98

Log variance 1.43 2.64 2.10 0.96 1.37 1.24

Geometric mean 6.13 69.46 111.94 39.20 50.07 11.82

14.1.7. SPATIAL STATISTICS

For Dikulushi, variography was analysed using composite data located within the mineralised

envelopes of each domain, based on the following methodology:

data was declustered prior to variogram modelling so as to remove the effect of closely

spaced blast hole and underground channel data

the principal axes of anisotropy were determined using semi-variogram (variogram) fans

based on normal scores variograms

normal scores variograms were calculated for each of the principal axes of anisotropy


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downhole normal scores variograms were modelled for each domain and adjusted to

determine the normal scores nugget effect

variogram models were then determined for each of the principal axes of anisotropy using

the nugget effect from the downhole variogram

the variogram models were back-transformed to the original distribution and used to guide

search parameters and complete ordinary kriging estimation.

Figure 14.6 Log histogram and probability plot for the main FW zone of mineralisation showing the results of robust domaining


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Variogram orientations were largely controlled by the strike of mineralisation and downhole

variography. Variogram models for silver and copper were similar, with silver tending to have a

slightly longer range of influence. Variogram models for the Footwall zone of mineralisation were

robust with a clearly-defined nugget value and well-defined structure (Table 14.4). Omni-directional

variogram models were derived for both HW zones and the upper oxidised FW zone. These domains

were not critical to this Mineral Resource estimate as this ore has already been mined. They were

included to ensure continuity with the deeper domains. Key variogram models for the main FW

zone are depicted in Figure 14.7.

Table 14.4 Dikulushi variogram models with angle1 about axis 3 (Z), angle2 about axis 1 (X) and angle3 about axis 3 (Z)

No. Assay Domain Angle1 Angle2 Angle3 Nugget St1 par1 St1 par2 St1 par3 St1 par4

1 CU 0 -5 130 10 0.06 4 6 5 0.54

2 AG 0 -5 130 10 0.06 10.5 5 6 0.51

3 CU 50 0 0 0 0.04 5 5 5 0.66

4 AG 50 0 0 0 0.04 5 5 5 0.63

5 CU 100 -10 100 -80 0.21 9 5 3 0.3

6 AG 100 -10 100 -80 0.2 11 4.5 1.5 0.29

7 CU 200 0 0 0 0.11 5 5 5 0.4

8 AG 200 0 0 0 0.12 4.5 4.5 4.5 0.33

9 CU 300 0 0 0 0.27 3 3 3 0.58

10 AG 300 0 0 0 0.28 4 4 4 0.46

11 CU 400 140 80 -100 0.07 5.5 5.5 5 0.7

12 AG 400 140 80 -110 0.06 3 3 3 0.79

No. Assay Domain St2 par1 St2par2 St2 par3 St2 par4 St3 par1 St3 par2 St3 par3 St3 par4

1 CU 0 11.5 15 9.5 0.29 191 39 10 0.12

2 AG 0 20 10 11.5 0.25 399 118.5 89 0.18

3 CU 50 34.5 34.5 34.5 0.3 - - - -

4 AG 50 57 57 57 0.33 - - - -

5 CU 100 26.5 18.5 8 0.25 84 49.5 15 0.25

6 AG 100 29 16 6.5 0.25 121.5 84.5 15.5 0.27

7 CU 200 15.5 15.5 15.5 0.32 38.5 38.5 38.5 0.17

8 AG 200 25.5 25.5 25.5 0.35 48 48 48 0.2

9 CU 300 16 16 16 0.15 - - - -

10 AG 300 23 23 23 0.26 - - - -

11 CU 400 40 33 14.5 0.23 - - - -

12 AG 400 31 31 31 0.15 - - - -


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Figure 14.7 Variogram models for copper % across the FW zone of mineralisation

14.1.8. BLOCK MODEL

The block model dimensions and parameters were based on the geological boundaries and average

drill grid spacing. Sub-blocks were used to ensure that the block model honoured the domain

geometries and volume. Block estimates were controlled by the original parent block dimension.

The individual parent block dimensions were 15 mE by 4 mN by 15 mRL, with sub-blocking allowed.

This dimension was supported by a kriging neighbourhood study which demonstrated little change

in the kriging efficiency or slope of regression (a measure of bias) from this block size to larger block

sizes.


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14.1.9. DENSITY ESTIMATES IN THE BLOCK MODEL

Density estimates were based on approximately 61 samples from the Footwall mineralisation and

1,236 samples from the surrounding waste material. These values have been tested and confirmed

via two mill feed samples. The assigned density of the Footwall ore zone was 2.8 t/m3 and the

surrounding waste material 2.6 t/m3.

14.1.10. DETERMINATION OF TOP CUTS

Top cut analysis was used to describe the maximum reasonable metal grade for a composite sample

value within a given domain. If the grade of a sample exceeded this value, the grade was reset to

the top cut value. The objective of applying top cuts is to minimise the risk of uniquely high metal

concentrations biasing individual block estimates, especially those located within areas of low

sample support.

Top cuts for Dikulushi were established by investigating univariate statistics and histograms of

sample values by domain. A top cut level was selected if it reduced the sample variance and did not

materially change the mean value. The following top cuts were applied to the data for resource

estimation (Table 14.5).

Table 14.5 Dikulushi - top cuts per domain

Domain Copper% Silver g/t

0 5 325

50 56 2000

100 - 1800

200 11 325

300 23 730

400 17 470

14.1.11. GRADE ESTIMATION

Grades for copper and silver were estimated into parent blocks of an empty domain coded block

model using ordinary kriging (OK). OK was deemed an appropriate interpolation technique owing to

near normal data distributions and differentiable grade ranges particular to the lode style

mineralisation. Estimation into parent blocks used a discretisation of 8 (X points) by 3 (Y points) by 8

(Z points) to better represent estimated block volumes.

14.1.12. ORDINARY KRIGING INTERPOLATION

Estimation parameters for kriging were based on variography, geological continuity and the average

spatial distribution of data. The first pass search radius was set within half to two thirds of the

variogram range to improve the quality of the local block grade estimate for areas of close spaced

drilling and to ensure that grade was not smeared laterally. Most blocks (75%) were estimated

within the first search radius. Subsequent search radii were set to ensure that remaining blocks

within the mineralised domain were interpolated with a copper grade.

For the ore domains, a minimum of 8 samples were required for a single block estimate and a

maximum of 40 samples to limit grade smoothing. Due to the long drillhole intercepts within the

orebody estimates were limited to a maximum of 10 samples per drillhole.


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Soft boundaries were created between the oxidised and fresh weathering domains in order to

represent the variable nature of this boundary and the transition in values. All other domain

boundaries were hard and data between domains was not included for estimation.

14.1.13. MODEL VALIDATION

The first pass of model validation included:

visual comparisons (Figure 14.8) of drillholes and estimated block grades

checks for negative grade estimates; if there were any, they were reset to a minimum 0.01 %

grade

checks to ensure that only blocks significantly distal to the drillholes remained without grade

estimates.

The model was further validated by statistical comparison of mean composite grades and model

grades, in addition to visual comparisons with drillholes. A table comparing the mean values for the

estimate with those of the data (Table 14.6) illustrates acceptable correlation.

Table 14.6 Mean statistics per domain comparing model estimates with data values

Domain Field Data Model % Variance

100 Ag g/t 219.01 201.22 8.12

100 Cu% 7.48 7.44 0.44

50 Ag g/t 172.11 169.34 1.61

50 Cu% 6.07 6.18 -1.81

400 Ag g/t 17.87 15.82 11.46

400 Cu% 0.42 0.42 0.19

Spatial statistical plots by domain are used to compare the mean model and drill grades data by

relative elevation slices (Figure 14.9). Model estimates respond well to changes in the composite

grade data, but local estimates are likely to be improved with additional drillhole intersections.

Based upon the summary statistics, visual validations and graphical plots, the OK estimates are

consistent with the drillhole composites, and are believed to constitute a reasonable representation

of the Footwall mineralisation.


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Figure 14.8 A plan view slice through the FW zone block model illustrating the good comparison between model estimates and the nearby drillhole data

Figure 14.9 A statistical plot of estimates versus drillhole data grades for successive 30m increments in elevation and the full strike length of the FW zone mineralisation


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14.1.14. MINERAL RESOURCE CLASSIFICATION

Classification of the Mineral Resource was primarily based on confidence in assayed grade,

geological continuity, and the quality of the resulting kriged estimates.

Geological confidence is supported by extensive open pit exposures and underground geological

mapping and channel data, which in turn reinforces drillhole logging and domain volumes.

Confidence in the kriged estimate is associated with drillhole coverage, analytical data integrity,

kriging variance and efficiency and regression slope values. Specifically, kriging variances below 0.2,

kriging efficiencies above 80% and regression slope values above 0.8 were considered appropriate

for a Measured Mineral Resource category of classification. Whereas the use of mean domain

density values is appropriate, subsequent models should make use of increased density data for

more robust estimates.

Regarding drillhole spacing, a Measured Mineral Resource category was considered appropriate with

a 20 m separation between drill holes and drill line spacing between 25 m to 50 m. An Indicated

Mineral Resource category was considered appropriate where there was a drill spacing of about 50

m to 75 m along drill lines and a line spacing of approximately 50 m. An Inferred Mineral Resource

category was considered where there was a drill spacing of about 75 m to 100 m along drill lines and

where the line spacing was around 100 m.

The Measured Mineral Resources are located below the pit and where underground sampling and

drilling is closely spaced. Indicated Resources extend as a consistent rim below the Measured

Resources. Confidence in the estimates deteriorates rapidly into Inferred Resources with the

increase in grid spacing and the short ranges of influence/grade continuity.

Figure 14.10 3D view of the Dikulushi model, looking south, and showing resource classification categories


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The Mineral Resource has been classified and reported using the guidelines of the JORC Code (JORC,

2004), which in turnalign with the Standards on Mineral Resources and Reserves of the Canadian

Institute of Mining, Metallurgy and Petroleum (CIM, 2000).

14.1.15. RESOURCE TABULATION AND INVENTORY

The Mineral Resource at Dikulushi is derived from that portion of the block model which occurs

below the current pit surface. Mineralisation appears to be open at depth, but is restricted to the

west by the Kiaka carbonates and is observed to pinch out to the east. Resources were depleted for

production and development from the underground mine, according to surveyed volumes. 112,000

tonnes of Mineral Resource was mined underground at an average of 8.5% copper.

The Measured and Indicated Resources for Dikulushi (Table 14.7) total 0.77 million tonnes at 6.6%

copper, and were determined above an economic cut-off grade of 1.0% copper. This is composed

of:

0.52 million tonnes at 7.0% copper in the Measured Resource category

0.25 million tonnes at 5.6% copper in the Indicated Resource category.

Table 14.7 Dikulushi Mineral Resource statement using a 1.0% copper cut-off grade as at August 2011

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)




Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)


14.2. MINERAL RESOURCE ESTIMATE COMPARISONS

14.2.1. MINERAL RESOURCE STATEMENT AUGUST 2011 VERSUS OCTOBER 2007

The August 2011 Mineral Resource estimates were compared to those of October 2007. These

results (Table 14.8) reflect an overall tonnage decrease of 23%, together with a 7% increase in

copper% and a 6% decrease in silver grade. Variance is against all resource categories: Measured,

Indicated and Inferred.

Notable category changes include a 124% increase in Measured Resource category tonnes and an 8%

increase in Inferred Resource category tonnes, associated with the presence of additional data from

underground exposures and from drilling. Most of these resources represent conversion from

Indicated Resource material.

There is a significant decrease in the Measured Resource copper % grades associated with

extensional drilling within deeper, lower grade areas. In contrast the deeper infill and extensional

drilling has supported an increase in the Inferred Resource copper grades.


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These comparisons were carried out using the 1.5% copper cut-off resource as the 2007 resources

were only available at that cut-off grade.

Table 14.8 Comparison of 2011 and 2007 Dikulushi Mineral Resource estimates

Dikulushi Mineral Resource statement as at August 2011, using a 1.5% copper cut-off

grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured 176 2.80 493 7.32 219

Indicated 86 2.80 241 5.79 118

Total Measured & Indicated 262 2.80 733 6.82 186

Inferred 129 2.80 361 7.11 94

Dikulushi Mineral Resource statement as at October 2007, using a 1.5% copper cut-off

grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured 78 2.83 220 9.63 289

Indicated 307 2.83 869 6.50 155

Total Measured & Indicated 385 2.83 1,089 7.13 182

Inferred 119 2.83 336 4.30 112

Comparison by percentage variation between the August 2011 and October 2007 results.

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured 126% -1% 124% -24% -24%

Indicated -72% -1% -72% -11% -24%

Total Measured & Indicated -32% -1% -33% -4% 2%

Inferred 9% -1% 8% 65% -16%

The 2011 Mineral Resource estimates have been guided by additional drillholes, underground

sampling, density, geological and in-pit blasthole data available as of November 2008. The

additional data has enabled an increase of some 21,000 copper tonnes from previous Indicated and

Inferred Mineral Resources to be upgraded to a Measured category.

Figure 14.11 illustrates the relative and cumulative change in copper tonnes between the 2007 to

2011 estimates. The 2011 Mineral Resource estimate has dropped by 14%, or a total of 13,000

tonnes of copper. Some of this is associated with mining depletion and significant changes to the

volumes of mineralisation. Grade reductions for the Measured and Indicated categories are offset

by increases in the Inferred category.


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Figure 14.11 A waterfall chart of cumulative Mineral Resource changes from 2007 to 2011

14.2.2. DEPLETION OF AUGUST 2011 MINERAL RESOURCES BY AUGUST 2013

OPEN PIT CUT BACK

The August 2011 Mineral Resource estimates were compared to those of August 2013, which

features the depleted Mineral Resources after the mining of the open pit cut back. These results, as

presented in Table 14.9 below, reflect an overall tonnage decrease of 37.3%, together with a 3.2%

decrease in copper% and a 9.3% decrease in silver grade (corrected for a previous reporting error).

The variance has been calculated for all resource categories, i.e., Measured, Indicated and Inferred.

Notable category changes include a 60% decrease in Measured Resource tonnes, a 41% decrease in

Indicated Resource tonnes, and a 3.9% decrease in the Inferred Resource category tonnes. All of

these changes are wholly due to depletion as a result of mining of the open pit cut back, which was

completed in July 2013.


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Table 14.9 Comparison of August 2011 and August 2013 Dikulushi Mineral Resource estimates, showing the Open pit cut back depletion


grade*

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured 184 2.80 516 7.0 211

Indicated 90 2.80 251 5.6 114


Inferred 136 2.80 380 6.8 91+


grade+

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured 74 2.80 207 5.4 163

Indicated 53 2.80 148 6.6 131


Inferred 130 2.80 365 7.0 160

Comparison by percentage variation between the August 2011 and August 2013 results.

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000)

Copper

(%)

Silver

(g/t)

Measured -51.1% 0% -59.9% -22.7% -22.7%

Indicated -41.1% 0% -41.0% +17.9% +14.9%

Total Measured & Indicated -53.7% 0% -53.9% -10.6% -16.2%

Inferred -4.4% 0% -3.9% +2.9% +3.2*

Note: + - The inferred silver grade was incorrectly reported at 91 g/t in the August 2011 Mineral Resource table and should have been 155 g/t.

Note: * - This % has been corrected to show the “Real” comparison % based on the correction note above.

Figure 14.12 is a waterfall chart which illustrates the relative and cumulative changes in Mineral

Resource tonnes between the 2011 and the 2013 estimates. Figure 14.13 shows tonnage-grade and

metal-grade curves for the depleted remaining Mineral Resource as at August 2013 as tabulated in

Table 14.9.


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Figure 14.12 A waterfall chart of cumulative Mineral Resource changes from 2011 to 2013

Figure 14.13 Grade tonnage curves for the combined remaining Measured and Indicated Mineral Resources

0.0

5.0

10.0

15.0

20.0

25.0

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0 2 4 6 8 10 12 14 16 18

Cu

%

Re

sou

rce

to

nn

age

an

d c

op

pe

r m

eta

l to

nn

age

Copper cut-off grade (Cu%)

Total Measured and Indicated tonnage and grade, remaining Dikulushi Mineral Resource

Resource tonnes (1,000,000) Cu tonnes (100,000) Cu %


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14.3. KAZUMBULA MINERAL RESOURCE ESTIMATE

The Kazumbula Mineral Resource estimates were prepared in November 2010 by Optiro in

conjunction with MWL geological staff, who worked on the Kazumbula Mineral Resources between

March 2009 and December 2010. The content of this technical report is guided by the reporting

requirements of the Canadian National Instrument 43-101 ‘Standards of Disclosure for Mineral

Projects’ and the JORC Code.

The Kazumbula deposit is located some 15 km north-northeast of the Dikulushi plant and has good

drillhole coverage from both Anvil and Mawson West. The Kazumbula deposit is relatively small

when compared to Dikulushi, and has a strike length of approximately 200 m.

RC and diamond drilling define the Kazumbula deposit with a grid spacing of approximately 20 m to

25 m. Only the MWL drilling data have been used for estimating the Kazumbula Mineral Resource

due to reliability issues with other drill campaigns.

14.3.1. GEOLOGICAL AND MINERALISATION MODELS

The Kazumbula mineralised volume (Figure 14.14) was delineated on vertical sections per drill line.

A 0.5% copper cut-off was used as a guideline for defining the mineralised volume. The delineated

string envelopes per section were linked with wireframe surfaces to define the mineralised volume

of the Kazumbula ore body. The Kazumbula mineralisation exhibits good downhole and between-

hole continuity. Mineralisation was also guided by the position of the interpreted fault surface. The

boundary between oxidised and sulphide mineralisation was modelled according to the logged

geology, oxidation and mineralogy; however, due to the relatively small size of the Kazumbula

orebody and the limited number of intersections within the oxide and sulphide domains, it was not

deemed appropriate to subdivide the mineralised domain.


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Figure 14.14 Kazumbula vertical section, looking north, highlighting the modelled mineralisation as per the RC and diamond drilling

14.3.2. DRILL DATA FOR MINERAL RESOURCE MODELLING

The deposit was drilled by Anvil during 2008, but MWL was not able to completely validate these

drillhole data (i.e. logging, sampling, assay, drillhole collars and downhole surveys), and as a result

the Anvil data were not used in this resource estimate. The Anvil drilling data has, however, assisted

MWL with drillhole planning and targeting of the Kazumbula deposit. MWL drilled RC and diamond

holes (Table 14.10) to define the near surface copper mineralisation during August and September

2010. The drillhole spacing was approximately 15 m on drill lines spaced 20 m apart (Figure 14.15).

Drillholes were drilled at approximately 60 degrees to the south-southeast to optimise the angle of

intersection with the orebody. No twin holes were completed for this programme, but significant

mineralised intersections are comparable with those completed by Anvil.


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Table 14.10 MWL drilling at Kazumbula

Prospect Type No Holes Metres Samples

Kazumbula RC 17 1676 1676

Kazumbula DDH/tail 10 674.4 674

The RC drilling was completed by Titan Drilling of Lubumbashi, utilising a truck mounted RC rig. A

supervising geologist was on site at all times during the drilling and industry standard procedures

were followed during the RC drilling programme. The diamond drilling was contracted and

completed by Chantete Emerald, who completed six diamond holes (HQ3) from surface and four

diamond tails from RC pre-collars (HQ3).

Figure 14.15 Plan showing the distribution of RC and diamond drillholes across the Kazumbula deposit.

14.3.3. DATA VALIDATION

Drillhole collar coordinates were surveyed by the qualified Dikulushi mine surveyor. Collars were

verified against the topography. Visual inspection of the downhole survey measurements was

completed in order to identify any anomalously different bearings and dips. Micromine software

was used to validate the drillhole logging and sampling data for any gaps or overlaps, with only

minor errors identified. These errors were associated with typing and were corrected immediately

in the MWL database.


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14.3.4. DATA PREPARATION FOR MODELLING

The de-surveyed 3D assay drillhole file was composited to two metre lengths. As discussed, a single

mineralised domain was used to model the Kazumbula orebody. As a result, the two metre

composites were deemed appropriate for defining the extents of mineralisation and reduced some

of the variability associated with the one metre samples.

14.3.5. STATISTICS

Statistical analyses of the two metre sample data, including spatial statistics, were carried out using

Supervisor software. The summary results are tabulated in Table 14.11.

Statistics for copper and silver grade data were investigated with histograms and probability plots

(Figure 14.16). The statistics highlight only minor internal variability and domain mixing, thereby

assisting with spatial analysis and supporting a reasonable estimate.

Table 14.11 Summary statistics of the two metre composite data for Cu% and Ag g/t for the Kazumbula deposit

Cu (%) Ag (g/t)

Samples 137 137

Min 0.03 2.50

Max 8.41 142.00

Mean 2.00 22.48

Std Dev 1.59 27.20

CV 0.79 1.21

Variance 2.53 739.86

Skewness 1.53 2.09

Log variance 0.83 1.42

Geometric mean 1.44 11.56

Log mean 137 137

14.3.6. SPATIAL STATISTICS

For the copper and silver variography, no definitive anisotropy was evident from spatial analysis. A

standardised nugget value of 0.37 was clearly defined from the downhole variography (Figure

14.17). Variography was oriented according to the plane of the fault and mineralisation, which dips

at 75 degrees towards 340. An isotropic variogram model (Figure 14.17) in this plane of

mineralisation was used to define the horizontal continuity of 60 m for copper and silver grades.

The true width variogram range was set to 20 m.

14.3.7. BLOCK MODEL

The block model dimensions were guided by the mineralised wireframe shape, orientation and

volume, together with the drill grid spacing. Sub-blocks were used to ensure that the block model

honoured the wireframe volume. The block model volume was within 1% of the mineralised

wireframe volume. The individual parent block dimensions were 15 mE by 10 mN by 2 mRL, with

sub-blocking allowed down to 5, 4 and 1 m respectively. Block estimates were controlled by the

original parent block dimension.


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Figure 14.16 Histogram and probability plots for the Kazumbula deposit two metre sample data.


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Figure 14.17 Variogram modelling for Cu % in the plane of mineralisation.

14.3.8. DENSITY ESTIMATES IN THE BLOCK MODEL

Diamond core samples were prepared by ‘squaring off’ the ends of approximately 10 cm to 20 cm

billets of half core. A total of 118 bulk density (BD) measurements were made of dried half core.

The same piece of core was then measured in water on a suspension cage below the same electronic

scale. The conventional formula for BD was then applied, viz.

BD = Dry Sample Weight / (Dry Sample Weight – Wet Sample Weight)

The average BD measurements and statistics for Kazumbula for mineralised and un-mineralised core

are shown in Table 14.12. The lower density for the mineralised oxide samples is explained by the

lack of hematite alteration in the mineralised samples, being replaced by an illite clay assemblage

and lowering the overall density of the rock.

Table 14.12 Density estimates for the Kazumbula deposit

Type SG

Mineralised Oxide 2.41

Unmineralised Oxide 2.47

Mineralised Sulphide 2.65

Unmineralised Sulphide 2.61

14.3.9. GRADE ESTIMATION

Grades for copper and silver were estimated into the parent blocks of an empty block model using

ordinary kriging (OK). OK is believed to be an appropriate estimation method due to the pseudo-

normal data distributions of the mineralisation. Estimation into parent blocks uses a discretisation

of 6 (X points) by 6 (Y points) by 2 (Z points) to better represent estimated block volumes, in addition

to applying an octant sample selection strategy of four sectors and a minimum of 4 samples and

maximum of 20 samples per sector.


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14.3.10. MODEL VALIDATION

The model was validated by visual techniques. Block estimates, while smoothed, do reflect the

average higher and lower grades of the drillhole samples. In addition, the model was also validated

with an inverse distance to the power of two estimate (ID2), using the same sample selection

parameters and search parameters. The mean grade estimate compared to the mean drillhole value

and the mean ID2 estimate are all close to within 10% of each other. The declustered mean of the

drillhole data has a value of 1.8% copper (Table 14.13).

Table 14.13 A table of mean statistics comparing model estimates with data values

Field Data ID2 Declustered data OK estimate

Copper % 2.0 2.0 1.8 1.8

From the summary statistics and visual validations, the OK estimates are consistent with the drillhole

composites, and while smoothed, are believed to constitute a reasonable representation of the

Kazumbula grade.

14.3.11. MINERAL RESOURCE CLASSIFICATION

Classification of the Kazumbula Mineral Resource was based on quality of sample assays, grid

spacing, the assigned density and the resulting kriged estimates. The Kazumbula deposit has been

classified in its entirety as an Indicated Mineral Resource. The mineralised volume is adequately

supported by a regular 20 m grid of drillhole intercepts and has been defined using an effective 0.5%

copper cut-off.

The Mineral Resource has been classified and reported using the guidelines of the JORC Code (JORC,

2004), which in turn comply with the Standards on Mineral Resources and Reserves of the Canadian

Institute of Mining, Metallurgy and Petroleum (CIM, 2005). A resource summary is given in Table

14.14.

Table 14.14 Kazumbula Mineral Resource statement as at November 2010.

Category Volume

(m3*1,000)

Density (t/m

3)

Tonnes (*1,000)

Copper (%)

Silver (g/t)

Indicated Oxide Mineral Resources 66 2.41 159 1.75 14.7

Indicated Sulphide Mineral Resources 60 2.65 160 1.89 22.9

Total Indicated Mineral Resources 126 2.52 318 1.82 18.8


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15. MINERAL RESERVE ESTIMATES

15.1. DEPLETION OF THE OPEN PIT RESERVES

The open pit Mineral Reserves have been depleted with the mining of the Dikulushi open pit cut

back, which commenced in August 2011 and was completed during July 2013. Processing of the

stockpiled cut back ore is continuing, and it is estimated that this ore will be processed by mid-

December 2013. Table 15.1 shows the Dikulushi Mineral Reserves statement as at August 2011.

Table 15.1 Dikulushi Mineral Reserve statement as at August 2011, using a 1.0% copper cut-off grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000) Copper (%)

Silver

(g/t)

Proven 66.6 2.8 184.7 7.27% 207

Probable 127.8 2.8 354.3 5.51% 169

Total Proven and Probable Reserves 194.4 2.8 539.0 6.12% 182

The Open Pit Mineral Reserve estimate was based on the Open Pit reaching the 810 mRL. Mining

ceased at the 825 mRL due to safety concerns, with some isolated sections of the pit wall

deteriorating beyond what was predicted. Table 15.2 shows the depleted Mineral Reserves post the

cessation of mining of the open pit cut back.

Table 15.2 Depleted Dikulushi Mineral Reserve statement as at August 2013, using a 1.0% copper cut-off grade

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000) Copper (%)

Silver

(g/t)

Proven 1.0 2.8 2.7 6.8 186

Probable 29.9 2.8 83.7 5.5 188

Total Proven and Probable Reserves 30.9 2.8 86.4 5.5 188

The above remaining Mineral Reserves have been incorporated into the Underground Mineral

Reserves, which are now presented in the following sections and discussions below.

15.2. UNDERGROUND MINE DESIGN AND SCHEDULE BASIS

15.2.1. EXISTING WORKINGS

Mine design for the Dikulushi underground mining operations has utilised the existing decline level

development completed by Anvil, which was completed prior to abandoning the underground. The

mining operations consisted of a decline and level development down to the 750 mRL. The capital

and development design criteria now used by MWL for the underground Mineral Reserve estimation

are the same as the original underground workings completed by Anvil, with the following design

cross-sectional areas:

decline and decline stockpiles; 5.5 m high by 5.5 m wide arched profile

level access / level stockpiles; 5.0 m high by 5.0 m wide arched profile


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return airway level drives; 5.0 m high by 5.5 m wide arched profile

cut and fill ramp access; 5.0 m high by 4.5 m wide square profile

ore drives / cut and fill; 5.0 m high by 4.5 m wide square profile

return airway rise; 5.0 m diameter.

Prior to the abandonment of the underground mining operations, Anvil completed stoping activities

in the upper levels of the mining operations between the 910 mRL and 850 mRL. The stopes were

mined using long-hole stoping. The success of stopes was limited due to ore dilution from the

surrounding waste material located around the hanging wall. Handheld stoping was also trialled in

the latter stages in some selected stopes, with better success.

Figure 15.1 shows the open pit cut-back and the original underground workings, as developed

previously by Anvil, which will now be utilised by MWL to form the basis of the underground Mineral

Reserve development and production. Additional development will be undertaken to access the un-

developed Mineral Reserves.

Figure 15.1 Existing workings, showing as built underground development (grey), and the as-built pit (green)


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15.2.2. BASIS OF THE UNDERGROUND DESIGN AND SCHEDULE

The Mining schedule and designs focussed on the Measured and Indicated Mineral Resources, and

were used in the development of the Mineral Reserves. No Inferred Mineral Resource Material was

used in the schedule or for economic evaluation of the Mineral Reserves. Figure 15.2

Underground reserve, showing as-built underground development (grey), as-built pit

(green), and measured (purple) and indicated (red) Mineral Resources and additionally shows the

relationship between the current development (as built), and the Measured and Indicated Mineral

Resources.

Figure 15.2 Underground reserve, showing as-built underground development (grey), as-built pit (green), and measured (purple) and indicated (red) Mineral Resources

15.3. CUT-OFF GRADE CRITERIA

Cut-off grade evaluation for the deposit has been completed using a Net Smelter Return (NSR)

calculation method due to the polymetallic nature of the deposit, as both the silver and copper

metals provide significant value to the final revenue of the mine.

The first part of the calculation process was to identify the NSR of one tonne of material with an

average grade of 5.38% copper and 128 g/t silver using Equation 15.1. The average grades for the

copper (x1) and silver (x2) were taken from the Mineral Reserve schedule.

Equation 15.1 Polymetallic NSR using average metal grades

( ) ( ) ( ( ) ( )


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x1 - is the average copper grade expressed as a percentage

x2 - is the silver grade expressed as a percentage

r1 - is the percentage of copper recovered when producing the concentrate

r2 - is the percentage of silver recovered when producing the concentrate

p1 - is the copper concentrate percentage received after smelting charges

p2 - is the copper concentrate percentage received after smelting charges

V1 - is the market value of copper per tonne

V2 - is the market value of silver per tonne

R1 - is the refining cost of the copper per tonne of copper produced

R2 - is the refining cost of the silver per tonne of silver produced

Cs - is the smelting and refining cost per tonne of copper concentrate

Ct - is the smelting and refining cost per tonne of copper concentrate

K - is the number of ore tonnes required to be mined to produce one tonne of copper

concentrate.

The values used in the calculation were derived from both the mining schedule created for the

Mineral Reserves Schedule and MWL’s financial model. The K value used was calculated using the

tonnes and grade from the mining schedule. From the application of Equation 15.1 a NSR value of

one tonne of ore averaging 5.38% copper and 0.0128% silver was calculated. The NSR value of one

tonne using the average copper and silver grades from the schedule was $329.

To allow the calculation of the NSR cut-off value for one tonne of rock with variable copper and

silver grades, Equation 15.2 has been derived from Equation 15.1. This equation provides a

mathematical relationship that allows the calculation of one variable component, if the other two

variable components are known, providing a breakeven revenue position. For example, if the

copper grade and an NSR cut-off value are known for a tonne of material, the silver content can then

be calculated. This allows the metal content of each tonne mined to be economically evaluated with

the aim of determining whether the material should go to the waste dump or be sent to the mill for

processing.

Equation 15.2 Variable grade NSR equation

A stoping (cut-off) value for the underground mining operations was determined using

Equation 15.3. The stope cut-off value relates to the drive/drift development used to extract ore

from the level development.

Equation 15.3 NSRc stope Stope cut-off equation

( )

- is the copper processing cost per tonne of ore mined and milled

- is the incremental silver processing cost per tonne of mined and milled

- is the waste processing cost per tonne mined

- is the ore mining cost per ore tonne mined


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The mining costs used in the evaluation of the were taken from the mining schedule

prepared for the Mineral Reserves, with the processing costs from MW’s financial model. As the

processing cost provided did not differentiate between the individual metals contained, the variable

for incremental cost of silver processing was set to $0 per ore tonne. When applying the mining and

processing costs, a NSRc Stope value of $143.44 per tonne is required to determine whether the

material should be mined or not.

The development cut-off was calculated using the Equation 15.4. The NSR value from this

calculation is used to determine whether material that must be mined from the underground

operation should be sent to the mill or waste dump. This calculation determines the NSR value

based on the difference between the mining and processing cost of ore and waste.

Equation 15.4 – NSRc dev Development Cut-off equation

( ) ( ) ( )

- is the overhead costs associated with the mining and processing 1 tonne of ore

- is the overhead costs associated with the mining and processing of 1 tonne of waste

- is the waste mining cost per waste tonne mined

When applying Equation 15.4, a value of $71.38 per tonne is required to determine if a

tonne of material that has to be mined should be processed or not.

Due to the polymetallic nature of the orebody, the cut-off grades are variable. Figure 15.3 shows a

relationship between the metal grades and the NSR cut-off for both the development and stope cut-

offs. Each of the lines shown on the graph has been calculated using the variable grade NSR

Equation 15.2, with the NSR value in the equation substituted with either the NSRc Stope or .

The silver content in g/t was then calculated for a variety of copper grades to produce the graph

lines.


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Figure 15.3 Relationship between cut-off NSR and metal grades

15.4. MINING RECOVERY AND DILUTION

Mining ore loss has been estimated at 5%. Ore loss is estimated to occur as a result of small gaps

left in the backfill process, caused by the inability to properly level off the top of backfill. This means

that when using the backfill as the floor for the next level, the fill will generally have a rounded

profile, and this will result in some material being lost in these corners as the next level is developed.

There will also be a small amount of ore loss resulting from general activities, such as rehandle and

material movement.

Mining dilution for the project has been estimated at 8%. This has been calculated by allowing a

200mm dilution skin on the walls of development drives, which has then been used to determine the

percentage dilution using the cross-sectional area of the drives. The figures use the average

dimensions for development drives, and also assume a square cross-section for calculation purposes.

Table 15.3 outlines the specific numbers for each development type.

Table 15.3 Mining dilution table

Development Width

(m) Height

(m) Drive

Area (m2) Dilution

width (m) Dilution skin

area (m2) Dilution

(%)

Ore

Ore Drive 4.50 5.00 22.50 0.20 2.00 8.9

Waste

Decline 5.50 5.50 30.25 0.20 2.20 7.3

Decline Stockpile 5.50 5.50 30.25 0.20 2.20 7.3

Level Access 5.00 5.00 25.00 0.20 2.00 8.0

Level Stockpile 5.00 5.00 25.00 0.20 2.00 8.0

Return Airway Drive 5.00 5.00 25.00 0.20 2.00 8.0

Cut and Fill Ramp 4.50 5.00 22.50 0.20 2.00 8.9


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Average 5.08 5.17 26.26 0.20 2.07 7.9

15.5. UNDERGROUND MINERAL RESERVE TABULATION

A financial model has been developed and analysis indicates that a positive return is expected.

There are several areas where it is considered by Optiro that conservative estimates of mining costs

have been made. It should be noted that the waste dilution rate of the ore is considered to be

conservative for this type of operation, mining style and deposit.

The resulting Mineral Reserves for the underground mine are only based upon Measured and

Indicated Mineral Resources from the depleted Mineral Resources post mining of the open pit cut

back. The Mineral Resources have been classified as Indicated due to the risks associated with

underground mining of this deposit.

The resulting Mineral Reserves are supported by historical production and current processing data

and are tabulated in Table 15.4, using a cut-off grade based on an NSR value of US$329/t, at a

copper price of US$3.08/lb and a Silver price of US$20 per oz. All stated Mineral Resources are

inclusive of Mineral Reserves. The Mineral Reserve, as per the CIM definition, incorporates mining

losses and diluting materials brought about by the mining operation.

Table 15.4 Dikulushi Mineral Reserve statement as at September 2013

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000) Copper (%)

Silver

(g/t)

Proven 0 0 0 0 0

Probable 62 2.8 173 5.2 127

Total Proven and Probable Reserves 62 2.8 173 5.2 127

Note:

1) Cut-off grade is based on a NSR value of US$329/t, at a copper price of US$3.08/lb and a

Silver price of US$20 per oz.

2) The above ore reserve does not include any Inferred category Mineral Resource material


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16. MINING METHODS

16.1. HISTORICAL MINING

Anvil commenced Underground Mining operations at the Dikulushi deposit in 2006 and continued

through until 2008, when falling copper prices and challenging mining conditions led to the

underground operations being abandoned. During the period of underground mining, mine

development was completed from the surface down to the 750 mRL, with extensive level

development completed from the 790 RL to the surface. Longhole open stoping activities were also

completed between the 900 and 850 mRL’s at the eastern end of the deposit.

MWL acquired the Dikulushi mining operations through its purchase of Anvil’s subsidiary AMC in

2010, and recommenced the open pit mining operations through the implementation of a cut back

of the original pit. The cut back was designed to reach the 805 mRL, and in doing so, would mine out

parts of the “old” underground workings that had been previously developed by Anvil. Mining of the

cut back commenced in late 2011 and continued through to July 2013, when mining was stopped at

the 825 mRL. This was 20 m short of the cut back design depth, due to safety concerns with some

isolated sections of the pit wall deteriorating beyond what was predicted.

16.2. PROPOSED MINING METHOD – CUT AND FILL

The planned mining method to be used at the Dikulushi underground operations is cut and fill, using

a combination of overhand and underhand variants. Initial ore production will be mined from the

800 mRL down to 755 mRL. Sections of the ore body between these two levels have good and poor

rock mass characteristics and contain the bulk of the Measured and Inferred material contained

within the Mineral Resource. Due to the variability of the rock mass in the zone, the backfilling of

drives will be completed using one of three backfill methods

Cemented rock fill (CRF),

Cemented aggregate fill (CAF)

Un-cemented rock fill (RF).

Use of the different backfill types as planned in the mining schedule has been determined by the

mining sequence and related mining activities that will be conducted after the fill has been placed.

16.2.1. OVERHAND CUT AND FILL

Overhand cut and fill is a variant of cut and fill which moves from the bottom of the orebody

upwards, using the fill from previous levels as the floor for the next level. Levels accessed are

excavated from the decline through to the orebody, perpendicular to the strike of the orebody.

From these accesses, ore drives are mined out along the strike of the orebody (shown in Figure

16.1), and once a level is fully extracted it is backfilled with CRF. Once the cement is cured, mining

recommences directly on top of the backfilled level, using the backfill as the floor for the next ore

drive. Overhand is the more productive of the two cut and fill variants suggested, and can be used

where the ore (which forms the backs of the mining levels) is competent enough to work under


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safely. Overhand cut and fill is also initially more capital-intensive than underhand, as it requires

development to the bottom of the ore block before commencing mining in an upwards direction.

Figure 16.1 Overhand cut and fill mining process

16.2.2. UNDERHAND CUT AND FILL

Underhand cut and fill mining works in a similar way to the overhand variety; however, mining

advances downwards underneath previously mined and backfilled levels, with backfill forming the

backs of each successive level. This is done by driving an access from the footwall of the deposit

through to the orebody. In the narrow sections of the orebody stoping is done by developing a

single strike drive to extract the ore on the level. The strike drives are developed and designed to

minimise openings created for the extraction of ore, with the openings kept as small as possible thus

minimising the amount of waste dilution taken with the ore being extracted. Once the orebody on

the level has been fully extracted using the drive, cemented fill (CAF) is then used to fill up the void

created by mining. After the backfill has been given sufficient time to cure and obtain the

appropriate strength requirements, the next level is then developed directly below the level above

with the cemented backfill forming the backs (roof) of the stoping drives developed below. Figure

16.2 shows the development sequence of this mining method in a horizontal long section view.

Underhand cut and fill is the less productive and a higher cost method; however, it has the

advantage of being able to control the condition of the backs. This means that it is suitable for any

areas where the ore is not competent enough for the overhand method, or anywhere where there is

a high propensity for rockbursts.


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Figure 16.2 Underhand cut and fill mining process

The primary mining method that has been selected and scheduled is the overhand cut and fill

method, which accounts for 70% of the ore extraction. The underhand mining method is primarily

used where the ore is extracted around previously developed levels and accounts for approximately

20% of the ore. The remaining 10% of the ore is extracted by Longhole stoping methods, which are

confined to stope areas within the crown pillar below the pit floor.

16.2.3. MINING OF WIDER SECTIONS OF THE OREBODY

In the sections of the orebody where the mineralisation exceeds the drive development width

(4.5m), multiple drives will be developed to extract the ore in these zones. The development

sequence of all drives will involve developing the initial drive along the hanging wall contact to the

end of the orebody. This drive will then be filled with either CRF or CAF. Once the fill has been

allowed to cure in the hanging wall drive, a second drive in the footwall will developed. Once the

footwall drive development has been completed, CRF or CAF will be used to backfill the drive.

Additional drives will then be developed to extract the remaining parts of the orebody between the

hanging wall and footwall contact drives. This process should allow for complete extraction of the

orebody where it exceeds the maximum drive width specified under the geotechnical requirements

(4.5 m). Figure 16.3 is an example of the extraction sequence for sections where the orebody width

exceeds the maximum stoping drive width.


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Figure 16.3 Diagrammatic representation of sequential mining in wide orebody areas

The current design work produced for the underground has access of the decline to the orebody at

20 m vertical spacing. Each of the level take-off points from the decline will provide access to four

stoping levels, with the levels developed from top to bottom (Figure 16.4 shows the access

development of the stoping levels from the decline). The bottom stoping drives from each level

access will sit directly above the top stoping drive from the decline access, 20 m below.

3

4

5

2

6

1

8m

1

3

15m

5

7

Plan view


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Figure 16.4 Orebody access development

As the mining sequence will require multiple levels to be mined in the same period to provide ore

feed to the processing plant, controls will need to be in place to maintain a minimum horizontal

distance between mining levels. The vertical spacing between all drives should be a small proportion

of the maximum void space created from the level above, in order to maintain a 1 to 1 void to pillar

ratio (Figure 16.5) between active mining levels. This has been employed to minimise the risk of

rock mass failure between the mining levels.

Figure 16.5 Pillar ratio diagram

16.2.4. PROPOSED MINING METHOD – EXTRACTION OF THE CROWN PILLAR

Extraction of the crown pillar from the 815 mRL to the base of the pit at the 825 mRL is to be

extracted using a longhole stoping method. The extraction of ore via this method will involve mining

through back filled sections of the 810 mRL to create a drive for the drilling of the ore from

underground, with the blasted ore transported to the mill via the underground decline. Extraction

and transportation of the ore via the open pit workings was not considered due to the possible

undercutting of the existing pit walls.

Removal of the crown pillar, as set out in the schedule, is undertaken at the end of the underground

mine’s operational life as it stands. The reason for this is to minimise the risk of water inrushes into

the underground and the potential destabilising of the open pit walls. MWL plans to continue


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underground exploration drilling of the orebody, and will review the success of those programs prior

to undertaking to remove the crown pillar.

16.3. GEOTECHNICAL DESIGN PARAMETERS

16.3.1. STOPE LAYOUT AND SEQUENCE

The mining sequence and layout developed for the mineral reserves schedule has been created in

consultation with both Mike Turner from Turner Mining and Geotechnical Pty Ltd and John Keogh

from Peter O’Bryan & Associates. For the purpose of the stope layout and designs, drives have been

limited to 4.5 m wide by 5 m high in the ore body. This is due poor ground conditions experienced

when mining was originally conducted.

16.3.2. DRILL AND BLAST

Rock breakage for the development of the underground mining operations will be conducted using

drill and blast. There are currently three development drills (Jumbos) located at the Dikulushi

operations from previous mining activities. These three jumbos will be used to complete all drilling

activities for horizontal mine development and stoping within the mining operations for ore and

waste. The three jumbos on site include two twin boom jumbos and one single boom jumbo, with

all three rigs able to drill holes using a 3.7 m steel. For the purposes of scheduling the underground

mining operations it was decided that each jumbo would be able to take a mining cut to a depth of

3.3 m.

Blasting of the rock will be done primarily using ANFO as the main explosive. In development cuts,

reduced charging will be used in the perimeter holes of each cut to minimise the impact of the

explosives on the drive walls and backs.

On completion of drill and blast activities, the installation of weld mesh and steel bolts or split sets

will be used to provide ground control for the drive walls and backs. Ground support work will be

predominantly done using the single boom jumbo, with the twin boom jumbos bolting and meshing

as required.

16.3.3. ORE EXTRACTION

Extraction of the ore and waste material after blasting will be completed using standard

underground loading and hauling practices. Ore material will be removed from the face of the drive

by an LHD (load –haul-dump unit) to stockpiles located on the level. The ore is then loaded onto

haul trucks and transported to the mill processing area on the surface. Waste material will be

removed from the underground workings in the same manner as the ore. The waste material used

for the production of CAF or CRF will be transported to surface where it will be screened or crushed

to provide the appropriately sized material for each of the filler types. Where RF is to be used in the

underground workings, this material will be transported to stockpiles in strategic locations

underground for use as backfill on levels where ore mining has been completed.


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16.3.4. BACKFILLING

Cut and fill mining methods rely heavily on the use of backfill, with the mining process using backfill

material to provide stable material for either the floor (overhand cut and fill) or the backs

(underhand cut and fill) or walls for the extraction of adjacent stoping drives. Backfilling operations

for the Dikulushi underground will use three different types of back fill for the safe extraction of the

orebody.

ROCK FILL (RF)

The use of RF within the mining schedule has occurred where all of the mining activities beside or

below the location backfill have ceased, and the RF is used to provide the floor for extraction of the

level directly above. This fill method is used extensively in certain sections of the orebody where a

single drive is required to extract the ore from the level, and mining activities below the drive have

been completed. Some levels use a combination of CRF and RF. RF is significantly cheaper than both

the CRF and CAF backfilling methods, as no secondary screening or crushing is required and the RF

can be directly hauled between locations with material dumped in a stockpile close to the backfill

location.

RF DESIGN AND PLACEMENT

Placement of the RF will be done using underground LHD units, with the rock placed in the each of

the drives working from the furthest end of the drive back towards the access. In addition to moving

the RF to the fill location the LHD unit can be used to push up and pack in the RF (commonly by using

a ‘rammer-jammer’ attachment on the bucket, as shown in Figure 16.6.

Figure 16.6 LHD loader with ‘rammer-jammer attachment

CEMENTED ROCK FILL (CRF)

Selection of CRF as the primary method for backfilling stopes was done on the basis that materials

required for the process were easily available, with waste development providing the rock fill and

cement being easily imported to site with minimal technical or specialised gear required to produce

the fill. The process for creating CRF involves the mixing of cement, water and rock from waste

headings using a LHD unit in a stockpile located close to the area to be backfilled.


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Once each of the ingredients has been mixed the LHD unit then transports the mix from the

stockpile and deposits it in the area being backfilled in a layered approach. A half to one meter layer

of CRF material is deposited onto the floor of the drive. The LHD unit then mixes more CRF, which is

deposited on top of the initial layer. As the LHD unit deposits the next layer of CRF it compacts the

layer below, helping to improve the strength of the CRF by removing any voids.

To maximise the fill in the drive and minimise any voids between levels, the CRF in the upper

sections of drive should be pushed up and packed in with the loader (commonly by using a ‘rammer-

jammer’ attachment on the bucket, Figure 16.6).

To ensure there is sufficient supply of rock material, additional waste will be sourced from the open

pit stockpiles. On completion of open pit mining operations, material in the pit was blasted, but not

removed; this has been identified as an alternate source of rock fill, should the ground waste

development not provide the volumes required to achieve the underground mining production.

CRF DESIGN AND PLACEMENT

CRF has been designed conservatively at a minimum of 10% cement, and rock fragments of 300 mm

diameter or less. A water to cement ratio is designed at 0.45, i.e. for a mix of 1 tonne cement and 9

tonnes rock (10 tonnes total), 450 litres of water is required (Table 16.1).

Table 16.1 CRF Specifications

CRF Specifications

Cement content ≥10%

Rock fragments ≤ 300mm

Water:Cement ratio 0.45

Curing time until full strength is obtained 28 days

The floor is to be covered in steel mesh, to be anchored into the sidewalls using resin encapsulated

rebar bolts (post-tensioned) at a spacing of 1.2 – 1.5 m and where mesh sheets overlap. The mesh

sheets are also to be shackled together along the overlaps.

The minimum curing time of the CRF prior to the development of adjacent development has been

set at 2 days within the schedule. This has been deemed as a significant time for the CRF to obtain

ample strength to allow mining activities to occur alongside them.

CEMENTED AGGREGATE FILL (CAF)

Cemented Aggregate Fill (CAF) is similar to CRF in design; however, the rock has been sized (and can

include sand) so as to achieve maximum strength. CAF is more expensive to produce than CRF due

to the additional requirement of crushing and screening the rock prior to use (as CAF) as backfill. It

is intended that CAF will be used on any levels that will subsequently form the backs for a level

directly below.

CAF DESIGN AND PLACEMENT

CAF has been designed conservatively at a minimum of 10% cement and rock fragments of 25 mm

diameter or less. The water to cement ratio is designed at 0.45, i.e. the CRF can be mixed at the


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surface using AGI trucks or by LHD using a mix of 1 tonne cement and 9 tonnes rock (10 tonnes

total), 450 litres of water is required (Table 16.2).

Table 16.2 CAF Specifications

CRF Specifications

Cement content ≥10%

Rock fragments ≤ 25mm

Water : cement ratio 0.45

Curing time until full strength is obtained 28 days

The floor is to be covered in steel mesh, to be anchored into the sidewalls using resin encapsulated

rebar bolts (post-tensioned) at a spacing of 1.2 – 1.5 m and where mesh sheets overlap. The mesh

sheets are also to be shackled together along the overlaps.

The minimum curing time of the CRF prior to the development of adjacent development has been

set at 2 days within the schedule. This is being deemed as significant time for the CRF to obtain

ample strength to allow mining activities to occur alongside them.

CEMENTED PASTE FILL

Testwork is being conducted by MWL to identify whether the tails material produced from the

production of the copper concentrate is suitable for use as paste backfill in the underground

operations. Initial results of the testwork have indicated that there is a high likelihood of this

material being suitable for backfilling operations. Should the testwork prove successful paste fill

could be used as an alternative to either CRF or CAF. The advantage of paste fill is its ability to

provide a better filling ratio in the backfilled drives, with minimal void space remaining after the

completion of the backfilling process. The use of paste fill is highly unlikely in the current reserve

mining plan due to the short mine life based on the Measured and Indicated Reserves. Should

Inferred material within the geological reserve model be able to be upgraded to higher confidence

categories, paste fill would provide a viable option for future mining activities.

16.3.5. ACCESSING THE OREBODY & REHABILITATION OF OLD WORKINGS

Initial access to the ore body will be provided by the existing workings left behind by Anvil. This

development includes extensive level development on the 810, 790 and 770 mRLs. The decline has

been developed down to the 750 mRL and includes the initial take-off drive development for level

access. Development of the decline below the 750 mRL will be completed using twin boom

underground development drills, with ground support implemented as per the geotechnical

recommendations from Mike Turner’s report ‘Geotechnical Input for re-opening Dikulushi

Underground’, August 2013.

Dewatering activities have made it easy to access the main decline. Prior to closure of the

underground mine in 2008, M. Turner (Geotechnical Consultant) undertook a detailed inspection of

the underground working areas. He highlighted a number of areas which required remedial support,

and the need for ongoing inspections for mining activities to proceed. In late 2010, MWL asked

Australian Mining Consultants (AMC) to undertake a geotechnical assessment of underground


P a g e | 117

mining options proposed by the previous owners Anvil. As part of this work AMC provided the

following comment: “Rehabilitation requirements below 900mRL are unknown, but for budgeting

purposes, it should be assumed that all of the (ground) support will require replacement.”

In December 2012 an inspection was performed by M. Turner to assess the decline down to the

groundwater level (830 mRL). This concluded that the rock mass and support in the decline was in

good condition, limiting the number of areas requiring support rehabilitation. Some areas would

require bleeding of scats from behind damaged and sub-standard mesh, but most areas could be

made safe by installing additional sheets of mesh over the existing support. Additionally, it was

observed that some areas below 900 mRL had corroded mesh. After testing the mesh strength,

MWL was satisfied that this would still perform as required. As dewatering proceeds, ongoing mesh

testing is planned.

As open pit mining has proceeded, the previously mined crosscuts and ore drives have been

intersected. The installed ground support, which includes galvanised Split Set bolds and mesh, is

exposed on the pit floor. This has allowed close inspection of installed support, and very little

corrosion was consequently noted. This has been consistent when mining down through the various

ore drives. Again, this lends weight to the premise that there has been only minor deterioration of

the installed ground support.

16.4. VENTILATION

16.4.1. PRIMARY VENTILATION

A series of primary ventilation rises will be developed in conjunction with the main decline. These

rises will be mined using hand-held mining equipment, with a rise driven between the return air

drives at a 1.5 m width and then stripped out from the top down to a diameter of 5 m. The rises will

be connected to the decline through a series of small return airway drives strategically positioned at

20 vertical metre intervals. Ventilation of the decline between return airway levels will be provided

through secondary fans in the development phase. A new section of the primary ventilation system

will be developed, and ventilation bulkheads will be used to seal off the return airway drives located

above the lowest return airway drive. This process will then provide primary ventilation to the

bottom of the decline, allowing for further development. Figure 16.7 shows the proposed

underground primary ventilation circuit down to the 520 mRL created by Red Rock Engineering.

Figure 16.8 shows the ventilation development required for mining the existing Measured and

Indicated material in the Mineral Reserve.


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Figure 16.7 Underground primary ventilation circuit (full)

Figure 16.8 Underground primary ventilation circuit required for the extraction of the measured and indicated material only

The ventilation system sets up the return airways from each of the levels up to the 825 mRL, where

the primary fans will be located. These fans will draw air out of the return airway system and

exhaust it into the current open pit, with fresh air being drawn down the decline to complete the

ventilation circuit. Figure 16.9 shows the proposed primary fan location.

Primary Ventilation

Return Air Rises

Primary Ventilation

Return Air Rises


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Figure 16.9 Primary ventilation fan location

16.4.2. SECONDARY VENTILATION

Secondary ventilation will be provided using smaller fans located in the decline, with ventilation bags

extending into the levels and secondary ventilated areas. MWL plans to use the secondary

ventilation fans and the electrical installations that were kept from the original underground

workings completed by Anvil. All this equipment will be refurbished and made fit for purpose prior

to re-installation in the underground workings.

16.5. DEWATERING

When mining the Dikulushi open pit cut back, MWL decided against dewatering from within the

open pit, and instead decided to access the existing decline and use it as a pumping platform. A

110mm diameter line was connected to a small floating pontoon within the decline, and as the

water level receded, the pontoon was shifted down the decline. This strategy successfully

minimised the impact of flood water on open pit mining. In the main decline, dewatering was rapid

until established levels were encountered at 20 m vertical intervals, below 900 mRL. The extensive

development encountered on these levels stored a significant amount of water and consequently

reduced dewatering rates. As the open pit was mined, depressurisation holes were drilled at 10m

vertical intervals, and 20 m horizontally apart within the walls. Significant volumes of groundwater

were encountered in the eastern and western walls of the pit, which was allowed to flow onto the

pit floor, eventually making its way into the underground workings. Prior to the 2012-13 wet

season, (October 2012), the groundwater level had been lowered to the 820 mRL. During the wet

season, a number of heavy downpours dramatically increased water flow into the underground,

such that by the end of April 2013, the water level had risen to 849.5 mRL. Dewatering activities at

the time were limited by the equipment available at the time of the rain events.

The commencement underground mining will depend upon the success in reducing the water level

to below the planned working areas. For initial mining, the water level will need to be below the 770

mRL level. The near term mining will need to effectively dewater all underground development

Underground as built

design information

Primary Ventilation

Fan Location 825 mRL


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down to the 750 mRL. For long term mining, ground water inflow will have to be carefully managed

to minimise its impact on mining activities.

Long-term dewatering of the underground mining operation would use the existing dewatering

sumps and pumping locations originally set up. Dewatering stations will be set up as the water level

is lowered throughout the existing mine workings. It is anticipated that dewatering equipment left

behind by Anvil, once refurbished, will be suitable to control the water entering underground

workings. In addition to the existing pumping locations and sumps, new sumps and pumping

locations will be developed as the decline is advanced down to lower regions of the orebody. Figure

16.10 and Figure 16.11 show the existing underground dewatering infrastructure locations and

proposed future dewatering locations respectively. The existing dewatering infrastructure

implemented by Anvil had a sump located along the decline, with vertical distances between 60 and

75 m between stations. All dewatering pipelines were run between pumping stations using the main

decline.

Figure 16.10 Existing underground dewatering infrastructure locations

Figure 16.11 Proposed underground dewatering infrastructure locations

New Sump Location


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16.6. MINING EQUIPMENT

With the purchase of Anvil, MWL acquired all the mining equipment associated with the mining

operation. This included all of the underground equipment previously used by Anvil in the

underground operations closed in 2008. MWL plans to refurbish this equipment and use it as the

underground mining fleet for extraction of the underground Mineral Reserves. Table 16.3 is a list of

the major pieces of mining equipment available. Additional mining equipment has also been

retained from the previous underground mining activities, including dewatering pumps, ventilation

fans, starter boxes, electrical supply and distribution equipment that was recovered as part of the

abandonment process.

Table 16.3 Dikulushi production mining equipment at site from previous mining activities

Equipment type Make Model Description

Development Drill Rig Sandvik Axera 6 Twin Boom Jumbo

Development Drill Rig Sandvik Axera 6 Twin Boom Jumbo

Development Drill Rig Sandvik Axera 5 Single Boom Jumbo

Haulage Truck Sandvik EJC533 30 t dump truck

Haulage Truck Sandvik EJC533 30 t dump truck

Haulage Truck Atlas Copco MT440 30 t dump truck

LHD Sandvik Toro1400 Underground Loader

Diamond Drill Boart Longyear LM75 Diamond Drill Rig

Diamond drill Kempe Diamond drill rig

Integrated tool carrier Caterpillar 924 Integrated tool corner

The composition of the mining fleet required for extraction of the underground Mineral Reserve is

shown in Table 16.4. In the equipment listed, the following items are still required to be hired or

purchased

one LHD

two light vehicles - will be taken from existing LVs on site

one charge up vehicle

six air leg drills and

a wire line scraper.

Due to the short mine life based on the Measured and Indicated Mineral Resource, the purchase of

major equipment such as the LHD has not been included as part of the capital cost. It is expected

that these items will be obtained on a hire arrangement, with the cost included as part of the main

operating costs.

It is expected that parts of the fleet will have low utilisation due to the small tonnages being

produced from the underground mining operations, providing ample coverage for breakdowns. The

additional pieces will still be required, as they will provide the operation with the flexibility to

increase mining rates as required.


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Table 16.4 Major mining fleet and equipment required for the extraction of the Dikulushi underground Mineral Reserves

Mining fleet Number

LHD 2

Truck 3

Jumbo twin 2

Jumbo single 1

Integrated Tool Carrier 1

Charge up vehicle 1

PC 1

Ute 2

Air leg drills 6

Wire Line Scraper 1

16.6.1. MINE DEVELOPMENT

The development of the underground operations will be completed using traditional drill and blast

methods utilising two twin-boom jumbos. A single-boom jumbo is to provide rock support and

rehabilitation capabilities, using weldmesh and rock bolts for the bulk of the mining development, as

per the geotechnical guidelines provided by Mike Turner. In addition to the standard ground

support outlined in the report, cable bolts will be installed at intersections and any areas of poor

ground encountered. Table 16.5 is a list of the horizontal development design parameters, including

the development location/type, width, height, back profile, ground support method and gradient.

Table 16.5 Underground horizontal development design parameters

Development type Width

(m)

Height

(m) Profile

Ground

support

method

Gradient

Decline 5.5 5.5 Arched Mesh & Bolt 1 in 7

Decline Stockpile 5.5 5.5 Arched Mesh & Bolt 1 in 50

Level Access 5.0 5.0 Arched Mesh & Bolt 1 in 100

Level Stockpile 5.0 5.0 Arched Mesh & Bolt 1 in 100

Return Airway Drive 5.0 5.0 Arched Mesh & Bolt 1 in 50

Cut and Fill Ramps 4.5 5.0 Arched Mesh & Bolt 1 in 6

Ore Drives 4.5 5.0 Square Mesh & Bolt 1 in100

Vertical development to be completed in the underground will be conducted using hand-held rising

methods. Escape way rises will be developed in a single pass from bottom to top. Other vertical

development, such as vent rises, will be developed initially with a single rise of 1.5 m in diameter

developed from bottom to top. Stripping of the rise will then take place working from top to

bottom, with all blast material scraped into the centre of the rise where it will fall down to the lower

level for removal. Installation of split sets and mesh will be done as required in the larger

development openings. Table 16.6 is a list of the expected vertical development required in the

mining operation.


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Table 16.6 Underground vertical development design parameters

Development type Diameter

(m)

Length

(m) Profile

Ground

support

method

Angle

(deg)

Vent Rise 5.0 15 to 20 Round N/A 60

Escape Way Rise 1.5 15 to 20 Round As Required 60

16.6.2. MINING SCHEDULE

For the purposes of determining reserves for the Dikulushi underground mining operation several

schedules were created. Each of the schedules created was produced in Enhanced Production

Scheduler (EPS) using information from the reserve block model and underground designs provided

by MWL.

Initial scheduling of the underground operations involved creating a Measured, Indicated and

Inferred schedule that would be the basis of identifying the economic prospects of the underground

mining operation. This schedule was created to identify the probability of the mining operation

having a longer life than would have been indicated by just evaluating the Measured and Indicated

ore reserve only. This approach was taken due to the limited amount of Measured and Indicated ore

reserve remaining after the completion of the open pit mining activities.

The initial Measured Indicated and Inferred schedule was evaluated by MWL using an existing

financial model to identify the project’s potential viability and to understand the impact of the

operation within the company’s portfolio of mining operations. From the information output from

MWL’s financial model, the following costs were used as a benchmark for evaluating the Measured

and Indicated tonnes contained within the Initial schedule with

an operational mining cost of $100 per ore tonne mined,

a processing cost of $55 per ore tonne and

an operational overhead costs of $68.20 per tonne.

For the process of identifying the viability of the Measured and Indicated ore tonnes contained

within the underground design, a total mining and processing cost per ore tonne was then estimated

at $223.20. This cost was estimated for an underground operation with an ore mining and

processing production rate of 183,291t per annum. In addition to the ore mining cost, a copper

price of $6,800 per tonne and silver price of $20 per ounce were then used to determine the

revenue from the contained Measured and indicated copper tonnes and silver ounces.

The tonnes and grade for each of the levels were exported from EPS into a spreadsheet where a high

level economic evaluation was performed. For the economic evaluation of the Measured and

Indicated ore tonnes, each level was split into two sections - east and west of the level access.

The revenue for each section was calculated by taking the Measured and Indicated copper tonnes

and silver ounces contained in the mined material multiplied by the processing recoveries and

product prices. Each section was then evaluated by subtracting the mining cost from the section

from the revenue obtained from the sale of the measured and indicated copper tonnes and silver


P a g e | 124

ounces to determine if it had a positive revenue. Each level that had a positive revenue was

included as part of the Mineral Reserve schedule.

From the evaluation process nine levels were identified as having positive revenues. Seven of the

levels included both the drives to the East and West of the access, with two levels having

development east of the access only. A second EPS schedule was set up to provide a practical

mining schedule for the basis of the underground reserves.

A second mining schedule was created to provide a new mining sequence, incorporating the

development of capital mining development to provide

primary ventilation of the mining operation

access to the design ore drives from the existing underground workings

development of the ore drives

the use of backfill.

This schedule excluded the higher risk ore contained in the crown pillar zone between the 805 and

825 mRL’s. The Measured and Indicated mining sequence created in EPS mined 121,000 ore tonnes

at 5.4% copper over a period of 19 months. This timeframe allows for one month lead for

rehabilitation of the existing underground workings and development of rises to establish primary

ventilation along with one extra month at the end of the schedule to complete ore backfilling

operations. The mining costs used in the schedule were a combination of time variables such as

labour and administration costs which are independent of production activities, and direct

production costs which are driven by the day-to-day production of the mining and processing

operations. All costs used in the schedule were provided by MWL, with direct production costs

created from first principle methods and the overhead costs calculated using historic site costs.

The final reserves schedule included the extraction of the crown pillar tonnes from the 805 to the

825 mRL’s. The economic selection of the drives to be mined as part of the Mineral Reserve was

completed using the same process as the second schedule. In addition an economic evaluation

process for the ore drives was evaluated, based on their interaction with the existing Dikulushi open

pit. Where the underground development could have a significant impact on the stability of the

open pit shell, these ore tonnes were removed.

The Measured and Indicated resource mining sequence created in EPS mined 173,000 ore tonnes at

5.2% copper over a period of 20 months. The extraction of the additional tonnes in the crown pillar

area was scheduled in parallel with the ore tonnes produced from a second mining schedule being

created.

CONSTRAINTS

The Measured and Indicated Mineral Reserve mining schedule produced in this report has tried to

replicate the same parameters as outlined in the Measured Indicated and Inferred schedule

produced for MWL. Mining activities were restricted to levels that provided a positive economic

evaluation as described, with the mining sequence altered to minimise the mining timeframe

required for the extraction of the ore. All drives were reduced in length where the copper grade fell


P a g e | 125

below 1.5% towards the end of the drive. Ore drive sections where the grade fell below 1.5% but

then rose above 1.5% further down have been retained in the schedule, and were treated as

marginal ore, requiring that only the processing costs were covered by putting the ore tonnes

through the processing plant as a tonnes would need to be mined regardless of grade to access the

high-grade sections of the orebody.

To ensure that the full cost of mining for each ore drive has been included as part of the economic

evaluation, the Inferred and unclassified ore tonnes mined have not been included as part of the

cost calculation process. The grade component from the Inferred and unclassified tonnes was not

included as part of the revenue from the drive, diluting the grade from the Measured and Indicated

tonnes mined.

The production drilling rates that are used, are the maximum scheduled advance per month rate, as

Jumbo development is the critical limiting factor in the schedule. Twin-boom Jumbo lineal advance

rates have been scheduled at 193m per month per jumbo, or a total of 386 m per month for both

twin-boom jumbos. Waste development has been scheduled at 100 m per month. Development

rates for ore drives have been scheduled at the following rates:

52 m per month in the western end of the orebody

65 m per month in the central and eastern areas of the orebody

91 m per month for capital development

Development rates used are outlined in greater detail in section 16.6.4.

DEVELOPMENT AND LEVEL DESIGN

The designs used in the schedule were in addition to work already completed by Red Rock

Engineering, which consisted of designs for capital development and an ore body wireframe shape

(stope6.dtm). Optiro has produced level designs in Datamine along the width of this stope shape

using 5m vertical intervals and drive dimensions of 4.5 mW x 5.0 mH. Levels have then been broken

into approximately 10m sections to allow more granular scheduling.

The level development design sections were then imported into Mine 5D Planner, where they were

evaluated against the block model and each section was assigned the corresponding average grade,

tonnes and material properties of the model cells contained within it. Additional properties were

also added such as development length and segment identification fields.

ORE LOSS AND DILUTION

For scheduling purposes, mining ore loss has been estimated at 5%. Ore loss will primarily occur as a

result of gaps left in the backfill process, caused by the inability to properly level off the top of

backfill as it is being placed in the drive. This means that when using the backfill as the floor for the

next level the fill will generally have a rounded profile, and this will result in some material being lost

to fill these corners as the next level is developed (see Figure 16.12). There will also be a small

amount of ore loss from general activities such as rehandle and material movement.


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Figure 16.12 Ore loss due to gaps left in the backfilling process

Mining dilution for the project has been estimated at 8%, which is the average dilution calculated for

the various drive dimensions. Dilution was estimated by allowing a 200mm dilution skin on the walls

of development drives, which has then been used to determine the percentage dilution using the

cross-sectional area of the drives. Drive profile calculations are outlined in Table 16.7. These

estimations assume a square cross-section for calculation purposes.

Table 16.7 Mining dilution

Development Width

(m) Height

(m) Drive

Area (m2) Dilution

width (m)

Dilution skin area

(m2)

Dilution (%)

Ore

Ore Drive 4.50 5.00 22.50 0.20 2.00 8.9

Waste

Decline 5.50 5.50 30.25 0.20 2.20 7.3

Decline Stockpile 5.50 5.50 30.25 0.20 2.20 7.3

Level Access 5.00 5.00 25.00 0.20 2.00 8.0

Level Stockpile 5.00 5.00 25.00 0.20 2.00 8.0

Return Airway Drive 5.00 5.00 25.00 0.20 2.00 8.0

Cut and Fill Ramp 4.50 5.00 22.50 0.20 2.00 8.9

Average 5.08 5.17 26.26 0.20 2.07 7.9

Ore loss has been applied to the schedule by reducing the contained metal for each segment by 5%.

Dilution has been applied to the in situ tonnes of the blocks, multiplying them by 1.08 and then

subtracting the ore loss material. This only occurs in areas producing ore, and is limited to the

Measured and Indicated tonnes only. Updated metal grades (including both ore loss and dilution)

are then calculated from the ore-loss applied contained metal and the new diluted tonnes. For this

schedule ore loss has been assumed for both copper and silver. Waste tonnes have been calculated

for the purpose of financial evaluation of the schedule, and have been determined by simply taking

the total tonnes for a segment and subtracting the ore tonnes.


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BACKFILL

Backfilling has been scheduled by calculating the mined volume of the drives and allowing for an

80% fill factor due to the difficulty of pushing material all the way up to the backs. Backfill has been

scheduled at a rate of 1,008 m3 per week per machine for all fill types, and minimum curing times

have been scheduled at 28 days for mining underneath fill and 2 days for mining alongside fill. A

density of 2.5 t/m3 has been assumed for backfill, which is conservatively high. Backfill sequencing

has allowed for 5 m pillars between active levels.

HAULAGE CALCULATIONS

The schedule has included production tonnes-km figures, which are a metric used to represent

haulage in terms of both the total tonnes and the total required haul distance for a given schedule

block. Surface haul distances were calculated by measuring the haul distance from the waste dumps

and ROM pad from the portal. Underground haul distances were calculated using differences in

vertical distance between the portal (1015 mRL) and the schedule level RL, and multiplying it by 7 to

estimate a decline distance. Total haul distances (in km) were then multiplied by the total tonnes

mined in each schedule block to calculate the haulage tonnes-km value.

SCHEDULE SEQUENCE

The Mineral Reserve schedule utilises the existing capital development completed in previous

underground mining activities. The mining activities that are required before a stoping operation

can re-commence are the development of ventilation and escape way rises in the upper levels of the

mine that were not previously completed. The development of the ventilation and escape way rise

system is to be completed within the first 4 months of the mining operation’s re-commencement.

This opens the opportunity to develop the maximum number of all drives from month 5 in the

schedule onwards.

Level development in the first 3 months of mining production is concentrated in the 800, 790, 765

levels. Towards the end of the quarter the 755 mRL opens up for mining activities with backfilling

operations. Backfilling operations are initially concentrated in the eastern section on the 790 mRL

for the first half of the quarter, with the 765, 770, and 800 RL backfilling activities commencing

towards the end.

Mining activities in the second quarter are scheduled to continue in the 790, 765 mRL’s. New mining

activity then commences on the 770, 780 and 750 mRL’s. Backfilling activities continue in the 790

and 800 mRL’s with backfilling activities starting in the 770 and 755 mRL’s. Mining activities in both

the 755 and 790 mRL’s is completed during this period.

Quarter three mining activities concentrate around the 770, 780, 795 mRL’s with backfilling activities

occurring on the same three levels. Mining of the 770 and 780 mRL’s is completed during this

period.

Quarter four mining activities are concentrated around the 795, 785 and 775 mRL’s and backfilling

activities are concentrated around the 780 and 795 levels, with backfilling activities also occurring in

the 785 and 775 levels towards the end of the period.


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The remaining mining activities for quarters five and six are concentrated in the 785 and 775 mRL’s,

extracting the remainder of the current Measured and Indicated material. Due to the requirement

for backfill to minimise the open stoping widths, mining and production significantly reduces.

Backfilling activities during this period include the 785 and 755 mRL’s along with the 795 mRL.

Mining activities are completed within the first month of quarter six, with backfilling activities

finishing midway through the third month of the quarter.

Figure 16.13 is a pictorial representation of the Measured and Indicated Mineral Reserve extraction

sequence as described.

Figure 16.13 Ore level schedule, by quarter

Table 16.8 details of the planned mining production physicals on a quarterly basis.

Table 16.8 Underground mine production physicals

16.6.3. MINING SHIFTS

Mining personnel shifts will be split up into three 8 hour panels; a day shift, afternoon shift and night

shift (outlined in Table 16.9). Day shift commences at 07H00 and shift changeover for the day shift

to afternoon shift will occur at 15H00. This is timed to coincide with the primary blasting time, and

afternoon shift start time will be dependent on re-entry periods following blasting. Shift changeover


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from the afternoon to the nightshift will occur at 23H00. Work rosters will be a 2 day shift / 2

afternoon shift / 2 night shift / 2 off basis.

Table 16.9 Work shifts

Shift Start time Finish time Duration

(hrs)

Day 07H00 15H00 8

Afternoon 15H00 23H00 8

Night 23H00 07H00 8

Project Management, Technical Services, and Support Personnel will generally work day shift only,

covering the day and part of the afternoon mining shifts. Table 16.10 to Table 16.12 shows the

estimated total number of personnel required for the site’s management, technical services and

operational support functions.

Table 16.10 Operational Management Labour

Operational Management Quota

Project Manager 1

Maintenance Manager 1

Electrical Supervisor 1

Senior Auto Electrician 2

Maintenance Crew Leader/ Trainer 3

Op. Crew Leader / Trainer 3

Table 16.11 Technical Services labour

Technical Services Quota

Technical Manager 1

Mine Survey 2

Mine Geology 2

Geotechnical Engineer 2

Mine Engineering 2


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Table 16.12 Support functions labour

Labour Quota

HSE Manager 1

Onsite Medical Doctor 2

Senior Environmental Officer 2

Onsite Medical Nurse 6

Environmental Technician 3

General Clerk 3

Ambulance Driver 3

General Driver 3

General Assistant 6

Table 16.13 and Table 16.14 detail the shift personnel numbers required for the operation of each

shift. The total numbers of personnel required to maintain the mining operations will be the

number of personnel per shift multiplied by the number of shifts. Where the manning number is

accompanied by an asterix, this means that this role is only filled during the day and afternoon shift,

with no personnel rostered into this position during the nightshift.

Table 16.13 Labour requirements: underground operations

Underground Operations Quota

Mine Foreman 1*

Shift Supervisor 1

Develop Drill Crew 2

Rehab Crew 1

General Service Crew 3

Material Haul 3

Material Load 2

Charge Crew 2

Air leg Drillers 2*

Backfill Crew 2


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Table 16.14 Underground Workshop personnel

Labour – Underground Workshop Quota

Workshop Foreman 1*

Crew Supervisor 1

Develop Drill Fitter 1

Workshop Fitter 2

UG Fitter 2

Trade Assistants 2

Senior Electrician 1

Mine Electrician 1

16.6.4. DEVELOPMENT / STOPING RATES

Development rates are outlined in Table 16.15, showing the estimated development turnaround,

length and maximum number of cuts available to be taken in each type of heading per week. All the

drive profiles can be found previously in Table 16.5 and Table 16.6.

Table 16.15 Jumbo/production drill rates by development type

Drive type Cut Length

(m) Turn Around

Maximum cuts per week

Decline 3.3 1.5 shifts 7

Decline Stockpile 3.3 1.5 shifts 7

Level Access 3.3 1.5 shifts 7

Level Stockpile 3.3 1.5 shifts 7

Return Airway Drive 3.3 1.5 shifts 7

Cut and Fill Ramps 3.3 1.5 shifts 5

Ore Drives 3.3 1.5 shifts 5

Table 16.16 details the individual Jumbo production rates applied in the scheduling process; these

rates have been created by applying basic mining principles on a conservative production rate. The

twin-boom development rate (193.1 m per month) has been applied to the schedule, and the single-

boom jumbo rates provided are intended as back-up rates only. Monthly development for 2 twin-

boom jumbos has been scheduled at a lineal rate of advance of 386.1 m per month, or 117 cuts

(shown in Table 16.17). The single-boom jumbo is assumed to be used primarily for cable bolting

and rehabilitation work.


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Table 16.16 Jumbo/production drill rates by individual machine

Jumbo

advance

Cut

length

(m)

Cuts

per

shift

No. of cuts

per week

Dist.

advanced per

week (m)

No. of cuts

per month

Dev. adv. per

month (m)

Dev.

status

Twin-boom 3.3 0.64 13.5 44.6 59 193.1 Primary

Single-boom 3.3 0.31 6.5 21.5 42 139.4 Back up

Table 16.17 Jumbo/production drill rates by fleet

Monthly

Development Units

Cut

length

(m)

Cuts

per

Shift

No. of

cuts per

wk.

Dist.

advanced

per wk.(m)

No. of

cuts per

month

Dev. adv.

per month

(m)

Development

Status

Twin-boom 2 3.3 1.29 27 89.1 117 386.1 Primary

Single-boom 1 3.3 0.31 6.5 21.5 42 139.4 Back up

16.6.5. AIR LEG DEVELOPMENT RATES

Development work using air leg miners will occur on an eight hour day shift only, as the work

conducted by these workers is regarded as high risk. Table 16.18 outlines the production rates that

have been used in the schedule. Air leg/rise mining is a critical part of the mining plan, as it provides

the return airways for the primary ventilation circuit and the alternate secondary escape a path for

an emergency.

Development of the air leg rises for escape ways has been split up into production rates for single

and double rises. The development rate of rises varies with the air leg miners’ access to multiple

headings. For a single rise being developed by itself the advance is 1.5 m per shift, if there are two

rises close together this rate is able to be doubled due to the availability of equipment.

This scheme was applied for the development of the ventilation rises in the schedule and involves air

leg miners initially developing a rise from the lower level of the ventilation rise to the upper level.

Once the rises can be completed, the air leg miner then proceeds to strip the surrounding parameter

of the final rise diameter into the hole created by the initial rise. The process used by the air leg

miner involves drilling and firing the material to be stripped into the hole during one shift, with the

following shift required to clean out the fired material from the stripping, and installation of

appropriate ground support around the rise as it is developed down. It is expected that the

development of a 20 m rise will take one air leg miner approximately 42 days to complete the

development from start to finish.


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Table 16.18 Air Leg development

Development type Length

advance (m)

Turn around Advance per

week (m)

Rise Development

Single 1.5 per shift 7

Double (side by side) 3 per shift 7

Return Air Way stripping

Drill and Fire 1.5 per shift 3.5

Bog and Ground support 1.5 per shift 3.5

16.7. GEOTECHNICAL

16.7.1. DATA

A Geotechnical study into re-opening the underground operation has been performed by Turner

Mining and Geotechnical (Turner, 2013). The study investigated geotechnical aspects of mining from

the pit bottom around 825mRL to 500mRL, and was based on the following data:

Diamond drill core logging databases (31 holes, Anvil and MWL).

Underground mapping data (7 sites, personal data, undertaken for Anvil).

Surpac files of current open pit and underground excavations (MWL).

Surpac files of planned open pit and underground excavations (MWL and Red Rock

Engineering Pty Ltd).

Previous geotechnical reports (Turner Mining and Geotechnical (2008(a), 2008(b), 2012 and

AMC Consultants (2004, 2011)).

Observations made during multiple site visits (2003 to December 2012).

16.7.2. GEOTECHNICAL DOMAINS

Rock Quality classification has been conducted by Turner Mining and Geotechnical, with the

following geotechnical domains identified:

The orebody ranges from Extremely Poor to Fair, with the majority classed as Poor.

The footwall rockmass averages Fair.

Hangingwall ranges from Extremely Poor to Fair, with the majority classified as Poor.

The contours of Q for the orebody (Figure 16.14) show the very poor ground above the750mRL and

west of 50300mE. This poor ground zone in the orebody is critically important to manage as it

indicates severe ground control problems could be encountered when trying to extract ore out of

this zone

The contours of Q (rock quality) for the footwall (Figure 16.15) show 2 zones of very poor ground,

centred around the 725mRL at 50125mE; and 525mRL at 50280mE. These two zones indicate that

even though split sets and mesh might still be appropriate, there could be a need for shorter

cuts/round lengths and secondary installation of grouted bolts.


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The contours of Q for the hanging wall (Figure 16.16) show a zone of very poor ground centred

around the 725mRL at 50130mE. This poor hanging wall zone eliminated the use of any longhole

stoping or benching in this zone.

Figure 16.14 Dikulushi orebody rock quality, Q (Turner, 2013)

Figure 16.15 Dikulushi footwall rock quality, Q (Turner, 2013)


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Figure 16.16 Dikulushi hanging wall rock quality, Q (Turner, 2013)

16.7.3. POTENTIAL FAILURES

From the previous mapping data gathered during mining activities in the 810 mRL in 2008 following

the fall of ground in 810 mRL W2 drive. Unwedge analysis was performed to evaluate the potential

for wedge instability in the walls and backs of excavations of the underground excavations. From

this analysis it was identified that the use of solid steel bolts compared to standard spit sets in

ground support activities provided an increase the factor of safety, reducing the likelihood of an

wedge failure occurring. Recommendations from the report by Turner ‘geotechnical input for re-

opening of the Dikulushi underground’ suggest the use of solid steel rock bolts to be used during

development of levels in the orebody where poor ground conditions have been identified (above the

755 mRL and West of the 50130 m E).

16.7.4. MAPPING, MONITORING AND ADDITIONAL DATA

Mapping and monitoring of the underground drive development will form a significant part of the

ground support risk mitigation activities for the underground development, especially where ore

drives are located in areas of poor rock mass quality.

16.8. GROUND SUPPORT REQUIREMENTS

The orebody rock mass quality in the poor ground section from 810 to 780 mRL west of 50130 mE

ranges from Very Poor to Extremely Poor and the hanging wall is also Very Poor. Only small voids

will remain stable in this zone, and will require very intensive support, including solid steel rockbolts,


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solid steel spiling bars and weldmesh or fibrecrete (Figure 16.17). Even with such support there will

still be a risk of collapse that will need to be managed by operations.

Figure 16.17 Dikulushi rock reinforcement chart (Turner, 2013)

Detailed pre-scheduling of drives is also essential to ensure sufficient working places are available,

taking into account the fill curing constraint. The strength and curing time need to correlate with

the scheduled extraction sequence to ensure sufficient ore is produced.

Testing of the different fill mixtures with varying cement percentages is another essential function

prior to the introduction of this method. The fill will need to have a guaranteed strength of at least

10 MPa before excavation can proceed under the fill. It is also common practice to increase the

cement content above that indicated from tests by up to 50% to cater for poor mix control.


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Stope drifts should be maintained at no greater than 4.5 m width and split sets and mesh

used as support

quality control on cement percentage and mixing is essential

the extraction sequencing of adjacent drifts in wide areas will be critical.

2.4 m solid steel bolts should be inserted into 1.2 m holes 0.3 m above the floor at 1.2 m spacing

along each wall of the drive prior to fill being introduced in order to improve the tensile strength and

wall adhesion in the critical lower corners of the filled drifts. Mesh should also be left at the same

height across the width of the drifts if paste fill is used.

16.8.1. SPLIT SETS

Most ore drives may be supported with split sets and mesh. Split sets (friction stabilisers) should be

galvanised, 2.4 m long, 46 mm diameter and of the type installed by jumbos. Hole diameter control

is an essential part of split set installation and bit sizes must be checked regularly to ensure that the

hole diameter is 44 to 45 mm. 0.9 m, SS39 stubby split sets may be used for mesh overlaps.

16.8.2. SOLID STEEL ROCKBOLTS

Solid steel rockbolts that are immediately active (no delay for grout curing) will be required for any

development in the extremely poor ground in the orebody (above 775 mRL and West of 50130 mE).

Suitable solid steel bolts include:

20 mm Posimix bolts

20 mm CT-Bolts

20 mm Gemini Bolts (South African version of the CT-Bolt)

20 mm MD Bolt (combination split set and mechanical wedge).

These are either anchored using a mechanical shell/wedge or with resin, and can be used to install

mesh to the face. Plain solid steel bars are also suitable for use as spiling bars in extremely weak

ground to stabilise the backs ahead of the drive.

16.8.3. CABLE BOLTS

All intersections should be cable bolted unless a geotechnical engineer is on site to map the

intersection and determine if it is stable, without a potential for wedge failure. Cable bolts should

consist of fully grouted, twin-strand 15.2 mm, plated and tensioned units on a 2.5 m spacing.

Historically cable bolts have never been installed correctly at Dikulushi and suitable equipment

should be purchased. Effective training and supervision will also be essential.

16.8.4. SHOTCRETE

Fibrecrete may be used instead of mesh but the logistics of maintaining an operational shotcrete

fleet at Dikulushi would probably preclude this option. Shotcrete with fibres would be useful for

intersections of extremely weak ground, but these are only expected in the orebody, and if the

ground is that weak there will be subsequent stope stability issues.


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16.8.5. GEOTECHNICAL FILL REVIEW

Brief comments on fill strength and sizing have been recommended by Turner Mining and

Geotechnical Pty Ltd in the 2013 report with additional comments provided by Peter O’Brian and

Associates Pty Ltd.

The 4 main types of fill considered for the operations in both reports were;

rock fill (RF)

cemented rock fill (CRF)

cemented aggregate fill (CAF) and

cemented paste fill (Paste)

Rock fill was recommended for overhand cut and fill stoping .It was recommended that the use of

standard underground development waste was sufficient, with the main function of the rock fill

being to provide a working floor for the next mining levels/lift.

Cemented rock fill (CRF) was recommended to be use in open stoping, bench stoping and cut and fill

mining activities. The percentage of cement recommended was 10% and this is significant more

than stand a cut and fill operations, which generally use 5 to 6% cement. In addition to the use of

rock fill it was also recommended that material no greater than 300 mm be used.

CAF was also recommended to be used in open stoping, bench stoping, underhanded cut and fill or

standard cut and fill mining activities.

In addition to the uses of CRF and CAF a review of the tailings from the Dikulushi processing plant

was also conducted to identify there suitability for their uses in cemented paste fill. This review

identified that the tailings from the plant would be suitable for the creation of paste fill for the

underground mining operations, the tailings from the processing plant had an even size distribution

which will lead to an increase in strength and fast curing times.

The use of paste fill requires the construction of both a paste fill plant and underground delivery

system either through pipework located in the decline or a series of boreholes to deliver fill to each

of the levels. Additionally barricades need to be constructed at the end of the voids being filled with

paste fill to contain the fill wallet sets. Construction of barricades could be undertaken using waste

material and hand packed cemented fill bags sealing the opening to the backs of the drive. Due to

the limited height of ore drives/Stopes in the underground working it is unlikely there will be a need

for engineered barricades to contain the paste fill.

16.8.6. CRF MIXING

Cemented Rock Fill (CRF) consists of waste rock mixed with cement. The source of the waste rock

should be unweathered and without an excess of large rocks or fine material. Underground

development waste is suitable but waste rock from the open pit waste dumps is not suitable due to

the much larger fragment size.

The method of mixing cement into the rock is critical and has a major impact on the cement content.


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Mixing methods can include:

Spray mixing with cement slurry at the entry to the stope

Spray mixing at the tipping point into trucks.

Batch mixing at the entry for trucks

Batch mixing by loader in stockpile bays (reasonably common in small mines in Australia)

A cement percentage of 5 to 6% is normally required for the scale of the mining methods proposed

at Dikulushi, but where there is a risk of poor mixing and excess water content this should increase

to 10%. A 10% composition is typically used in small-scale filling where the fill is mixed by loaders in

stockpiles bays, calculated by adding the required number of cement bags to each loader bucket of

waste placed in the stockpile bay.

16.9. GROUND SUPPORT STANDARDS

Ground support standards have been developed for the Dikulushi underground mining operation by

Turner mining and geotechnical Pty Ltd. The standards include ground support for the declines,

access drives, ore drives and intersections and are shown in sections to 16.9.1 to 16.9.5


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16.9.1. DECLINE SUPPORT STANDARD


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16.9.2. ACCESS SUPPORT STANDARD


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16.9.3. ORE DRIVE SUPPORT STANDARD WITH MESH


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16.9.4. 3 WAY INTERSECTION SUPPORT STANDARD


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16.9.5. 4-WAY INTERSECTION SUPPORT STANDARD


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16.10. WASTE DUMP DESIGN

Waste material from the underground mining activities is expected to be minimal due to the use of

waste material as a major component of the backfill required for the ore mining activities. In

addition to the underground waste produced it is expected that suitable waste rock material will

need to be sourced from the base of the open pit and surrounding waste dumps to make up any

shortfall in waste material required for the production of backfill.

Should the underground mining activities change to mining methods which require significantly less

backfill material, it is expected that the existing waste dump surrounding the Dikulushi open pit can

be modified to accept additional waste material.

16.11. SURFACE WATER MANAGEMENT

The existing surface water management arrangements for the Dikulushi open pit mining operations

are in conjunction with the underground dewatering system and it is expected to be capable of

dealing with any surface water inflow into the underground mining operations.


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17. RECOVERY METHODS

17.1. PLANT FLOWSHEET

The processing plant and associated infrastructure was refurbished prior to start up in June, 2010,

and has been in continuous operation since.

The crushing plant consists of 3 stages; primary jaw crushing, followed by 2 stages of cone crushing

in a closed circuit with a double deck vibrating screen, producing a minus 20 mm product for the

grinding circuit feed. The grinding circuit consists of two overflow ball mills in parallel configuration

in closed circuit with a 250 mm hydrocyclone. Each ball mill is powered by a 750 kW motor. The

grind sizing parameter is 70% passing 106 microns. The mill is capable of treating in excess of

520,000 tonnes of ore per annum.

Both ball mills discharge to a common sump, and the slurry is pumped to a single 250 mm diameter

cyclone. The cyclone underflow gravitates to an Outokumpu SK240 Unit Flotation Cell to recover

coarse liberated copper sulphides, which report directly to the final concentrate. The cyclone

overflow reports to conditioning and conventional flotation at 35% solids.

A relatively simple flotation circuit is in place; the circuit consists of two sections, a primary sulphide

flotation and a secondary sulphide/oxide flotation. See figure 17.1.

Collector and frother addition is conventional when processing low grade ore. The splitting of the

circuit is due to the presence of oxide minerals in some of the ore blends which require activation

using sodium hydrosulphide (Na2S) to enable them to be recovered. As sodium hydrosulphide can

depress some sulphide minerals, the majority of the sulphide minerals are recovered in the primary

sulphide flotation circuit.

The tailings from the primary sulphide flotation circuit are sulphidised and the liberated oxides and

additional sulphides are recovered. In the event of the ore blend containing little or no oxides and

thus not requiring sulphidising, the secondary sulphide circuit acts as a sulphide scavenger. The

primary rougher circuit has provision for bypassing initial rougher concentrates directly to final

concentrate. Lower grade rougher concentrates report to the cleaning flotation cells for upgrading.

Final tailings from the secondary rougher circuit are pumped to the tailings storage facility.

Supernatant water is recovered from the tailings dam and recycled to the processing plant. The

circuit is based on a nominal flotation time of 20 minutes in each of the rougher flotation stages and

a minimum 15 minutes in each of the cleaner stages.

Final concentrate is pumped to a thickener and the underflow is pumped to a concentrate storage

tank. The storage tank has sufficient capacity for 8 hours of concentrate production. A filter press

with a capacity of 194t per day is operated in batch mode. Filter cake discharges directly onto a

concrete floor below the filter where it is recovered and transported to a simple hopper/bagging

arrangement with a skid steer loader. Concentrate is loaded into two tonne capacity bulk bags.

Moisture content is near 10%.


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Each bag is weighed ready for despatch by truck to the Kilwa port as outlined in (Figure 17.1)

Figure 17.1 Dikulushi Plant flow diagram

17.2. TAILINGS STORAGE FACILITIES (TSF)

The first TSF for HMS tailings, designed by Knight Piésold Consulting, South Africa, covers 1.8

hectares and has been dormant since September, 2004. A particularly coarse portion of the HMS

tailings was recovered and processed through the flotation plant by Anvil. MWL has recovered and

processed approximately 15,000t of coarse sand and slime material from this dump

A second TSF (TD2), designed by D.E. Cooper and Associates, Australia, was built during 3Q, 2004 to

receive flotation tailings. This facility is located ~100m North of the HMS TSF, covers about 12

hectares and is 12 m high on the eastern embankment. This facility has also reached capacity.

A third TSF (TD3), designed by Knight Piésold, is located adjacent to and north of the second TSF and

covers a 21 hectare area. This dam is a typical hillside impoundment and provides the needed area

to limit the rise rate of tailings at acceptable norms.

Supernatant tailings water is reclaimed via penstock arrangements for use in the processing plant.

The third TSF (TD3) was utilized until December, 2008 and lay dormant until it was recommissioned

in July 2010. At this juncture Knight Piésold was employed to carry out a volumetric assessment

study to determine the storage capacity of the dam to accommodate 840,000t of tailings resulting


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from the processing of the low grade ore stockpile. The study concluded that a 2 m embankment

raise would be required during 2011.

Deposition continued until October 2010 when Knight Piésold was commissioned to further assess

TD3 expansion capabilities to make provision for an additional 1,500,000t of deposition. The study

concluded that the walls would require to be raised by 6m to accommodate this quantity of tailings.

The raise will be carried out in 2 stages of 3m each, with the first raise completed in mid-2012.

Sub aerial tailings deposition to TD 3 has continued since completion of this first stage 3m raise and

as of June 2013, the TD 3 survey pick up indicated a residual storage volume capacity of 425,000 m3.

Based on operational beach densities and freeboard management allowance, the residual storage

capacity allowance is ensured until the TD 3 stage 2 raise, which is planned for during the dry season

in 2014.

In preparation for uninterrupted TD 3 works during this construction period, an interim storage

capacity lift of 2m is proposed in TD 2 before TD 3 works commence. This will also allow for ongoing

back-up operational capacity.

This will provide a tailings storage facility capable of supporting the underground mining operation.

The information on the upgrading of TD 3 has been supplied by MWL and Sedgman has not reviewed

this data.

17.3. PROCESSING STATISTICS

ANVIL PROCESSING

Anvil processed 137,256 tonnes of low grade between May 2008 and December 2008 when the

open cut run of mine ore ran out prior to full production from underground. Some production

results from the February 2007 to April 2008 can be seen in Table 17.1.

MWL PROCESSING

MWL blended material from surface stockpiles and the HMS Tails through the plant to maximise

copper output between June 2010 and February 2012. This was followed by the treatment of ore

from satellite orebodies such as Boom Gate. The recoveries from this activity were much lower than

from the fresh ore material from either the open pit or underground. It is reasonable to say that

process recoveries and values are associated more with those from the previous open pit and

underground mining operations carried out by Anvil. Processing statistics for the LG material

completed by MWL are shown in Table 17.2.

Commercial production from mining the Dikulushi open pit cutback commenced in November 2012

and was completed in July 2013. Processing of the cut back material is incomplete and open pit cut

back ROM feed material will be processed till the end of 2013. Underground ore processing of ore

will commence for approximately 18 months to Mineral Reserve Completion.


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Table 17.1 Dikulushi processing summary relevant to ore to be mined in the pit cut back

Month Blend

% ROM

Plant Feed

copper%

Silver g/t

Copper Rec%

Silver Rec %

RL mined (Ore Only)

Grade Grade

Conc copper%

Conc silver g/t

Feb-07* 60 5.93 181 85.7 86.9 860 pit stockpile 56.0 1730

Mar-07 100 8.41 273 91.5 87.5 860 pit stockpile 56.0 1745

Apr-07 100 7.65 233 90.7 92.2 850 pit stockpile 55.0 1696

May-07 100 7.61 231 90.2 90.1 850 pit stockpile 55.0 1670

Jun-07 100 7.74 233 91.0 90.5 870 Dev 55.0 1654

July 07* 91.4 7.28 214 89.5 89.8 stockpile 56.5 1668

Aug-07 100 7.92 245 91.1 89.4 850 Dev 56.0 1695

Sep-07 100 7.98 262 91.3 90.2 850 Dev & 890 Stoping 54.0 1890

Oct-07 100 8.18 272 92.4 92.5 870 Dev & 890 stoping 55.0 1821

Nov-07 100 7.81 250 92.2 91.8 870 Dev & 890 stoping 56.0 1793

Dec-07 100 8.45 266 92.8 92.0 830 Dev & 890 stoping 57.0 1772

Jan-08 100 6.00 187 90.1 89.4 830 Dev & 870 Stoping 55.0 1694

Feb 08* 81.4 5.09 154 87.2 87.7 830 Dev & 890/870

Stoping 56.0 1687

Mar-08 100 5.45 188 88.1 79.2 830 Dev & 870 Stoping 54.0 1668

Apr-08 90 4.76 139 87.0 88.4 830/810 Dev & 870

Stoping 54.0 1601

* Low grade ore blended in with the development or stoping ore.

Table 17.2 Processing statistics for the LG material completed by MWL – June 2010 to May 2011

Jun-

10 Jul-10

Aug-

10

Sep-

10 Oct-10

Nov-

10

Dec-

10

Jan-11 Feb-

11

Mar-

11

Apr-

11

May-

11 YTD

Ore Processed tonnes 5,387 36,157 43,882 40,839 27,450 49,029 41,111 49,650 42,839 46,054 40,855 44,705 467,958

Mill Feed Grade Cu % 1.28 1.45 1.04 1.27 3.78 1.52 1.17 1.33 1.32 1.28 1.40 1.32 1.46

Mill Feed Grade Ag g/t 35.87 40.4 27.63 31.72 77.17 41.2 28.5 32.6 29.2 27.8 34.6 33.31 35.31

Tails Grade Cu Cu % 0.34 0.39 0.35 0.46 1.64 0.63 0.52 0.46 0.43 0.46 0.44 0.52 0.54

Tails Grade Ag Ag g/t 10.1 11.1 10.5 10.9 23.0 13.6 10.70 11.3 7.95 7.85 8.7 9.1 10.95

Conc Tonnes dmt 128 896 719 783 1,380 1,066 684 1001 890 893 906 865 10,211

Conc Grade Cu Cu % 38.7 43.5 42.7 43.0 44.1 41.5 39.74 40.1 41.6 39.35 40.2 41.7 41.66

Conc Grade Ag Ag g/t 1,067 1,138 1,107 1,119 1,139 1,188 1139 1070 1033 941 1092 993 1089

Cu metal in Conc dmt 51.45 389.6 306.9 336.7 608.5 442.7 272 400 366 351 365 361 4,251

Ag metal in Conc oz 4,384 32,778 25,581 28,177 50,534 40,726 25,057 32,737 29,385 27,279 31,904 27,559 356,101

Recovery Cu % 74.62 74.31 67.25 64.92 58.64 59.41 56.54 64.13 66.91 62.66 67.46 61.37 64.05

Recovery Ag % 70.57 69.88 65.62 67.65 74.20 62.68 66.45 62.70 73.09 63.34 71.25 61.3 66.68


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Table 17.3 Processing statistics for the LG material completed by MWL – June 2011 to June 2012

Table 17.4 Processing statistics for the Open pit cut back ROM completed by MWL – Jul 2012 to July 2013

As stated in Section 13.3 the production data for the Period June 2011 to July 2013 was supplied by

MWL and has not been reviewed by Sedgman.

Jul-12 Aug-12 Sep-12 Oct-12 Nov-12 Dec-12 Jan-13 Feb-13 Mar-13 Apr-13 May-13 Jun-13 Jul-13 YTD-13




Tails Grade Cu Cu % 0.65 0.59 0.55 0.45 0.42 0.44 0.40 0.43 0.36 0.44 0.51 0.61 0.46 0.48


Concentrate Tonnes Produced dmt 1,448 1,334 1,476 1,200 1,138 2,711 2,685 2,939 2,936 3,087 2,470 2,798 3,723 29,944


Concentrate Grade Ag Ag g/t 620 677 622 725 1,733 1,462 1,532 1,787 1,620 1,499 1,903 1,952 1,864 1,515

Cu metal in Concentrate dmt 505 487 554 486 696 1,584 1,596 1,804 1,829 1,936 1,552 1,701 2,195 16,925

Ag metal in Concentrate oz 28,876 29,014 29,496 27,950 63,410 127,429 132,235 168,837 152,974 148,824 151,060 175,621 223,058 1,458,783

Recovery Cu % 76.9 75.2 77.4 78.7 92.4 92.6 93.3 93.8 93.4 93.9 95.8 95.4 94.2 91.5

Recovery Ag % 76.9 75.4 76.0 83.2 87.1 90.2 90.5 94.2 90.3 88.7 93.9 94.0 92.1 90.3


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18. PROJECT INFRASTRUCTURE

The Dikulushi operation is an operating mine and the infrastructure remains in place. It has been

used and maintained by MWL since it took over the project site. The infrastructure is considered

adequate for the continuation of the operations with the resumption of underground mining

activities.

18.1. SURFACE FACILITIES

The existing surface facilities (Figure 18.1) remaining from the previous underground operations and

open pit cut back operations will be suitable for use by the underground mining personnel (Figure

18.1 and Figure 18.2).

Figure 18.1 On-site office facilities at Dikulushi

Figure 18.2 On-site Underground change room facilities at Dikulushi


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18.2. POWER

The project is located in a remote area where there is no electrical utility grid. The mine power is

supplied by diesel generators. Power for the Dikulushi operation will be provided by the existing

diesel powered electricity generation installation. This installation has previously supplied power to

the camp and the processing plant. Current production plans will not exceed previous levels and the

installed capacity is expected to be sufficient for future activities. There is sufficient back-up

capacity.

The existing power station at Dikulushi comprises the following generators: 4 x 1.2 MW FG Wilson (

new units and installed during the 3rd quarter 2013 ), 1 x 2.0 MW Caterpillar, 1 x 1.6 MW Caterpillar

and 1 x 0.8 MW Mirrlees, for a total capacity of 9.2 MW. The current power demand for the plant

and infrastructure is in the order of 1.8 MW. The 2.0 MW Caterpillar and 1.6 MW Caterpillar

generators currently require major overhauls, which will be completed during 2014. The 1 x 0.8 MW

Mirrlees will be decommissioned during the 4th quarter of 2013. The new FG Wilson generating sets

were installed to supply power to the operations as well as dewatering of the underground and

normal underground operations.

18.3. PROCESS WATER SUPPLY

Lake Newton on the perennial Dikulushi stream provides storage for dewatering and serves as a

reservoir for the supply of process water.

Meteorological data has been collected between August 2005 and August 2013, with the exception

of 2009 and 2010, where little or no data was collected due to reduced activities on site. The mine

water flow regime has changed over the past 3 years and the current water supply and balance

system is shown in Figure 18.3.

There are several sources of water on site:

The Dikulushi stream, which traverses adjacent to the mine, has two abstraction points. The

first abstraction point feeds water to the Process Plant, with the flow being measured. The

second abstraction point is at Lake Newton which stores the water before routing it to the

Return Water Dam (RWD) as make-up water. The flow between Lake Newton and the RWD

is measured.

The second source of water is from the Stream Borehole which supplies the Process Plant.

The borehole is not currently in operation.

The current main source of water is from the Open Pit which has a single supply pipeline to

the Process Plant which is metered.

Water from Tailings Dam 3 (TD3) is captured at the RWD where it is routed to the Process

Plant, This flow is also metered.

Other flow metered points are the Admin Building and the Power House which are internal plant

meters. The Truck Feed receives water from Lake Newton and is used for dust suppression around

the mine. A recent review and update was carried out on the full water balance during June and July


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which are amongst the driest months in the year and this has been used to update the current

average water balance model, which is presented in Figure 18.3

Figure 18.3 Average water balance

Based on the water balance review the following were noted:

The total inflow onto the Tailings dam under present conditions is 3,607 m3/d of which

1,856 m3/d is returned back to the return water pond for reuse in the plant;

The RWD gets about 2,433 m3/d from Lake Newton via the river and 1,856 m3/d from the

tailings dam, while a total of 1,940 m3/d is sent to the plant for reuse;

The main losses from the Tailings dam are seepage, evaporation and interstitial storage;

The rainfall onto the open pit that is collected in the sumps below is reused in the plant and

only when water cannot be reused in the plant is the water discharged into Lake Newton

after the water is settled;

The extended waste footprint means that there will be runoff from the dump that will need

to be settled in paddocks and evaporated where possible;

Borehole water is currently not being used but will be used as potable water and make-up

water when it is in operation;

Approximately 1,589 m3/d will need to be supplied from Lake Newton or from boreholes to

sustain the mine during the dry months;

During the wet season there will be times where the water will discharge from the RWD into

the perennial stream as the plant will not be able to use all the water in the process.

Process Plant Water SourcesFirst option, must keep the water level down.

Last option.

150m3/Day

1715m3/Day

To the Plant

3387m3/Day KEY

Main saurce for Process and Raw water. A Lake Newton

The river pond feed also the Process only B Return Dam Water

when required. C Boom Gate water

2205m3/Day D River

E U/G Water 1

F U/G Water 2

515m3/Day G Process Water

H Admin/gardens

I Power House

J Gland service/lube cooling water

Flowmeter

32m3/Day 117m3/Day

To admin/garden Pump

1956m3/Day

To Power House 32m3/Day

764m3/Day

Emergency backup for Raw water.(Gland service)

Boom gate is the backup for Raw water and also Process water.

Can feed also the river pond and overflow to Lake Newton.

Returned

Water Pond

Tailing

Dam

Lake

Newton

Process Water

U/G

B

A

C

D

E

DIKULUSHI

RIVER

Raw water

BG pit

Camp

BoreHole

F

G

H

I

J


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19. MARKET STUDIES AND CONTRACTS

19.1. MARKETS

The Dikulushi plant is currently configured to produce a copper/silver concentrate which contains

approximately 50% copper.

MWL has not yet committed any of the concentrate that will be produced from the underground.

This will be done via a tender process.

In July 2013, the Company entered into a copper hedging transaction, forward selling 3,500 tonnes

of copper at $6,875 per tonne for financial settlement at intervals commencing in November 2013

and ending in March 2014.

19.2. CONTRACTS

MWL currently has a contract to sell the copper concentrate produced from the open pit project to

Transamine Trading. This contract was awarded after a tender process.

There are various contracts either already in place or required to be entered into for the following

major areas:

Explosives

Diesel supply

Transport

Reagents

Spares

MWL recognises that a consistent reliable fuel supply is crucial to the success of the Dikulushi

operation. The operation currently uses approximately 500,000 litres of diesel per month and will

need additional fuel for the underground operations. This fuel is supplied by four DRC based

companies; two receive supplies from the port of Beira in Mozambique and the other two receive

supplies from the port of Dar Es Salaam in Tanzania. MWL had no interruptions during the open pit

project when it was receiving up to 1,200,000 litres per month. Thus MWL believes that it has

mitigated the risk of fuel supply by having a number of suppliers whom source fuel from different

ports.

MWL has a number of supply contracts for various inputs required for operations. MWL also has

contracts to transport of concentrate from Dikulushi to Kilwa.

The current revenue estimates include the concentrate being sold to the Concentrate trader who

will on sell the concentrate to smelters, where it will be converted to metal and sold to the market.

MWL receives 90% provisional payment for material delivered to Nchelenge in Zambia. A further

10% is received after finalisation of QP pricing and assay exchange.


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Costs for the transport, as well as the treatment and refining charges for the copper and silver

concentrates, along with the final net smelter returns have been sighted and used in calculations for

the Mineral Reserves, but due to the commercial in confidence agreements with the smelters are

not shown here.

The study has used a copper price of $6,800/tonne copper ($3.08/lb. copper) and a silver price of

$20 /oz silver.

No formal off-take agreements have been confirmed to support these assumptions, but the

expected revenue parameters are based on assessments completed by Mawson West of likely

conditions and forward price curves

The average cost per tonne of copper product for transport, treatment, refining and clearing is

estimated to be $1,153 per tonne of copper metal sold.

Long term commodity price projections have not been evaluated due to the short life of the

underground and processing operations being less than 24 months, at this stage.


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20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL

OR COMMUNITY IMPACT

An Environmental Impact Assessment (EIA) for the Dikulushi project was lodged in 2003. In 2009,

an EIA for the underground Project was submitted to the DRC Government. Both of these reports

were compiled by African Mining Consultants of Kitwe, Zambia, an environmental company that was

licensed to work and report in the DRC. In 2011 an EIA for the cutback project was prepared by EMIS

sprl, a DRC environmental company licences to work and report in the DRC. All three environmental

reports received DRC Government approval. A revised EIA, extending underground mining beyond

2013, has been submitted to the Government for approval.

MWL is required to provide annual environmental reports and demonstrate that it is in compliance

with the EIA. Mine remediation is one of the compliance items in the EMP.

MWL has lodged an environmental bond of $1.19M. The financial guarantee is a contribution

towards an estimate of the total costs of closure, rehabilitation and re-vegetation of the Dikulushi

mine. The development of the financial guarantee is conducted in compliance with:

Articles 410 of the Mining Regulations

Articles 124 and 125 of Appendix XI of the DRC Mining Regulations 2003; and

Appendix II of the Mining Regulations 2003.

The company recently had completed an annual review of the EIA which has been lodged. This

review did not find any non-compliant items, or any breaches of the permitted conditions and

requirements.

MWL has number of corporate social responsibility programs that are run on the Dikulushi Property.

The key programs are –

A) The Dikulushi-Kapulo foundation – a community foundation to initiate, develop and support

development projects for the benefit of local communities in health, education,

infrastructure and reinforcement capabilities. The foundation acts as a catalyst to support

community initiatives and development projects.

B) Employment and training – MWL employs approximately 900 local employees. MWL has

introduced various training programs that are also available to the local community. A

teacher has been employed to assist women with language education in French, English and

Swahili.

C) Community Health – MWL has joined with the Australian Government’s DAP program and

the program has contributed funds to improve the available health care facilities present in

the Dikulushi community clinic.

D) Education – MWL has contributed funds towards the upgrade of the Dikulushi School, to

enclose classrooms and provide classroom equipment and resources. This project benefits

1,100 local school students.

E) Kilwa Electrification – this project is working with the Kilwa community to provide power to

the hospital and surrounding buildings in the village.


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F) Kipeto Community garden project – This project is to assist the community to establish a

vegetable nursery in the community.

G) MWL is committed to supporting local business by sourcing certain supplies from local

villages surrounding Dikulushi and the Kapulo projects. Figure 20.1 is an example of this

commitment.

Figure 20.1 Community Business making work clothes for the mine.


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21. CAPITAL AND OPERATING COSTS

21.1. CAPITAL COST ESTIMATE

Capital expenditure for the underground mining operations is estimated at around US$9M. The bulk

of the Capital expenditure for the project is focussed on the mining equipment related to the

underground operations. All other capital for the processing plant, infrastructure and administration

is already in place and ongoing from the Open pit cut back operations.

The list of items included in capital expenditure cost is shown in Table 21.1. Due to the short time

period that the Mineral Reserves will be mined over, it has been planned that any additional major

mining equipment required will be obtained on a hire arrangement, with costs covered as part of the

annual operating cost. This has provided a significant saving in the capital spend for the re-

establishment of the underground operations. Additional capital savings have been achieved

through the refurbishment of the previous underground mining fleet and this is to provide the bulk

of the mining fleet required to mine the underground Mineral Reserves.

Table 21.1 Dikulushi underground capital expenditure cost estimate.

Item Expenditure US$

000’s

Refurbishment of mobile equipment 1,109

2 x UG Loader & 2 x Truck (new) UG dewatering system & bores Vehicles & ancillary units

3,362 2,570 560

Ventilation fans 400

UG compressors 150

Safety equipment and systems 590

Ancillary support equipment 865

Total 9,606

The total composition of the mining fleet is shown in Table 21.2. The majority of this fleet is already

located on-site and was used as part of the previous underground mining activities conducted by

Anvil.

It is expected that parts of the fleet will have low utilisations due to the small tonnages being

produced from the underground mining operations, and this will provide ample coverage for

breakdowns. The additional pieces of equipment required will provide the operation with the

flexibility to increase mining rates as required.


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Table 21.2 Major mining fleet and equipment required for the extraction of the Dikulushi underground Mineral Reserves

Mining fleet Number

LHD 2

Truck 3

Jumbo twin 2

Jumbo single 1

IT 1

Charge up vehicle 1

PC 1

Ute 2

Air leg drills 6

Wire Line Scraper 1

The majority of the above capital is spent over the first 6 months of the project.

21.2. OPERATING COST ESTIMATE

The underground mining operations at the Dikulushi project will be run under an owner operator

model. MWL will provide all the equipment and personnel to complete the mining operations. For

the purpose of estimating the operating costs, the costs have been broken up into mining variable

costs and overhead/fixed costs. This approach has been chosen so that the costs that are directly

involved with development and ore production from the underground, will vary with the advance of

the physical mining operation and are separated from other costs such as labour, management and

supervision that are generally fixed operating costs on a month by month basis.

21.2.1. MINING OPERATING COST

Table 21.3 shows the fixed mining and processing costs associated with the Dikulushi underground

operation. The costs included in this table were provided by MWL and are based upon cost

estimates in financial modelling already completed to evaluate underground mining at the site. G&A

costs are included in the Site Costs item, along with camp costs, OH&S, site wide power and water

supply for the camp.

Variable costs are listed in Table 21.4, including development, haulage backfill and rehabilitation.

Capital development, vertical development and production development costs are all inclusive of

jumbo drilling, blasting, loading, ground support and services, including all the equipment operating

costs involved. Costs have been calculated on a lineal metre basis.

Decline costs have been calculated assuming drive dimensions of 5.5 mH by 5.5 mW.

Level access and return airway drives have been calculated using dimensions of 5.0 mH by

5.0 mW.

Production access and ore drives assume dimensions of 5.0 mH by 4.5 mW.

Vertical development rises assume dimensions of 1.5 mH by 1.5 mW for escape ways and

5m diameter for vent rises.


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Haulage costs are inclusive of loading, hauling to the waste dump or mill, and equipment

operating costs.

Table 21.3 Mining overhead and fixed costs

Mining overhead and

fixed costs Units

Rate

(million) Comments

Mining Personnel $/yr $0.75

Workshop Personnel $/yr $0.33

Technical Support Roles $/yr $2.60 Engineering, Geology, survey etc.

Fixed Plant Costs $/yr $0.66

Power Consumption $/yr $1.76

Table 21.4 Mining variable costs

Mining variable costs Units Rate Comments

Capital Development

Decline $/m 1,739 5.5mH by 5.5mW

Level access / Stock piles $/m 1,677 5.0mH by 5.0mW

Return Air Way Drives $/m 1,739 5.0mH by 5.5mW

Vertical Development

Air leg Rises Escape way $/m 83 1.5mH by 1.5mW - including equipping

Vent Rise $/m 568 5.0m diameter, air leg pilot + airleg strip

Production Development

Cut and Fill Ramp Access $/m 1,330 5.0mH by 4.5mW

Ore Drives / Cut and fill $/m 1,330 5.0mH by 4.5mW

Haulage

Ore $/tkm 2.21 Loading and haulage to waste dump or mill –

incl. equipment operating costs Waste $/tkm 3.19

Backfill

CRF $/m3 44.57 -

CAF $/m3 55.70 -

RF $/m3 3.23

Rehabilitation

Drive rehab $/m 814 Assumes all ground support is replaced

21.2.2. PROCESSING OPERATING COSTS

A $55.00 per ore tonne processing cost has been applied to all tonnes processed from the

underground mining operations. This was provided by Mawson West and represents the overhead

and variable processing costs associated with the operation of the processing plant for the

production of the copper concentrate.

21.2.3. MANAGEMENT AND ADMINISTRATION COSTS

The management and administration costs are based on the existing cost structure of the current

operations and includes management and administration personnel, OH&S costs, logistics costs,

camp costs and sustaining capital required for non-production related activities. The total annual

cost is estimated at $10.6M per year.


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21.2.4. TRANSPORT AND SMELTING COSTS

Costs for the transport, as well as the treatment and refining charges for the copper and silver

concentrates, along with the final net smelter returns have been sighted and used in calculations for

the Mineral Reserves, but due to the commercial in confidence agreements with the smelters are

not shown here.

21.3. METAL PRICES

Financial modelling has used a copper price of US$6,800/t and a silver price of US$20/oz. Details are

provided in Table 21.5.

Table 21.5 Metal prices used in modelling

Product Units Rate

Copper $/t 6,800

Sensitivity Low $/t 6,120

Sensitivity High $/t 7,480

Silver $/oz 20

Sensitivity Low $/oz 18

Sensitivity High $/oz 22

22. ECONOMIC ANALYSIS

22.1. OPERATIONS SUMMARY

Table 22.1 below provides a summary of the mining and cashflow performance of the Project mine

including the extraction of the Crown pillar ore. The total material mined from the underground

operations is 221 kt of material of which 173 kt is ore at a diluted and recovered grade of 5.15%

copper.

The average mining cost $83 per total tonne and $106 per ore tonne. The processing cost is $55 per

ore of tonne processed. Management and administration costs are $97 per tonne of ore processed.

Due to the long history of mining operations at the Dikulushi project, capital expenditure is minimal

as the site already has the processing and major infrastructure in place from previous open pit and

underground mining activities. In addition to the facilities the majority of the mining equipment

required for the underground operations exists on site. The total capital costs allowed for as part of

the underground mining project is $9M. Sustaining capital for the operations has been included in

the above capital cost.

The mine life is 19 months with the majority of the ore mined in the first 12 months of the mining

operation, and ore production reduces towards the end of the mine life, as the number of

production ore headings available reduces. Processing, copper production and sales have been

scheduled in the month of extraction for the purpose of this economic evaluation.

The processing recovery for copper used for this estimate was 94%, with silver recovery at 90% to

produce a copper concentrate grading approximately 60% copper.


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The total cash cost of the operation is $44.6M. A total of 8.1 Kt of copper is sold along with 573 Koz

of silver to produce net revenue of $57.7M, and net cashflow of $3.0M. Due to the short life of the

existing underground workings no discount rate has been applied to the revenues in Table 22.1


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Table 22.1 Dikulushi mining and financial summary

Pre-mining Qtr 1 Qtr 2 Qtr 3 Qtr 4 Qtr 5 Qtr 6 Total

Physical Schedule

Total material mined tonnes 3,564 50,672 30,517 25,044 38,069 38,713 34,548 221,128

Waste Mined tonnes 3,564 17,346 4,271 6,084 5,576 11,044 0 47,885

Ore Mined tonnes 0 33,327 26,246 18,961 32,493 27,669 34,548 173,243

Copper mined grade copper% 0.00 5.08 4.97 6.16 5.78 4.38 4.83 5.15

Silver mined grade silver g/t 0.0 142.7 102.9 126.9 151.1 94.3 133.7 127.0

Copper mined t 0 1,693 1,304 1,168 1,877 1,212 1,667 8,921

Silver mined oz 0 152,913 86,871 77,338 157,804 83,929 148,518 707,373

Costs

Mining

Decline Rehabilitation $ 134,790 0 0 0 0 0 0 134,790

Jumbo development $ 42,538 1,186,964 829,203 349,937 702,380 923,050 324,052 4,358,124

Airleg Development $ 3,811 28,678 0 0 0 0 0 32,488

Longhole Stoping $ 0 0 0 0 0 0 427,327 427,327

Ore Haulage $ 0 140,714 114,775 40,237 94,240 115,531 59,772 565,269

Waste Haulage $ 24,091 119,642 31,200 37,816 37,977 68,728 0 319,454

Backfill $ 0 307,128 994,495 306,330 586,851 472,853 172,134 2,839,791

Mining & Workshop Labour $ 88,110 264,331 264,331 264,331 264,331 264,331 264,331 1,674,093

Fixed Plant $ 55,414 166,243 166,243 166,243 166,243 166,243 166,243 1,052,874

Power $ 146,792 440,376 440,376 440,376 440,376 440,376 440,376 2,789,046

Technical Support $ 216,313 648,938 648,938 648,938 648,938 648,938 648,938 4,109,938

Total Mining Opex $ 711,859 3,303,014 3,489,560 2,254,206 2,941,334 3,100,050 2,503,172 18,303,195


P a g e | 164


Total Mining Opex $/ore tonne mined 0.00 99.11 132.95 118.89 90.52 112.04 72.45 105.65

$/total tonne mined

199.72 65.18 114.35 90.01 77.26 80.08 72.45 82.77

Processing

Processing $ 0 1,832,958 1,443,546 1,042,835 1,787,108 1,521,768 1,900,139 9,528,355

$/ore tonne milled 0.00 55.00 55.00 55.00 55.00 55.00 55.00 55.0

Management & Admin

Administration $ 883,333 2,650,000 2,650,000 2,650,000 2,650,000 2,650,000 2,650,000 16,783,333

$/ore tonne milled 0.00 79.52 100.97 139.76 81.56 95.78 76.70 96.88

Total Operating Costs $ 1,595,192 7,785,972 7,583,106 5,947,041 7,378,442 7,271,818 7,053,311 44,614,883

Total Capital Costs $ 4,990,000 3,921,000 495,000 200,000 9,597,000

Sustaining Capital $ 26,316 78,947 78,947 78,947 78,947 78,947 78,947 500,000

Revenue

Metal in concentrate copper t 0 1,592 1,225 1,097 1,765 1,139 1,567 8,385

silver oz 0 140,680 79,922 71,151 145,180 77,215 136,636 650,783

Metal sold copper t 0 1,537 1,184 1,060 1,705 1,101 1,514 8,100

silver oz 0 124,309 70,157 62,448 128,109 67,845 120,705 573,573

Sales & Transport Costs $ 0 1,394,211 1,059,225 948,462 1,540,389 986,344 1,371,896 7,300,528

Duties and Taxes $ 0 297,711 223,249 199,859 327,811 208,192 292,751 1,549,574


P a g e | 165


Copper NSR $ 0 10,454,484 8,049,030 7,208,970 11,591,067 7,484,181 10,294,229 55,081,959

$/t mined 0 206.31 263.75 287.85 304.48 193.33 297.97

Silver NSR $ 0 2,486,180 1,403,131 1,248,965 2,562,176 1,356,902 2,414,105 11,471,459

$/oz mined 0 49.06 45.98 49.87 67.30 35.05 69.88

Total Revenue

Revenue from Sales $ 0 11,546,452 8,392,935 7,509,473 12,612,854 7,854,738 11,336,438 59,252,891

Royalties/taxes $ 0 297,711 223,249 199,859 327,811 208,192 292,751 1,549,574

Net Revenue $ 0 11,248,741 8,169,686 7,309,614 12,285,043 7,646,546 11,043,687 57,703,317

Cashflow from Operations $ -6,611,508 -537,178 12,632 1,083,626 4,827,653 295,781 3,911,429 2,982,434

NPV NPV $2,982,434 8% IRR


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22.1.1. SENSITIVITY ANALYSIS

MWL has carried out a sensitivity analysis on the cash flow forecasts, and this is provided in Table

22.2.

Table 22.2 Sensitivity analysis on the cash flow forecast for underground mining and treatment at Dikulushi

Table

Dikulushi Copper Project

Project Sensitivity to a Change in copper Price

NPV (US$ million)

Change in copper Price

-10% -5% 0% 5% 10%

-9.8 0.7 3.0 5.7 8.4

Table


Project Sensitivity to a Change in Silver Price

NPV (US$ million)

Change in Silver Price

-10% -5% 0% 5% 10%

1.8 2.4 3.0 3.6 4.1

Table


Project Sensitivity to a Change in Operating Costs

NPV (US$ million)

Change in Operating Costs

-10% -5% 0% 5% 10%

7.4 5.2 3.0 0.8 -1.5

22.2. PAYBACK

As discussed the refurbishment cost of the mill has already been covered by the revenues from the

LG stockpile treatment and the previous open pit mining activities, thus there is no formal capital

payback period. The development of the underground mining is to be fully funded out of MWL’s

current existing cash reserves. The maximum negative cashflow (including capital costs) is -$1.7M

(end of the pre-mining period) and cash flow moves back into positive territory during the third

quarter of operation.


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22.3. MINE LIFE

Mine life is based on the Probable mining Mineral Reserve schedule and is approximately 19 months.

This allows a one month lead for rehabilitation of the existing underground workings, and the

development of ventilation rises to establish primary ventilation along with one extra month at the

end of the schedule, to complete ore backfilling operations.

22.4. TAXATION

The Dikulushi mine operates under the Dikulushi Mining Convention, which provides for

concessionary rates of taxation for each new mine. The first five years of production were tax free,

the effective tax rate from the sixth through tenth years of production is 16% and for the eleventh

through fifteenth years of production 18%, thereafter 40%. Dikulushi has been producing for

approximately eleven years.

In addition to the usual deductions of expenses and accruals, the Dikulushi Mining Convention

provides that taxable income is adjusted by allowances for:

depreciation of moveable and immoveable fixed assets,

a “depletion allowance” equal to 15% of gross sales up to 50% of net profit, and

all exploration and evaluation expenses.

AMC also receives the benefit of concessionary import duty rates. During the construction phase,

2% import duties are applied and then during production import duties are applied at the rate of 3%

for fuel, lubricants and mining consumables and 5% of all other supplies.


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23. ADJACENT PROPERTIES

There are no significant mining properties adjacent to the Dikulushi Property.

MWL’s Kapulo copper exploration and development project is also part of the Dikulushi Mining

Convention that includes the Dikulushi Property. However, MWL considers the projects to be

separate and non-contiguous. The Dikulushi Mining Convention applies to an area of approximately

7,300 km2 and the two projects are located 124 kilometres apart, are on distinctly separate leases,

and are separated across this distance by Lake Mweru and the Luapula River. Road access to the

Dikulushi Mine is from the DRC side of Lake Mweru, while road access to the Kapulo Project is from

the Zambian side of Lake Mweru. There is only rudimentary road access between the two projects.

Development of the Kapulo Project is not dependent on, and will not share infrastructure with, the

Dikulushi Mine. Each of the projects will have their own separate mills, facilities, equipment and

administration, and will conduct independent processing operations. A definitive feasibility study on

the development of the Kapulo Project is the subject of a NI 43-101 technical report dated June 30,

2011 entitled “Kapulo Copper Project, DRC, National Instrument 43-101 Technical Report”.


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24. OTHER RELEVANT DATA AND INFORMATION

Historically, Dikulushi was a producing open pit operation from 2002 until 2006. It continued for a

period of time supplying ore from underground operations until closure in November 2008.

The Dikulushi mine was acquired from Anvil by Mawson West Limited in April 2010 and work started

immediately on refurbishment of the plant, which was completed in June 2010. Since June 2010,

MWL has produced copper-silver concentrate from a feed of blended HMS tails and reclamation of

the LG stockpile, as well as the successful mining and processing of the Open pit cut back.

The next stage of the operations is to now re-establish the underground workings and re-commence

underground production as outlined in this report. It is also MWL’s intention to continue

exploration drilling of the Dikulushi orebody from accessible locations within the underground.


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25. INTERPRETATION AND CONCLUSIONS

The Dikulushi Property is a producing and developing property. Current processing of the Dikulushi

LG stockpile reserves and open pit cut back ore has provided MWL with a robust cash flow, and

production results demonstrate reliable grades of remaining ore stocks when compared with

Mineral Reserve estimates.

The Dikulushi deposit has a history of exploration and successful mining. Data quality across the

unmined volume of the deposit is of good quality and has representative sample values for reliable

Mineral Resource estimates. Mineral Resource classification supports both Proven and Probable

Reserve categories within the underground Project. The pre-feasibility study and resulting Mineral

Reserves from the underground further extends MWL’s production life from the Dikulushi Project.

MWL has an opportunity, with the re-establishment of the underground, to actively pursue

exploration drilling from selected locations in the underground to be able to upgrade Inferred

Mineral Resource material and extend the total Mineral Resource. MWL intends to continue

processing the open pit cut back ROM material during the re-establishment and build up phase to

production from the underground.

MWL’s strategy is to continue to develop satellite deposits around Dikulushi, such as Kazumbula, in

addition to extending the remaining Dikulushi Mineral Resources located below the underground.

The recent exploration drilling at Kazumbula has provided geological and sample information to

support a robust Mineral Resource estimate. Upon completion of the mine design, scheduling and

financial analysis, the Kazumbula deposit is most likely to be of reasonable size and grade to be able

to contribute feed to the Dikulushi plant. Additional satellite deposits within 50 km of Dikulushi are

currently being drilled by MWL.


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26. RECOMMENDATIONS

It is recommended that MWL continues with the planned underground Project. The key aspects for

the success of the underground Project are the establishment and maintenance of sound

underground mining practices, including a key focus on the drilling and blasting operations and the

post blasting ground support regimes. The backfilling operations and the selected fill support

method for the ground conditions, will require on-going attention and review, as well as close

attention to the fill specifications as recommended. Ongoing test and study work is recommended

on the investigation of using the tails dam material as underground paste fill. The initial pre-

feasibility tests showed positive results.

With the recommencement of undergrounds operations, it is recommended that exploration drilling

be continued from selected underground positions to test the orebody at depth and to assist with

the re-classification of the remaining Mineral Resource up to the Indicated Category and thus

Probable Reserves.

Ongoing annual reviews are required for the environmental approvals and permitting and a key

component is to ensure the integrity of the tailings dam continues to be maintained.

The development of additional targets within the 50 km radius of Dikulushi has good synergies with

the overall MWL strategy.


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27. REFERENCES

DevMin Pty Ltd (Feb 2004): Anvil Mining Ltd “Dikulushi Copper-Silver Deposit, NI34-101 Technical

Report. February 16, 2004.

Franey, N., Hillbeck, M. and Fahey, G. (2006): Technical Report, Dikulushi Copper – Silver Deposit.

February 21, 2006

JORC (2004): Australasian Code for Reporting of Mineral Resources and Ore Reserves, Effective

December 2004. Prepared by the Joint Ore Reserves Committee of The Australasian Institute of

Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia (JORC).

National Instrument 43-101, Standards of Disclosure for Mineral Projects, Supplement to the OSC

Bulletin, April 8, 2011

Form 43-101F1 Technical Report, Supplement to the OSC Bulletin, April 8, 2011

Munro, K.D. & Associates (1998): Dikulushi Copper-Silver Project. Geological Review and Mineral

Resource Estimate for Dikulushi Copper-Silver Project.

Lemmon, T., Boutwood, A., Turner, B., (2003) The Dikulushi copper-silver deposit, Katanga, DRC. In,

Proterozoic Sediment-hosted base metal deposits of Western Gondwana, ed., J. Cailteux, Abstract of

the IGCP 450 conference and field workshop, July 14-24. Lubumbashi, DRC.

Dewaele, S., Muches, P., Heijlen, W., Lemmon, T., Boutwood, A., (in press), Reconstruction of the

hydrothermal history of the CU-Ag vein-type mineralisation of Dikulushi, Kundelunga foreland,

Katanga, DRC.

Fahey,G.,Franey,N., Anvil Mining Limited Dikulushi Copper-Silver Mine Katanga Region Democratic

Republic of Congo technical Report (NI43-101), December 22nd, 2006

Mawson West Ltd Pre-Feasibility study, July 2011

Independent Metallurgical Laboratories (IML): Metallurgical Ore Characterisation of Dikulushi

Copper Ores for Anvil Mining NL, August 2003

Independent Metallurgical Laboratories (IML): Confirmatory Metallurgical Testwork on ROM

Dikulushi Copper Ore for Anvil Mining NL, June 2004

Metallurgical Design and Management Pty Ltd; Dikulushi Copper Silver Project, Stage 2 Flotation

Project Interim Metallurgical Rreport, July 11, 2003

F Chikosha, Dikulushi Copper Mine Tailings Disposal Facility TD3 Expansion Study, June 2011

A J Strauss, Dikulushi Copper Mine Tailings TD3 Volumetric Assessment, July 2010

M.Turner, Indpendent geotechnical consultant: Dikulushi north wall cable bolts 270711, July 2011

M.Turner, Indpendent geotechnical consultant: MHTurner Project stability 260711, July 2011.

SRK Consulting: Project No: 436159 Water Balance for Dikulushi Mine – 2011 Update


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Mawson West, Dikulushi Underground Mining Study, June 2013

Peter Wade, Capital Mine Consulting; Dikulushi Mining Operating Cost Review, June 2013

M.Turner, Dikulushi Underground Re-opening Geotechnical Study Report No. 0713, August 2013.

J.Keogh, Dikulushi Underground Underhanded Cut & Fill Cemented Rockfill Preliminary Design

Technical Note 13032, August 2013


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28. CERTIFICATES

OPTIRO PTY LTD

CERTIFICATE OF QUALIFIED PERSON – ANDREW LAW

As the lead author and a Qualified Person of the report entitled “Technical Report on the Dikulushi

Underground Project, Democratic Republic of Congo” (the Study) dated 12 December 2013, on the

Underground Project of Mawson West Limited, I hereby state:

1. My name is Andrew Law and I am a full time employee of the firm Optiro Pty Ltd of Level 4,

50 Colin Street, West Perth, WA, 6005, Australia.

2. I am a practising Mining Engineer and a Fellow of the AusIMM (107318), also a Fellow of the

Institute of Quarrying Australia (991004), and a Member of the Australian Institute of Company

Directors (0044149).

3. I am a graduate of the Witwatersrand Technikon, Johannesburg, South Africa, with a HND

Metalliferous Mining, in 1982.

4. I have practiced my profession continuously since 1983.

5. I am an “independent” and “qualified person” as the terms are defined in National Instrument

43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have performed consulting services and reviewed files and data associated with the Dikulushi

Project from August 2011 to the present.

7. I visited the Dikulushi Project property and the underground as far as the 850Mrl (est water

level) in February 2012. I have performed consulting services and reviewed files and data

associated with Dikulushi between August 2011 and the present time

8. Based on the information provided by Mawson West Ltd and reviewed by myself, I contributed

to Sections 1,4,5,6,15, 16, 19, 20, 21, 22, 24, 25, and 26.

9. As of December 12, 2013, the effective date of the Study, to the best of my knowledge,

information and belief, the Study contains all scientific and technical information that is required

to be disclosed to make the Study not misleading.

10. I have read the National Instrument and Form 43-101F1 (the “Form”) and the Study has been

prepared in compliance with the Instrument and the Form.

11. I do not have nor do I expect to receive a direct or indirect interest in the Dikulushi property of

Mawson West Ltd, and I do not beneficially own, directly or indirectly, any securities of Mawson

West Ltd or any associate or affiliate of such company.

Dated at Perth, Western Australia, on the 20 December 2013.

Andrew Law

Director - Mining (Optiro Pty Ltd)

FAusIMM


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OPTIRO PTY LTD

CERTIFICATE OF QUALIFIED PERSON – IAN GLACKEN.

As one of the authors of the report entitled “Technical Report on the Dikulushi Underground Project,

Democratic Republic of Congo” (the Study) dated 12 December 2013, on the Underground Project of

Mawson West Limited, I hereby state:

1. My name is Ian Glacken and I am a full-time employee of the firm Optiro Pty Ltd of Level 4,

50 Colin Street, West Perth, WA, 6005, Australia.

2. I am a practising geologist and a Fellow of the AusIMM (107194) and a Chartered Professional

Geologist. I am also a Member of the Institution of Metals Mining and Materials (IMMM, 46394)

and a Chartered Engineer of this Institution.

3. I am a graduate of Durham University in the United Kingdom with a BSc (Hons) in Geology in

1979, the Royal School of Mines in the United Kingdom with MSc in Mineral Exploration in 1981

and Stanford University in the USA with an MSc in Geostatistics in 1996.

4. I have practiced my profession continuously since 1981.

5. I am an “Independent” and “qualified person” as the terms are defined in National Instrument


6. I have not visited the Dikulushi Project property. I have performed consulting services and

reviewed files and data associated with the Dikulushi and Kazumbula Projects between May

2009 and the present.

7. I take responsibility for Sections 1 (in part), 7, 8, 9, 10, 11, 12 and 14 of the Study and have

contributed to Sections 17.1 and 17.3 and the associated text in the summary, conclusions and

recommendations.




9. I am independent of Mawson West Ltd pursuant to section 1.4 of the Instrument.






Dated at Perth, Western Australia, on the 20 December, 2013.

Ian Glacken

Principal Consultant (Optiro Pty Ltd)

BSc (Hons) (Geology), FAusIMM(CP), MIMMM, CEng


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TURNER MINING AND GEOTECHNICAL PTY LTD

CERTIFICATE OF QUALIFIED PERSON – MIKE TURNER




1. My name is Mike Turner and I am a full-time employee of Turner Mining and Geotechnical Pty

Ltd of 3B Valley Road, Wembley Downs, WA, 6019, Australia.

2. I am a practising geotechnical and mining engineer and a Chartered Professional Fellow of the

AUSIMM (205399).

3. I am a graduate of Imperial College, London University with a BSc (Eng) (Hons) in Mining in 1979.

I also obtained a Master of Science in Mineral Production Management at the Royal School of

Miners in 1984 and a Chamber of Mines Certificate in Rock Mechanics in South Africa in 1987.

4. I have practiced my profession continuously since 1979, apart from a 12 month period from

1983/1984 during which I completed the MSc course at Imperial College.



6. I visited the Dikulushi Project from 10th to 13th December 2012, the most recent visit prior to

completion of the Study. I have performed consulting services and reviewed files and data

associated with Dikulushi between July 2003 and the present.

7. I am responsible for the geotechnical section in Section 16.










Dated at Perth, Western Australia, on 20 December 2013

Michael Harry Turner, MSC, DIC, BSc (Eng) (Hons) (Mining), ARSM, FAusIMM (CP), RPEQ

Director, Turner Mining and Geotechnical Pty Ltd


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KNIGHT PIESOLD

CERTIFICATE OF QUALIFIED PERSON – DUNCAN JOHN GRANT-STUART

As a reviewer of the report entitled “Technical Report on the Dikulushi Underground Project, Democratic Republic of Congo” (the Study) dated 12 December 2013, on the Underground Project of Mawson West Limited, I hereby state:

1. My name is Duncan John Grant-Stuart and I am a full time Engineer with the firm of Knight Piesold (Pty) Limited of PO Box 221, Rivonia, 2128, South Africa.

2. I am a practising Civil Engineer and member of the Institution of Civil Engineers (UK)(MICE) and am registered with the Engineering Council (UK) (C.Eng) and the Engineering Council of South Africa (PR.Eng).

3. I am a graduate of the University of the Witwatersrand with a BSC (Eng) degree completed in 1976. 4. I have practiced my profession continuously since 1976. 5. I am an “Independent” and “qualified person” as the terms are defined in National Instrument 43-

101 (Standards of Disclosure for Mineral Projects) (the “Instrument”). 6. I visited the Dikulushi Project property. I have performed consulting services and reviewed files and

data supplied by Mawson West Ltd in 2011. 7. I reviewed Section 17 of the Study, as well as the associated text in the summary, conclusions and

recommendations. 8. I am responsible for the geotechnical section and the associated text in the summary conclusions

and recommendations. 9. I am not aware of any limitations imposed upon my access to persons, information, data or

documents that I consider relevant to the subject matter of the study. 10. I am not aware as at 12 December 2013, the effective date of the Study, of any material fact or

material change with respect to the subject matter of the Study, which is not reflected in the Study, the omission of which would make the Study misleading.

11. I am independent of Mawson West Ltd and AMC SARL pursuant to section 1.4 of the Instrument. 12. I have read the National Instrument and Form 43-101F1 (the “Form”) and the Study has been

prepared in compliance with the Instrument and the Form. 13. I do not have nor do I expect to receive a direct or indirect interest in the Dikulushi property of

Mawson West Ltd, and I do not beneficially own, directly or indirectly, any securities of Mawson West Ltd or any associate or affiliate of such company.

Dated at Rivonia, South Africa, on 20 December 2013

Duncan John Grant-Stuart

PR Eng 900014

C.Eng

Technical Consultant


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SEDGMAN

SEDGMAN

CERTIFICATE OF QUALIFIED PERSON – PETER HAYWARD




1. My name is Peter George Hayward and I am Senior Process Engineer with the firm Sedgman

Limited, Suite 3, 3 Craig Street, Burswood, 6100.

2. I am a practicing Metallurgist and a Fellow of the Australian Institute of Mining and Metallurgy.

3. I am a graduate of the Ballarat Institute of Advanced Engineering and hold a Diploma of

Metallurgy.

4. I have practiced my profession continuously since February 1974.



6. I have personally visited the Dikulushi property in February 2012. I have reviewed files and data

supplied by Mawson West Ltd in September 2011 and in January 2013.

7. I have contributed to the Sections 13 and 17.

8. I am not aware of any limitations imposed upon my access to persons, information, data or

documents that I consider relevant to the subject matter for the Study (apart from as indicated

in the text).

9. I am not aware as at 12 December 2013, the effective date of the Study, of any material fact or

material change with respect to the subject matter of the Study, which is not reflected in the

Study, the omission of which would make the Study misleading.







Dated at Perth, Western Australia, on 20 December 2013

Peter Hayward

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