mawson west limited

Mawson West Limited Technical Report on the Dikulushi Open Pit Project, Democratic Republic of

Congo – September 16, 2011

Technical Report on the Dikulushi Open Pit Project, Democratic Republic of Congo – September 16, 2011

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Doc Ref:

110916_Optiro_Dikulushi_43-

101_Reserve_Draft_Base.docx

Print Date: 16 September 2011

Number of copies: 2

Optiro: 1

Mawson West Limited: 1

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: David Gray BSc Hons (Geology),

MAusIMM, PrSciNat

Signature:

Date: 16 September 2011

Principal Reviewer: Rick Stroud FAusIMM

Contributing author: Andrew Law FAusIMM

Signature:

Date: 16 September 2011

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 entire Report 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, 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 Open Pit Project, Democratic Republic of Congo

A technical report on the open pit cutback

Prepared for

Mawson West Limited

Authors

David Gray Principal Consultant, Optiro Pty Ltd BSc Hons (Geology), MAusIMM, PrSciNat

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

Date of report: 16 September 2011


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TABLE OF CONTENTS

1. SUMMARY 11

1.1. LOCATION 11

1.2. OWNERSHIP 12

1.3. MINERALISATION 12

1.4. MINERAL RESOURCES & RESERVES 12

1.5. METALLURGICAL 13

1.6. ENVIRONMENTAL 13

1.7. CONCLUSIONS AND RECOMMENDATION 14

2. INTRODUCTION 15

2.1. SCOPE OF THE REPORT 15

2.2. AUTHORS 15

2.3. PRINCIPAL SOURCES OF INFORMATION 16

2.4. SITE VISIT 17

2.5. INDEPENDENCE 18

2.6. ABBREVIATIONS AND TERMS 18

3. RELIANCE ON OTHER EXPERTS 25

4. PROPERTY DESCRIPTION AND LOCATION 26

4.1. DEMOGRAPHICS AND GEOGRAPHIC SETTING 26

4.2. PROJECT OWNERSHIP 26

4.3. PROJECT LOCATION 26

4.4. THE PROJECT TENEMENT AREA 26

4.5. ENVIRONMENTAL PERMITS 29

5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY 30

5.1. ACCESS 30

5.2. SITE TOPOGRAPHY, ELEVATION AND VEGETATION 30

5.3. CLIMATE, PHYSIOGRAPHY, LOCAL RESOURCES AND INFRASTRUCTURE 30

5.4. SURFACE RIGHTS 30

5.5. SITE INFRASTRUCTURE 31

5.5.1. WATER SUPPLY 31

5.5.2. POWER SUPPLY 31

5.5.3. MINE PERSONNEL 32

5.5.4. TAILINGS STORAGE FACILITY 32

5.5.5. ADMINISTRATION AND PLANT SITE BUILDINGS 32

5.5.6. ACCOMMODATION 32

5.5.7. COMMUNICATIONS 33

5.5.8. MOBILE EQUIPMENT 33

5.5.9. SECURITY 33


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6. HISTORY 34

7. GEOLOGICAL SETTING AND MINERALISATION 35

8. DEPOSIT TYPES 36

9. EXPLORATION 37

10. DRILLING 38

11. SAMPLE PREPARATION, ANALYSIS AND SECURITY 39

12. DATA VERIFICATION 40

13. MINERAL PROCESSING AND METALLURGICAL TESTING 41

13.1. INTRODUCTION 41

13.2. ANVIL MINING TEST WORK 41

13.2.1. EARLY TEST WORK 41

13.2.2. LATER TEST WORK 42

13.1 PLANT OPERATIONAL RESULTS 46

13.2 METALLURGICAL PROPERTIES OF THE CUTBACK ORE 47

14. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES 49

14.1. GEOLOGICAL AND MINERALISATION MODELS 50

14.2. DRILL DATA FOR MINERAL RESOURCE MODELLING 51

14.3. DATA VALIDATION 53

14.4. DATA PREPARATION FOR MODELLING 53

14.5. DATA COMPOSITING 54

14.6. STATISTICS 55

14.7. SPATIAL STATISTICS 55

14.8. BLOCK MODEL 58

14.9. DENSITY ESTIMATES IN THE BLOCK MODEL 59

14.10. DETERMINATION OF TOP CUTS 59

14.11. GRADE ESTIMATION 59

14.12. ORDINARY KRIGING INTERPOLATION 59

14.13. MODEL VALIDATION 60

14.14. MINERAL RESOURCE CLASSIFICATION 62

14.15. RESOURCE TABULATION AND INVENTORY 63

14.15.1. GRADE TONNAGE CURVES 63

14.16. MINERAL RESOURCE ESTIMATE COMPARISONS 64

15. MINERAL RESERVE ESTIMATES 67

15.1. PIT OPTIMISATION 67


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15.1.1. OPTIMISATION PARAMETERS 67

15.1.2. OPTIMISATION RESULTS 69

15.2. PIT OPTIMISATION SENSITIVITY ANALYSIS 72

15.3. MINE DESIGN 73

15.4. CUT-OFF GRADE CRITERIA 74

15.5. MINING INVENTORIES 75

15.6. MINING RECOVERY AND DILUTION 75

15.7. RESERVE CLASSIFICATION 75

15.8. MINERAL RESERVES TABULATION 77

16. MINING METHODS 78

16.1. MINING STRATEGY 79

16.1.1. CONTRACTORS FLEET 82

16.2. OTHER MINING FLEET 83

16.3. GEOTECHNICAL 83

16.3.1. DATA 83

16.3.2. GEOTECHNICAL DOMAINS 84

16.3.3. SLOPE GUIDELINES 85

16.3.4. POTENTIAL FAILURES 89

16.3.5. OTHER FACTORS AFFECTING STABILITY 91

16.3.6. MAPPING, MONITORING AND ADDITIONAL DATA 93

16.4. IN-PIT SUPPORT REQUIREMENTS 93

16.4.1. EXISTING UNDERGROUND EXCAVATIONS 94

16.4.2. MAJOR STRUCTURES 97

16.4.3. CABLE BOLTS AND CATCH FENCES 98

16.5. ROM PAD DESIGN 99

16.6. WASTE DUMP DESIGN 100

16.7. SURFACE WATER MANAGEMENT 101

17. RECOVERY METHODS 103

17.1.1. PLANT FLOWSHEET 103

17.1.2. TAILINGS STORAGE FACILITIES (TSF) 104

17.1.3. PROCESSING STATISTICS 105

18. PROJECT INFRASTRUCTURE 107

18.1. SURFACE FACILITIES 107

18.2. POWER 108

18.3. PROCESS WATER SUPPLY 108

19. MARKET STUDIES AND CONTRACTS 111

19.1. MARKETS 111

19.2. CONTRACTS 111

20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR

COMMUNITY IMPACT 113


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21. CAPITAL AND OPERATING COSTS 114

21.1. CAPITAL COST ESTIMATE 114

21.2. OPERATING COST ESTIMATE 114

21.2.1. MINING OPERATING COST 115

21.3. PROCESSING OPERATING COSTS 117

21.3.1. OVERHEAD OPERATING COST 117

21.3.2. CAPITAL EXPENDITURE 117

21.3.3. OPERATING COSTS 117

22. ECONOMIC ANALYSIS 118

22.1. MINING SUMMARY 118

22.1.1. SENSITIVITY ANALYSIS 120

22.2. PAYBACK 121

22.3. MINE LIFE 122

22.4. TAXATION 122

23. ADJACENT PROPERTIES 123

24. OTHER RELEVANT DATA AND INFORMATION 124

25. INTERPRETATION AND CONCLUSIONS 125

26. RECOMMENDATIONS 126

27. REFERENCES 127

28. CERTIFICATES 129

TABLES

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

off grade 12

Table 1.2: Dikulushi Mineral Reserve statement as at August 2011, using a 1.0% copper cut-off

grade 13

Table 2.1 Glossary of terms 19

Table 4.1 Mawson West Limited tenement schedule 28

Table 13.1 Details of Dikulushi drillcore used in Mintek metallurgical testing 41

Table 13.2 Head grades of chalcocite composites 43

Table 13.3 Relative abundance of significant minerals 43

Table 13.4 Comminution test work results 44

Table 13.5 Effect of grind size on flotation performance (high grade chalcocite) 44

Table 13.6 Effects of collector addition on flotation performance (high grade chalcocite) 44


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

chalcocite) 44

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

chalcocite) 45

Table 13.9 Effect of grind size and Eh level on flotation performance (Pb/Zn rich chalcocite)45

Table 13.10 Head grades of chalcocite composites 46

Table 13.11 Locked cycle flotation test results 46

Table 13.12 Dikulushi processing summary (February 2007 – April 2008) 48

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

off grade 50

Table 14.2 Domain codes for Dikulushi modelling 54

Table 14.3 Summary statistics for copper % and silver g/t per domain 55

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

and angle3 about axis 3 (Z) 57

Table 14.5 Dikulushi - top cuts per domain 59

Table 14.6 Mean statistics per domain comparing model estimates with data values 60

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

August 2011 63

Table 14.8 Comparison of 2009 and 2007 Dikulushi Mineral Resource estimates 65

Table 15.1 Pit Optimisation Parameters 67

Table 15.2 Pit Design Parameters 73

Table 15.3 Pit design fleet parameters 74

Table 15.4 Dikulushi mined material 76

Table 15.5: Dikulushi Mineral Reserve statement as at August 2011 at a 1% copper cut-off

grade. 77

Table 16.1: Major Equipment List – Dikulushi Open Pit Project 82

Table 16.2 Slope design guidelines 86

Table 16.3 Weathering depth from new holes 86

Table 17.1 Dikulushi Processing Summary relevant to ore to be mined in the pit cutback105

Table 17.2 Processing statistics for the LG material completed by MWL 106

Table 21.1 Capital cost estimates 114

Table 21.2 Drill and blast unit costs 115

Table 21.3 Load and Haul unit costs 116

Table 21.4 Operating costs 117

Table 22.1 Dikulushi Mining and Financial Summary 118

Tables 22.2 to 7 Sensitivity analysis on the cash flow forecast for the open pit cutback and

treatment at Dikulushi 120

FIGURES

Figure 1.1 Locality plan of the Dikulushi Open Pit Project 11

Figure 4.1 Exploration Licences of the Dikulushi copper silver project 27

Figure 4.2 Dikulushi mine infrastructure within the PE 606 28

Figure 5.1 Dikulushi airstrip and the G1 Charter plane provides safe staff transportation to and

from site 31

Figure 13.1 Underground sources of ore presented in Table 13.2 48


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

50

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

mineralisation lenses and current drilling 51

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

from the pit have been excluded 52

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

samples within a common area 53

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

length 54

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

the results of robust domaining 56

Figure 14.7 Variogram models for copper % across the FW zone of mineralisation. 58

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 61

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 mineralisation61

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

categories 62

Figure 14.11 The grade tonnage curves for the combined Measured and Indicated Mineral

Resources 64

Figure 14.12 A waterfall chart of cumulative Mineral Resource changes from 2007 to 2009 66

Figure 15.1 Pit optimisation plot (undiscounted) 69

Figure 15.2 East-west section 70

Figure 15.3 North-south section 70

Figure 15.4 North-south section 71

Figure 15.5 Oblique view showing fault planes 71

Figure 15.6 Final pit design 72

Figure 15.7 Pit optimisation sensitivity analysis plot (discounted @ 10%). 73

Figure 15.8 Mineral Resource and Mineral Reserve classification 76

Figure 16.1 The existing Dikulushi open pit in 2011 78

Figure 16.2 The cutback stages 80

Figure 16.3 The cutback width at the 880Mrl. The white outline is the existing pit. Red lines

show the old underground workings and the ore blocks are in blue. 81

Figure 16.4 Location of geotechnically logged drillholes 84

Figure 16.5 Pit slope design domains 85

Figure 16.6 Domain E Wedge Potential 88

Figure 16.7 Factor of Safety Sensitivity Analysis, Domain E, 75° Bench Face Angle 89

Figure 16.8 Run-off control domains 92

Figure 16.9 Underground development holings in the North Wall of the pit design 95

Figure 16.10 Underground development holings in the South Wall of the pit design 95

Figure 16.11 Stope Intersection Indicating Rockbolt support. 96

Figure 16.12 Area of pit wall requiring rockbolt support to prevent unravelling 96

Figure 16.13 Plan view indicating development associated with the 870 to 830 Ventilation Rise

97

Figure 16.14 North wall cable bolts and catch fence 98

Figure 16.15 North wall bolting patterns. 98

Figure 16.16 South wall catch fence on 830m RL 99


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Figure 16.17 Location of waste dump relative to expanded pit 100

Figure 16.18 Surface water management – general arrangement 102

Figure 17.1 Dikulushi Plant flow diagram 103

Figure 18.1 On-site office facilities at Dikulushi 107

Figure 18.2 Lake Newton 108

Figure 18.3 Average water balance 110


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1. SUMMARY

Mawson West Limited’s (MWL) Dikulushi Open Pit Project (the Project) is located in the Katanga

Province of the Democratic Republic of Congo (DRC). The pit comprises Mineral Resources from the

main Dikulushi deposit’s “Footwall” zone, which has a 230 m strike length and true widths up to 25

m. The pit is planned as a cutback extension of the existing Dikulushi pit left by Anvil Mining Limited

(Anvil) during its tenure of the Dikulushi deposit. MWL has completed open pit studies to mine out

the remaining high grade resource at the Dikulushi deposit and is in the process of developing a pre-

feasibility study in order to access the mineralisation below the current open pit.

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).

Figure 1.1 Locality plan of the Dikulushi Open Pit Project


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1.2. OWNERSHIP

The Dikulushi mine is part of 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 100% owned by Anvil Mining Congo SARL which is in the process

of being renamed to CMCC SARL (CMCC). Mawson West Investments Ltd, a wholly owned subsidiary

of Mawson West Limited holds 90% of the issued capital of CMCC, 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 and LG ROM stockpile form part of the PE. This report presents technical

information on the Dikulushi open pit cut back only.

1.3. MINERALISATION

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

of the Dikulushi deposit has been mined out.

1.4. MINERAL RESOURCES & RESERVES

The current Mineral Resources of the Dikulushi ore body have been modelled using a mineralisation

based interpretation of copper. A block model estimate was completed in May 2009 by David Gray

of Optiro and was depleted in August 2011 according to updated surveyed volumes of historical

mining. The resulting Mineral Resources are stated for a 1.0% copper cut-off grade as per Table 1.1

below.

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

Mineral Reserves are shown in Table 1.2 and are stated for a 1.0% copper cut-off grade. Mineral

Resources are 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 1.2: 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 66633 2.8 184719 7.27% 207

Probable 127790 2.8 354258 5.51% 169

Total Proven and Probable Reserves 194423 2.8 538977 6.12% 182

1.5. METALLURGICAL

Several metallurgical test work 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 Dikulushi, and have subsequently been compared against actual production

results during the period of operation by Anvil.

The most recent metallurgical test work was managed by Sedgman Metals, a metallurgical

consulting company of Perth, Western Australia. Test work was completed at AMDEL Laboratories

in Perth. Metallurgical test work was carried out previously by Anvil on the main resource ore body.

Additional test work was reported on in June 2004 by Independent Metallurgical Laboratories IML

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

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

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

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

results during operation by Anvil were higher.

The float plant at Dikulushi operated from 2004 to 2008, fed with high grade ore from the open pit

and underground mine, giving recoveries of 93% copper and 90% silver, producing a concentrate

with 55% copper and 2,100 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 recoveries previously experienced for the fresh ore

from the open pit will be achieved.

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

50% copper.

1.6. ENVIRONMENTAL

An ESIA and EMP was lodged in 2003 and was completed by African Mining Consultants of Kitwe,

Zambia, an environmental company licensed to work and report in the DRC.

MWL has lodged $368,409.50 as an Environment bond. This financial guarantee is a contribution

towards environmental rehabilitation costs for the Dikulushi mine.

An updated ESIA was completed for the Project and will be submitted as part of the environmental

requirements of the mining lease.


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1.7. 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

developing the Dikulushi deposit via an open pit cutback. Since this was previously an operating

open pit the remaining ore zones have the same risks and are somewhat mitigated for the

mineralogy, metallurgical properties and the processing aspects. There will remain risks for the

mining operations and constant recognition of changing conditions will need to be ensured and

appropriate changes made as mining progresses. Geotechnical knowledge will increase with the

physical mining activities and a better understanding of the ground conditions will be established.

There is likely to be continued resource 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 implement the cut-back, and any other required

development, MWL intends to continue processing the LG stockpile and open pit mining of satellite

resources. In addition MWL is currently in the process of defining additional deposits within 50 km

of the Dikulushi plant.


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2. INTRODUCTION

2.1. SCOPE OF THE REPORT

Mawson West Limited (MWL) commissioned Optiro Pty Ltd (Optiro) in May 2011 to review the

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

copper-silver Mineral Reserves at the Dikulushi deposit based on this 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” of 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).

The technical report has been written to provide the market with an update on Mineral Resources

and Reserves for the Dikulushi Open Pit Project (the Project) which is now entering an expansion

cut-back of the original Anvil open pit.

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

otherwise stated.

This report presents technical information relevant to the Project’s open pit cut back only.

2.2. AUTHORS

The key authors for compiling this report are:

Mr David Gray is the principal author and Qualified Person and takes overall responsibility

for this report. Mr Gray, of Optiro, is a professional geologist and has a BSc (Hons) degree

(1988) from Rhodes University, South Africa. He has more than 20 years experience in

exploration and mining geology. Mr Gray is a Member of the Australasian Institute of

Mining and Metallurgy (AusIMM) and a member of the South African Council for Natural

Scientific Professions (PrSciNat, 400018/4) and has the relevant qualifications, experience

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

Instrument 43-101. Mr Gray has visited the Dikulushi deposits in November 2010 and has

generated 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.

Mr Andrew Law is a Qualified Person and takes responsibility for Mineral Reserves

estimation portion of 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 28 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,


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experience and independence to be considered as a “Qualified Person” as defined in

Canadian National Instrument 43-101. Mr Law has not visited the Dikulushi deposits and at

this stage does not intend to visit the deposit. He has however reviewed previous NI 43-101

reports generated by Anvil and Optiro, for MWL, and has held discussions with both David

Gray of Optiro, and staff members from 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 the pre-feasibility study by MWL.

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

Dikulushi deposit. 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

David Gray Principal Consultant, Optiro Pty Ltd Principal Qualified Person

Andrew Law Director – Mining, Optiro Pty Ltd Qualified Person and author of

chapters 15, 16, 19, 20, 21, 22, 24, 25,

26.

Nick Hunt-Davies Principal Consultant, Optiro Pty Ltd Contributing author of chapters 15,

16, 19, 20, 21, 22, 24, 25, 26.

Rick Stroud Director, Optiro Pty Ltd Peer review

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

geotechnical submission in chapter 16

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

storage facilities in Chapter 17

Peter Shepard SRK Consulting Hydrological, QP and author of

hydrological submission in chapter 16

and 18

Ray Creese Sedgman Ltd Metallurgical, QP and input into

chapter 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 covering Mineral

Resources at Dikulushi. This pre-feasibility information was provided to Optiro by MWL. The

Mineral Resource information has been provided from the previously submitted NI 43-101 Technical

Report, by Optiro, on the Dikulushi Project, Democratic Republic of Congo, February 3, 2011 and

subsequently revised March 7, 2011. The Mineral Resource has recently undergone a review based

on recently supplied and updated underground survey information and revised cut-off grades.

In summary, the following are primary data sources:

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

February 3, 2011 and subsequently revised March 7, 2011

2. Historical and current production and processing data


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3. A pre-feasibility study 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

Mr Jan Dharma-bandu Mawson West Ltd Mining

Adam Anderson 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 Haywood Sedgman Ltd Metallurgical

Andries Strauss Knight Piesold Consulting Tailings storage facilities

Glen Zamudio Mawson West Ltd Finance, Marketing and Legal

2.4. SITE VISIT

Mr David Gray 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 and the LG ROM Stockpile

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

observe and inspect operational activities related to processing the LG stockpile

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. Site visits

have been carried out by the following persons:

Name Company Section Date of Visits

David Gray Optiro Resource NI 43-101 November 2010

Mr Jan Dharma-bandu Mawson West Mining Various as employee of MWL

Adam Anderson Mawson West Geology Various as employee of MWL

Mike Turner Turner Mining and

Geotechnical Pty Ltd

Geotechnical August 2008

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

Ray Creese Sedgman Ltd Metallurgical Twice during 2006 to 2008

Peter Shephard SRK Consulting Hydrology, Water Management Once during 2007

Glen Zamudio Mawson West Ltd Finance, Marketing and Legal Various as employee of MWL


P a g e | 18

Mr Andrew Law has not visited the Dikulushi deposits and at this stage there is no intention for him

to visit the deposit. He has however reviewed previous NI 43-101 reports generated by Optiro for

MWL and has held discussions with both David Gray, of Optiro, and staff members from MWL.

Mr Andrew Law has reviewed all sections of the “Pre-Feasibility” study generated by various

Qualified Persons, many of whom were independent of Mawson West. The report was collated by

MWL.

2.5. INDEPENDENCE

Neither Mr David Gray nor Mr Andrew Law, 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 | 19

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 | 20

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

EPCM Engineering, Procurement, Construction and Management

Equator Principles A financial industry benchmark for determining, assessing and managing social and

environmental risk in project development

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


P a g e | 21

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

GST Goods and Services Tax

ha Hectares

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 Thousands of grams

kL Thousands of litres

km Thousands of metres

kt Thousands of tonnes

kV Thousands of volts

kW Thousands of watts

kWh Thousand watt hours

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


P a g e | 22

structure

MAX Maximum

mbgl Metres below ground level

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)


P a g e | 23

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

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


P a g e | 24

TN True North

TOFR Top of fresh rock

tpa Tonnes per annum

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 | 25

3. RELIANCE ON OTHER EXPERTS

This technical report has been prepared and approved under the supervision of Mr. David Gray,

Principal Consultant, Optiro Pty Ltd., and Mr Andrew Law, Director Mining, Optiro Pty Ltd. Mr David

Gray, who is the principal author of the report, and Mr Andrew Law are both independent Qualified

Persons as defined in National Instrument 43‐101.

In preparing this report, the Qualified Persons have relied upon 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 Section 4 and 20 of this report.

Environmental permit and bond information which is relevant to Section 4 and Section 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 information leading to the assessed head grade of the copper-silver

concentrate produced from treatment of the mined ore which is relevant to Section 13 and

17 of this report.

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

sections 15, 16, 21, and 22.

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

authenticity of the information provided and drafts of this report were provided to MWL with a

request to identify any material errors or omissions prior to filing.


P a g e | 26

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 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 66 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 part of 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 owned 100% by Anvil Mining Congo SARL (which is in the process

of being renamed CMCC SARL) (“CMCC”). Mawson West Investments Ltd a wholly owned subsidiary

of Mawson West Limited, holds 90% of the issued capital of CMCC, the remaining 10% is held by the

Dikulushi – Kapulo Foundation (NPO).

4.3. PROJECT LOCATION

The Project is located within the Katanga Province in the southeastern 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).

4.4. THE PROJECT TENEMENT AREA

CMCC holds title to the Dikulushi mine and surrounding exploration tenements through the

Dikulushi mining convention. Under the Mining Convention the exploration tenements known as

“PR’s” were issued for a five year period and are renewable a further three times, each time for a

period of five years. The PR’s shown in Table 4.1 and Figure 4.1 below were first granted on the


P a g e | 27

22 May 2001 and currently showing as expiring in 21 May 2011. The PR’s were renewed for a

further five years on the 22 May 2011. The company relinquished PR’s 1695 and 1704 during the

renewal process. CMCC holds 19 Exploration Permits and one exploitation permit under the

Dikulushi Mining Convention, covering 7,283km². CMCC holds title to the Kapulo exploration

tenements through the Dikulushi Mining Convention.

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

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

exploitation in respect of each mine are for a period of 20 years from the respective dates of

commencement of production from each mine.

Figure 4.1 Exploration Licences of the Dikulushi copper silver project

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

by Ministerial Decree on 31 January 2002. The PE covers an area of 40.77 km2 over the Dikulushi

mine area (Figure 4.1 and Figure 4.2).


P a g e | 28

Table 4.1 Mawson West Limited tenement schedule

Tenement Schedule

Project Group Entity Permit No. Area km² Type Granted Expiry

Dikulushi AMC PE606 40.77 Mining 31-Jan-02 30-Jan-22

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

Kapulo AMC PR1684 399.1 Exploration 12-Apr-11 11-Apr-16


Kapulo AMC PR1686 395.7 Exploration 22-May-11 21-May-16




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


Kapulo AMC PR1697 398.7 Exploration Held Held










Total Area 7,283

Figure 4.2 Dikulushi mine infrastructure within the PE 606


P a g e | 29

4.5. ENVIRONMENTAL PERMITS

An Environmental Impact Assessment (EIA) was completed by African Mining Consultants of Zambia

in April 2003, along with an Environmental Management Plan (EMP), which have been approved by

the DRC Government. The EMP 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 $368,409.50 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 updated EAP was completed for the Project in July 2011 and will be

submitted as part of the requirements

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

4. wet/dry, min/max temperatures

5. rainfall

6. dust exposure

7. noise levels

8. ground and surface water quality

9. groundwater levels

10. Tailings Dam piezometer water levels

11. 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.


P a g e | 30

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 refurbished 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. An all weather airstrip is located at the Dikulushi

mine and charter flights 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 340 t capacity barge owned by CMCC; the water journey takes 5 hours. Access from

Kilwa port to the mine is via a 54 km refurbished gravel road and takes approximately 1 hour by light

vehicle.

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.

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. Weather data

collected at Dikulushi over 3 full years from 2006 to 2008 shows an average annual rainfall of 1,127

mm. 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 18°C

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

The wet season does not affect mining or processing operations at Dikulushi but does inhibit

exploration activities, and access to some areas within the PR with flooded roads and rivers and

terrain becoming difficult to access with vehicles. Labour is sourced locally for camp, geological and

drilling assistant type work.

5.4. SURFACE RIGHTS

The Dikulushi mine is based on Exploitation Licence (PE606) granted on 31 January 2001. The lease

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


P a g e | 31

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

4. the staff village

5. the airstrip

6. the process water dam

7. the tailings storage facility.

These items of infrastructure are depicted in Figure 4.2. This infrastructure was in place for the

previous operations under Anvil and is well established and sufficient in size for current and planned

requirements.

Figure 5.1 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.

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.


P a g e | 32

The existing power station at Dikulushi comprises the following generators: 1 x 2.0 MW Caterpillar, 1

x 1.6 MW Caterpillar, 4 x 0.8 MW Mirrlees for a total capacity of 6.8 MW. The current power

demand is in the order of 1.8 MW and only the 2.0 MW Cat, 1.6 MW Cat and 2x Mirrlees are being

utilised on rotation. This is sufficient to supply the extra demand of 0.6-1.0 Megawatts for

dewatering purposes during the cut-back project.

CMCC 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 Saleem. CMCC has contacted a further supplier from Dar Es

Saleem whom would be able to supply fuel to Dikulushi. During the cutback project the demand for

diesel will increase to 1,200,000 l/month for a four to five month period. CMCC is regularly speaking

to suppliers to guarantee no interruptions in supply. Thus CMCC believes that it has mitigated the

risk of fuel supply by having a number of suppliers whom source fuel from different ports.

5.5.3. MINE PERSONNEL

As at December 2010, Dikulushi mine employed 270 people, of which 22 were expatriates. The

requirements for the cut-back of the open pit and other associated activities will require a total

workforce of 145 - 40 employees and 105 contractors.

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 the coarse portion reclaimed and retreated in recent

operations. The second TSF is dormant whilst the third is in use to accommodate the tailings

resulting from the treatment of the HMS material and other low grade stockpiles. The third TSF has

been reviewed for extended use beyond its current life. This will be built up to accommodate

tailings resulting from the open pit cut back mining 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, a warehouse, mining equipment and

maintenance garages, mechanical workshops and a service area with access pit for inspection and

repair of vehicles.

There is an infirmary on site and a hospital at Kilwa, approximately 50 km from the mine. An assay

laboratory has been constructed on site in order to facilitate metallurgical, exploration and grade

control sampling.

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.


P a g e | 33

5.5.7. COMMUNICATIONS

Mobile phone coverage is possible 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.

5.5.8. MOBILE EQUIPMENT

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

vehicles (including an ambulance), quad bikes, light trucks, forklifts, buses and generators.

5.5.9. SECURITY

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

and explosives.


P a g e | 34

6. HISTORY

For information on the history of the project/property refer to item 8 in the NI 43-101 Technical

Report on the Dikulushi Project, Democratic Republic of Congo, February 3, 2011 and subsequently

revised March 7, 2011.


P a g e | 35

7. GEOLOGICAL SETTING AND MINERALISATION

For information on the Geological Setting and the Mineralisation refer to item 9 and item 11 of the

NI 43-101 Technical Report on the Dikulushi Project, Democratic Republic of Congo, February 3,

2011 and subsequently revised March 7, 2011.


P a g e | 36

8. DEPOSIT TYPES

For information on the Deposit Types refer to item 10 of the NI 43-101 Technical Report on the

Dikulushi Project, Democratic Republic of Congo, February 3, 2011 and subsequently revised March

7, 2011.


P a g e | 37

9. EXPLORATION

For information on the Exploration of the project/property refer to item 12 of the NI 43-101

Technical Report on the Dikulushi Project, Democratic Republic of Congo, February 3, 2011 and

subsequently revised March 7, 2011.


P a g e | 38

10. DRILLING

For information on the drilling used for the generation of the resource refer to item 13 of the NI 43-

101 Technical Report on the Dikulushi Project, Democratic Republic of Congo, February 3, 2011 and

subsequently revised March 7, 2011.


P a g e | 39

11. SAMPLE PREPARATION, ANALYSIS AND SECURITY

For information on the Sample Preparation, Analysis and security used for the generation of the

resources refer to item 15 of the NI 43-101 Technical Report on the Dikulushi Project, Democratic

Republic of Congo in February 3, 2011 and subsequently revised March 7, 2011.


P a g e | 40

12. DATA VERIFICATION

For information on the data verification used for the generation of the resources refer to item 16 of

the NI 43-101 Technical Report on the Dikulushi Project, Democratic Republic of Congo, February 3,

2011 and subsequently revised March 7, 2011.


P a g e | 41

13. MINERAL PROCESSING AND METALLURGICAL TESTING

13.1. INTRODUCTION

Historically Anvil has completed a significant amount of test work for Dikulushi, and a summary of

this work is presented below. Relevant operational data from the Dikulushi processing plant is also

tabulated.

As the cut back ore will be mined from the same, or close to the same areas as the ore previously

treated at or below the current pit floor, it is not unreasonable to expect that it will exhibit similar

metallurgical characteristics during processing through the existing Dikulushi Processing plant.

13.2. ANVIL MINING TEST WORK

13.2.1. EARLY TEST WORK

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 test

work reports and as such cannot verify the information in this sub-section.

A significant amount of metallurgical test work 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 test work 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 test work 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 test work 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 test work 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.

Table 13.1 Details of Dikulushi drillcore 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


P a g e | 42

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 gray 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 test work 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 -

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 TEST WORK

Additional test work was performed by Independent Metallurgical Laboratories (IML) in Perth during

2003. The related test work 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 test work:


P a g e | 43

High Grade Chalcocite

Disseminated and Low Grade Chalcocite

Lead & 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.

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.


P a g e | 44

Table 13.4 Comminution test work 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 P80’s 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 140g/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.

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


P a g e | 45

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

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%.


P a g e | 46

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 headfeed 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

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%.

TEST WORK 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 test work 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.

13.1 PLANT OPERATIONAL RESULTS

Mawson West 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 1659 g/t silver. The


P a g e | 47

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 ore and HMS tails. The recoveries vary between 60 - 70% for the low grade and 50 – 65% for

the HMS tails. Concentrate grades average 45% copper and 1,100g/t silver. This information on

recent operations has been provided by Mawson West. Sedgman has not been able to review the

data relating to Mawson West operations to verify the accuracy of these statements.

13.2 METALLURGICAL PROPERTIES OF THE CUTBACK ORE

The Dikulushi deposit was mined and processed by Anvil Mining 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 Mining’s monthly production reports, for the period where

ore from at or below the existing pit floor was being processed have been summarised in Table

13.12 to demonstrate the metallurgical response of this material.

Table 13.12 shows that ore from above and below the crown pillar achieved flotation recoveries

around 90% for both copper and silver from February 2007 to April 2008. The average treatment

rate on an annualised basis during this period was 365,317 tpa due to constraints associated with

underground mining.

Figure 13.1 shows the RL’s relative to the resource which is targeted by the cutback and the

locations of UG drives where material in Table 13.2 would have been sourced from.

As the cutback ore will be mined from the same or close to the same areas as the ore at or below the

present pit floor, it is not unreasonable to expect that it will exhibit similar metallurgical

characteristics. However concentrate recoveries will be marginally reduced as the planned

treatment rate is 500,000 tpa (or 1370 tpd). Flotation residence time will consequently be reduced

by approximately 27% compared to the operating period February 2007 to April 2008.

Reviewing historic operating data it can be seen that when average monthly treatment rates were

equivalent to around 1,370 tpd (Jan-06, Feb-06, Aug-06) the copper recovery was approximately 86-

87%. Based on historic production data, flotation grades should not be significantly affected.


P a g e | 48

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.

Figure 13.1 Underground sources of ore presented in Table 13.2


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14. MINERAL RESOURCE AND MINERAL RESERVE

ESTIMATES

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 according to the latest

available survey data of the historical mined volumes and the updated pre-feasibility study cut-off

grades.

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 extend to the Kiaka Carbonates.

Since the October 2007 estimate an additional 23610 m of underground, infill and extensional

drilling has been 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 October 2007 estimate was updated in May 2009 and includes all available data as at the end of

November 2008, with no outstanding core logging, sampling or assay results remaining. Lower

confidence in the results from underground sludge (open hole) drilling resulted in their exclusion

from the latest estimate.

Dikulushi mineralisation (Figure 14.2, showing footwall mineralisation in green and hangingwall

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 lode is largely

absent below the base of the open pit.

Grade interpolation was undertaken for total copper percent (%) 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 | 50

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

The resulting Mineral Resource statement has been depleted for open pit and underground material

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 current 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

14.1. 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


P a g e | 51

of the overlying hangingwall. 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; Weathering was 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 extension of the

Footwall Fault zone.

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

14.2. DRILL DATA FOR MINERAL RESOURCE MODELLING

Drill data was stored using 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.


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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 metres. 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 true width 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

FW 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 | 53

Figure 14.4 Quantile Quantile (Q-Q) plot of Diamond (DD) drilled samples versus sludge drilled samples within a common area

14.3. 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.

14.4. 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 mineralization zone and


P a g e | 54

weathering profile, providing mineralised domain codes for estimation (Table 14.2). The coded

drillhole data was exported for subsequent geostatistical analysis and grade interpolation.

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

14.5. 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


P a g e | 55

14.6. STATISTICS

Statistical analyses of the data, including spatial statistics, were carried out using Snowden’s

Supervisor software. The statistical analysis of composite copper grades was undertaken within

each of the final domains and the summary results are tabulated 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 FW 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

14.7. SPATIAL STATISTICS

For Dikulushi, variography was analysed using composited 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


P a g e | 56

normal scores variograms were calculated for each of the principal axes of anisotropy

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


P a g e | 57

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 FW 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.

14.8. 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.

Dikulushi’s 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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14.9. 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.

14.10. DETERMINATION OF TOP CUTS

Top cuts were 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 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

14.11. 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.

14.12. 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 in order to limit grade smoothing. Due to the long drillhole intercepts

within the orebody estimates were limited to a maximum of 10 samples per drillhole.


P a g e | 60

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.

14.13. MODEL VALIDATION

The first pass of model validation included:

visual comparisons (Figure 14.8) of drillholes and estimated block grades

checks for negative 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 FW 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


P a g e | 62

14.14. 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


P a g e | 63

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, 2000).

14.15. 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

east by the Kiaka carbonates and is observed to pinch out to the west. 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 1.5% 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)

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

14.15.1. GRADE TONNAGE CURVES

The grade tonnage curves for the total Dikulushi Mineral Resource are presented in Figure 14.11.


P a g e | 64

Figure 14.11 The grade tonnage curves for the combined Measured and Indicated Mineral Resources

14.16. MINERAL RESOURCE ESTIMATE COMPARISONS

The May 2009 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 resources, 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 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.

These comparisons were carried out using the 1.5% cut-off resource as the 2007 resources were only

available at that cut-off grade.


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Table 14.8 Comparison of 2009 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

Measured & Indicated 262 2.80 733 6.82 186

Inferred 129 2.80 361 7.11 94

Total MII 391 2.80 1095 6.91 155

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

Measured & Indicated 385 2.83 1,089 7.13 182

Inferred 119 2.83 336 4.30 112

Total MII 504 2.83 1425 6.46 166

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%

Measured & Indicated -32% -1% -33% -4% 2%

Inferred 9% -1% 8% 65% -16%

Total MII -22% -1% -23% 7% -6%

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 21,186 copper tonnes from previous Indicated and

Inferred Mineral Resources to be upgraded to a Measured category.

Figure 14.12 illustrates the relative and cumulative change in copper tonnes between the 2007 to

2011 estimates. The 2011 Mineral Resource estimate has dropped by 14%, 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 14.12 A waterfall chart of cumulative Mineral Resource changes from 2007 to 2009


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15. MINERAL RESERVE ESTIMATES

The method adopted by MWL in the development of the mining reserves was to commence with an

optimisation of the resource block model based on set parameters. The output from this

optimisation is a series of economic “shells”. The shell selected for mine design is based on both

economic and strategic principals (such as the ability to change production rates according to the

economic climate).

The practical mine design applied to the optimised shell selected provides the volumes of material

that can be potentially economically mined.

15.1. PIT OPTIMISATION

A pit shell optimisation exercise was conducted for the Dikulushi pit using “NPV Scheduler” software.

This software uses industry standard techniques to identify an optimised pit shape for a given set of

physical and economic parameters

A number of geotechnical and operational considerations were required in the development of the

final pit shell. There was a previous failure of the existing pit wall caused by structural features in

the North wall. The pit optimisation was initially developed to excavate the North wall to the extent

of the identified fault in order to improve the stability of the wall. However, subsequent design

reviews have resulted in further geotechnical reviews which have resulted in a cable bolt and mesh

support regime for areas identified as potential failure spots.

The minimum cut-back width used for the final optimisation is 25 m to enable safe access for

trucking to and from the working face.

15.1.1. OPTIMISATION PARAMETERS

Error! Not a valid bookmark self-reference. show the parameters used in generating the optimised

pit shell for the Dikulushi deposit. This table covers Physical Mining, Processing, Cost and Revenue

parameters.

Table 15.1 Pit Optimisation Parameters

Parameter Unit

PHYSICALS

Limits

Mining – Total Movement Mtpa 18

Processing Rate Mtpa ore 0.5

Mining

Pit Slope (Weathered) ° 40

Pit Slope (Fresh)

@ 000 deg mine grid brg ° 40





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Parameter Unit

Mining Recovery % 95%

Mining Dilution % 15%

Processing

copper Recovery

Weathered % 70%

Transitional % 90%

Fresh % 90%

Silver Recovery

Weathered % 70%

Transitional % 90%

Fresh % 90%

CAPEX

Capital

Infrastructure & dewatering Equipment

USD M 1.6

Plant & Equipment USD M 5.10

Sustaining Capital USD M 2.0

OPEX

Mining

variable

Waste / ore USD / BCM mined avg 9.62

MCAF USD / BCM / 10m Bench avg 0.41

Processing

Variable

Ore Processing USD / t ore 37.29

Rehabilitation

Variable USD / BCM waste 0.03

Selling

Variable

copper Sales USD / t Cu 1,720

Silver Sales USD / t Ag -

Administration USD M / y 7.2

REVENUE

copper

Base Price USD / lb 3.50

USD / t 7,716

NSR % 96.75%

Royalties % -

Realised Price USD / lb 3.39

Siver

Base Price USD / oz 30.00

USD / g 0.96

NSR % 91%

Royalties % -

Realised Price USD / oz 27.30

FINANCIAL

Discount Rate % 10%


P a g e | 69

Selective mining is not considered feasible or necessary due to the “massive” nature of the orebody.

This approach resulted in the geological model being regularised for the mining model.

Regularisation is where the mineral grades are averaged over the mining block selected and is a

means of including internal waste dilution within these blocks. The block size used is normally

associated with a pre selected mining unit (SMU) being the block that can best be mined effectively

using the equipment chosen. As a result the optimisation was run unconstrained by grade and the

result was based on the economic mining of the material from within the resource block model. The

selected shell requires adjusting in order to fit a practical mine design and the volumes are also likely

to be adjusted.

15.1.2. OPTIMISATION RESULTS

NPV Scheduler software was used to produce an optimum pit shell for the above parameters and

based on Measured and Indicated Resources only and at a copper price of US$3.50/lb and a silver

price of US$30.00/oz, the optimum pit shell, based on the maximum un-discounted cash flow, for a

practical minimum cut-back width is pit shell 32. Pit shell 32 contains some 540k tonnes of ore at a

grade of 6.1% copper and 182g/t silver, for approximately 29,700t of recovered copper and

2.81M ounces of recovered silver. Some 20 million tonnes of waste are contained within the pit

shell with a stripping ratio of 37:1. The undiscounted operating cashflow, inclusive of capital and

start up costs, is $143 million. To ascertain the likely discounted cashflow derived from a realistic

mine production schedule, the average discounted cashflow at 10% discount was calculated and is

$116 million.

Figure 15.1 Pit optimisation plot (undiscounted)

On analysis of the optimum pit shell it was found that the eventual design would have to deviate

from the optimum pit shell to address:


P a g e | 70

A minimum safe cutback width

Mining to remove additional waste material from the faulted Northern wall

The main ramp which cannot cross the less stable northern wall and thus will switch-back

across the southern wall.

The following figures show the existing pit (green), the optimised shell (grey) and the final pit design

(blue). The ore is represented by the red blocks at the bottom of the pit.

Figure 15.2 East-west section

Figure 15.3 North-south section


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Figure 15.4 North-south section

Figure 15.5 shows an oblique view of the pit with the two fault planes in the northern wall.

Figure 15.5 Oblique view showing fault planes

Figure 15.6 shows a plan view of the final pit design where the switch-back access ramp can be seen

on the southern wall


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Figure 15.6 Final pit design

15.2. PIT OPTIMISATION SENSITIVITY ANALYSIS

A sensitivity analysis is carried out on the selected pit shell to determine how selected major

elements will affect the economic viability of the operation. The sensitivity analysis for this

estimation can be seen below (Figure 15.7).


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Figure 15.7 Pit optimisation sensitivity analysis plot (discounted @ 10%).

The results indicate that the project is most sensitive to the copper price and least sensitive to

changes in operating cost. This is investigated and reported in further detail in Section 22, Economic

Analysis.

15.3. MINE DESIGN

The basic pit design parameters for the extension of the existing Dikulushi pit are described in Table

15.2 and Table 15.3.

Table 15.2 Pit Design Parameters

Design Parameter Weathered Fresh

North West North East East South West

Bench Width (m) 5 4 5 5 5 5

Bench Height (m) 15 10 20 20 20 20

Bench Face Angle (°) 50 55 60 60 60 50

Inter-Ramp Slope Angle (°) 40 42 50 50 50 50

80

90

100

110

120

130

140

150

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

NP

V (

US

$ m

illio

n)

Percentage Change for Base Case

Dikulushi Copper ProjectSummary Project Sensitivity Analysis

Cu Price Ag Price Operating Costs


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Table 15.3 Pit design fleet parameters

Haulage Fleet Characteristics

Machine Payload (t) Optimum

Width Tyre

EH1700-3 90 6.5 27.00R49

CAT740 40 3.5 26.50R25

On the HW side of the orebody a 190t excavator (EX1900C) will be used down to the

960RL. On the FW side an 80t excavator will be used with 40t ADT’s

Footwall

Ramp From mRL To mRL Gradient Width (m) Capacity

805 850 1:8 9 Single CAT740

850 875 1:10 15 Single EH1700 or

dual CAT740

From mRL To mRL Gradient Width (m) Capacity

875 900 1:8 9 Single CAT740

900 968 1:8 15 Dual CAT740

968 Ground

level 1:10 15 Dual CAT740

Passing Bays every 20m in RL

HW Ramp From mRL To mRL Gradient Width (m) Capacity

875 900 1:10 15

EH1700

2 CAT740

900 950 1:10 15

EH1700

2 CAT740

950

Ground

Level 1:10 23 2 EH1700 or CAT740

Passing Bays every 20m in RL

The above design parameters used in conjunction with the Mining Strategy as outlined in Section

16.2, a staged cut back approach, has been used to design a practical pit based on the selected

optimised shell (pit shell 32).

15.4. CUT-OFF GRADE CRITERIA

The differentiation between ore and waste in the Dikulushi pit has been determined by a

profitability determination process within the NPV Scheduler software system which is the same

system used to develop the Loerch-Grossman pit optimisation shells.

The mining and financial parameters used in the pit shell optimisation are also used to assess the

profitability of each block of material within the block model and within the optimum pit shell. The

revenue and cost associated with each block is determined in turn and if the total associated cost is

lower than the net income, the material is sent to the mill as ore. For units of material where the


P a g e | 75

total cost is greater than the income and the material falls within the optimum pit shell, the material

is sent to the waste dump.

This technique allows the individual costs for each block located at various depths within the pit (and

therefore varying costs) to be determined. This avoids a common cut-off grade being applied to

every block within the pit but allows each block to be evaluated on its merits.

For mining operational purposes a marginal cut-off grade will be applied based on grade control

results and the edges of economic material will be determined by this. In order to perform this

function a cut-off grade value of 1.0% has been determined based on a copper price of$3.50 / lb.; a

silver price of $30 / oz and metallurgical recovery of 90%.

15.5. MINING INVENTORIES

The resource model for the Dikulushi deposit has been developed to define the deposit from surface

to a depth of approximately 350 m below surface. The current open pit strategy targets the top-

most portion of the identified Mineral Resources. The extraction of the deeper resources will

depend on a viable underground mining approach.

15.6. MINING RECOVERY AND DILUTION

Internal mining dilution has been built in to the mining block model through regularisation with the

intent being to mine the blocks above cut-off grade criteria in their entirety. There will be some

additional dilution at the edges of this economic mineralised zone, such as between the classified

ore and the waste material. This will be at waste copper and silver grades according to the resource

model. The average additional dilution of 15% (after ore loss) at 0.5% copper and 20g/t silver has

been allowed in the financial modelling to account for this peripheral dilution and is effected after

the mining recovery of 95% has been applied. The dilution is considered conservative at this stage.

15.7. RESERVE CLASSIFICATION

The classification of ore reserves according to the JORC code follows a process as described in Figure

15.8. Reserves declared from an Indicated Mineral Resource are only allowed to be classified as

Probable Reserves. Typically reserves declared from a Measured Mineral Resource are classified as

Proved reserves. However, a provision exists where Measured Resources can be declared as

Probable Reserves if there is insufficient certainty in the modifying factors to meet the requirements

of the proved category but still sufficient certainty to meet the probable category. In the case of the

Projects reserve categorisation, all Measured and Indicated Mineral Resources below the 850m RL

are classified as probable due to the following areas of uncertainty and thus risk:

Further work is required on the final designs regarding the support of the North pit wall in

the lower section where it interacts with the major, known fault, existing underground

stopes, and underground development. Initial work has indicated that it appears feasible,

based on good mining practices being maintained, but there remains an elevated risk related

to the actual physical properties of the rock and the location of faults zones.


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It is considered that mining down to the 850m RL bench can be effectively managed, and

possibly below to the 840m RL with no major ore loss. However, should the ground

conditions be significantly different to those used then the steeper pit walls may not be

practicable and would render mining of material below this point to be at a higher risk.

Stope blasting may have adversely affected the wall rock between 870m RL and 850m RL. It

is considered that mining can effectively take place down to the 850m RL since the stope is

confined to the eastern end and access to the bulk of the ore can still be maintained.

Since the inter-ramp wall angles have been steepened over initial parameters there will be a

need for continued monitoring of the pit slope stability. This is especially so in the areas of

faulting, interaction with underground excavations, wall support and catch fences.

It is considered that practical operating constraints may prevent final extraction as planned

due to the steepening and narrowing of the pit below 850m RL, especially when considered

in conjunction with the points above.

Figure 15.8 Mineral Resource and Mineral Reserve classification

There is a small portion of Inferred Mineral Resource material within the pit design, down to 840m

RL, which is likely to be mined and treated. This is not included in the mining inventory or the

Mineral Reserves and has no effect on the economic viability of the reserve.

Table 15.4 Dikulushi mined material

Mining Loss and Dilution applied

tonnes

Waste 20,102,963

Ore (diluted and recovered) 538,978

Strip Ratio 1:37

Total Material 20,641,941


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15.8. MINERAL RESERVES TABULATION

A financial model has been developed and analysis indicates that a positive return will be made.

There are several areas where it is considered that conservative estimates in mining costs have been

made. It has also been noted that the waste dilution of the ore is considered conservative for this

type of operation, mining style and deposit.

Resulting Mineral Reserves at the Project are only based upon Measured and Indicated Mineral

Resources. The Mineral Resources located below the 850m RL that fall within the optimised pit

shells, are classified into the Probable Mineral Reserve category due to the areas of uncertainty and

higher risk relating to maintaining good mining practice coupled with geotechnical risk relating to

faults and underground void interaction. It is recommended that the risk profile is reviewed at the

850m RL.

The resulting Mineral Reserves are supported by historical production and current processing data

and are tabulated in Table 15.5 using a 1.0% copper cut-off grade. All stated Mineral Resources are

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

mining losses and diluting materials brought about by the mining operation.

Table 15.5: Dikulushi Mineral Reserve statement as at August 2011 at a 1% copper cut-off grade.

Category Volume

(m3*1,000)

Density

(t/m3)

Tonnes

(*1,000) Copper (%)

Silver

(g/t)

Proven 65,971 2.8 184,719 7.27% 206.85

Probable 126,521 2.8 354,258 5.51% 169.46

Total Proven and Probable Reserves 192,492 2.8 538,977 6.12% 182.28

The above reserve does not include any inferred material.

It is noted that finance for the cut-back has already been made available and that the planned cut-

back is already in progress.


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16. MINING METHODS

Mining at the Project will be by conventional open-cut methods, and will be carried out by a

contractor under the supervision of CMCC staff. Tenders for the mining contract were invited from

three African based earthmoving companies who submitted detailed tenders for review by the

company. Following a comparison of the tenders and further clarification of their submissions, the

mining contract has been awarded to Mining Company Katanga (MCK). MCK’s primary earthmoving

fleet at Dikulushi will comprise 85 tonne to 190 tonne hydraulic backhoe excavators and a fleet of 40

tonne articulated dump trucks (ADTs) and 90 tonne off-highway trucks (OHTs). Production drilling

and blasting will be carried out by CMCC staff; blasting agents and accessories will be supplied by

African Explosives Limited, Lubumbashi (AEL) the DRC registered extension of the South African

headquartered international explosives supplier. Atlas Copco, who supplies the production drill rigs,

will be contracted to provide maintenance for the drill fleet. A list of primary and ancillary mining

equipment is given in the contractor’s fleet section.

Mining activity is carried out on two 12 hour shifts for the first 12 months and then reverts to a

single day shift as the larger fleet is demobilised. The equipment operator’s roster is 9 weeks on and

4 weeks off. Ore loading, assisted by spotters, will be highly selective and restricted to the day shift

to improve mining recovery of the high grade ore which is visual in nature. The Dikulushi sulphide

ore is grey in colour compared to the red hematite altered waste rocks and thus visual spotting

during ore mining will allow for improved selectivity.

The existing Dikulushi open cut was mined between 2002 and 2007 using open pit methods by the

previous owner Anvil. The open cut was mined to a final depth of 150m vertical and the operation

then changed to underground. The current Dikulushi pit is presented in Figure 16.1 below as of

2011.

Figure 16.1 The existing Dikulushi open pit in 2011

The mining study has looked at both UG and pit cutback options to extract the remaining high-grade

ore at the bottom of the current pit which served as the crown pillar for underground mining. The


P a g e | 79

open cut method was chosen as it posed a lower risk, less specialised workforce, shorter lead-time

for contractor mobilisation, and higher reclamation of existing resource. The decision to go with

contract mining was made based on the capital outlay required being too high to go with an owner

operated mining fleet and the supplier lead times being too long.

Mining equipment selection was done to allow mining of the waste in a suitable time frame to make

the cutback project viable and to allow small enough equipment to operate in a tight area at the

bottom of the pit. The cutback requires two sets of equipment to allow the pit to be mined to final

planned depth.

The cutback requires a total of 20.1Mt of waste to be mined over 18 months. However in order to

mine the cutback economically, 18Mt is required to be moved in 12 months. The bulk earthmoving

requirement over a short period combined with the spatial limitations of mining a cutback meant

larger excavators and haul trucks were required. The larger equipment was sized to meet these

requirements.

The final pit design narrows significantly at depth as the ore body itself has an average width of

about 10 m. The deeper part of the pit design dictates smaller equipment is required to mine the pit

past the 850mRL level. Therefore two sets of equipment were necessary to be able to mine the

Dikulushi cutback economically. The mining contractor selection process then required the

contractor to have suitable large scale equipment available in the DRC. Most mining contractors in

the DRC only have access to smaller capacity mining equipment.

Selection of the blasting and explosives mixing and storage equipment was completed and supplied

by AEL, based on the remnant storage blasting infrastructure at Dikulushi and the proposed

explosives consumption rate. AEL supplied explosives to Dikulushi under the previous owners. AEL

is the only reliable international explosives supplier in the region.

16.1. MINING STRATEGY

The current Mawson West strategy is to process the current low grade stockpile on surface and

thereafter, open pit mining of satellite resources until the transition into processing ore from the

active mining faces of the pit. The low grade stockpile is expected to be exhausted by December

2011.

The mining strategy for the Dikulushi open-pit is to use conventional drill and blast techniques using

an excavator and truck fleet to load and haul the mined material to the ROM pad or waste dump

site. A mining contractor is to be engaged to carry out this work.

The pit is planned to be mined as a series of staged cut-backs of all the walls to enable deepening of

the pit and mining of ore in the lower regions.

Portions of the deposit have been mined from underground. Some of these existing underground

workings, both existing underground stopes and development headings, will be intersected by the

deepened open pit. In all cases the geotechnical and physical interaction of the pit operation and the

pit walls with the existing underground voids has been considered.


P a g e | 80

Mining will be done both on a BCM rate and on an hourly hire rate. The footwall side of the cutback

has a number of underground workings which will create working areas requiring elevated safety

compliance. DOIR/MOSHAB Guidelines (2000) for open pit mining through underground workings is

the minimum Mawson West internal safety compliance standard. Rather than place the emphasis

on the local earth moving contractor to adhere to these standards. Mawson West mining staff will

strictly supervise the mining of these areas by utilising the mining fleet on an hourly hire basis.

The split between BCM and hourly hire is achieved by considering the pit in three stages. The white

lines in the Figure 16.2 represent the existing pit.

Stage 1, shown in Figure 16.2 as green horizontal lines. Stage 1 does not contain any known ore and

extends from the surface to 850mRL and will be mined on a BCM rate.

Stage 2, shown in Figure 16.2 as blue lines. Stage 2 generally lies on the footwall side of the pit and

has ore and all underground workings including footwall drives, ore drives, and access cross cuts,

decline and stopes and will be mined on an hourly rate. Stage 2 extends from 940mRL to 850mRL.

Stage 3, Stage 3 is shown in Figure 16.2 in yellow lines at the base of the pit. Stage 3 commences

from 850mRL where both Stage 1 and Stage 2 simultaneously end. Stage 3 is mined on an hourly

hire basis only.

The total quantity of material mined by hourly hire is 18% of the total volume moved.

Figure 16.2 The cutback stages

Stage’s 2 and 3 of the cutback will be mined by hourly hire of MCK equipment. The areas mined by

hourly hire were identified as not suitable for the big fleet due to either safety or grade control

considerations. These areas were identified by individually inspecting flitches at 5 m intervals

through the pit and assigned as being extracted on hourly hire using the small 85t digger.


P a g e | 81

For costing purposes the 880m RL bench was analysed with the Talpac simulator assuming that in

the area designated to be mined on an hourly rate an 85t digger can operate at normal productivity

over 50% of the material and at half its loading productivity over the rest of the 50%. The plant

times thus obtained were multiplied by the corresponding hourly rates that MCK had provided and

the unit costs of the load and haul were derived. The ratio between these hourly hire rates thus

obtained were compared with the BCM rate for same (880m RL) bench and was estimated to be very

close to 1.25. This ratio was assumed to be a fair representation of the cost increase from mining by

hourly hire for all areas designated to be mined on the hourly hire rate.

Figure 16.3 The cutback width at the 880Mrl. The white outline is the existing pit. Red lines show the old underground workings and the ore blocks are in blue.

The bench between 840m RL-850m RL is a transition zone from big to small mining equipment. The

840m RL is a solid 200m x 100m bench with a 5:1 strip ratio but still contains a considerable amount

of waste that can be mined by the larger EX1900 excavator. After the waste in this bench is taken

out the remainder of the pit will be mined exclusively by the 85t digger and 40t ADTs.

All waste will be hauled from the pit and placed on a waste dump or used in the construction of a

drainage diversion bund.

The perennial Dikulushi stream flows past the mine site. The original stream course crossed the

current open pit and was diverted via a channel dug around the pit. The extended pit identified in


P a g e | 82

this study intersects the stream diversion channel and this will now have to be moved further away

from the pit by excavating another channel and possibly the placement of some waste material as a

flood bund to avoid any possible flood waters from entering the pit. The previous diversion required

for the original open pit has worked well over the years and no failures have been recorded.

The number of working faces will be restricted by the annulus geometry of the pit cutback.

Essentially, for each active bench, there can be a maximum of two working faces working off a single

ramp.

For the majority of the operations the haul trucks will be required to perform a full 180° turn at the

operating face to enable them to be loaded and return back along the operating bench. The

minimum cutback width has been chosen to suit the equipment to be used. The selected contractor

will be using two basic haul trucks. Hitachi EH1700 Dump Trucks (rigid chassis) will be used in the

main upper section of the pit to move the bulk of the material and in particular the waste. Cat 740

articulated dump trucks will be used in the ore and lower sections where manoeuvrability is required

due to more confined working areas.

16.1.1. CONTRACTORS FLEET

A mining contractor based in the DRC has already been selected to operate the open pit under a

negotiated contract. The equipment provided by this contractor is as follows (Table 16.1):

Table 16.1: Major Equipment List – Dikulushi Open Pit Project

Item No Description Quantity

1 Hitachi Excavator EX1900 2

2 Hitachi 870 Excavator 1

3 Hitachi EH1700 Dump Truck 10

4 CAT D9R Dozer 3

5 CAT 14M Grader 2

6 CAT 740 Dump Truck 6

7 6x6 Volvo Service Truck 1

8 6x6 Volvo Service Truck 1

9 CAT773WT 1

10 BELL B40D Water Cart 1

11 Tyre Handler 1

12 Hiab Truck - 1

13 Lighting Towers 8

The selected mining contractor has reviewed the mine designs and has confirmed their ability to

safely mine the cut back.


P a g e | 83

16.2. OTHER MINING FLEET

The remaining mining equipment fleet of drills, charging vehicle and ancillary equipment will be fit

for purpose and depend on the requirements of both Mawson West and the mining contractor. The

selection of these items is not critical to the reserve estimation.

16.3. GEOTECHNICAL

Geotechnical analysis and recommendations were provided by Turner Mining and Geotechnical Pty

Ltd (Turner) for slope and design guidelines in designing the Dikulushi pit. Mike Turner is a Qualified

Person and has signed off against the geotechnical portion of the pre-feasibility study.

16.3.1. DATA

The data used for this stability analysis included all previously available data plus diamond drill core

logging from a drilling programme completed in late 2010 (Figure 16.4). The drillhole programme in

2010 targeted areas where there were gaps in geotechnical data.

Data was filtered to ensure only reliable, quality measurements were used and many of the old

orientated measurements from core for the hanging wall were not used due to poor orientation

quality and potential measurement errors. The holes drilled in 2010 were logged by Mawson West

geologists under the guidance of geotechnical engineers from AMC Consultants (Perth) and was of a

higher quality than older data. Alpha and beta angles were taken of joints per core run (“joint” in

this regard includes joints, bedding, open veins, fault related fractures and other open structures).


P a g e | 84

Figure 16.4 Location of geotechnically logged drillholes

All structural data used in the stability analyses was derived from open pit mapping and orientated core logging and only included data where the original source could be verified with surveyed coordinates. The data was analysed using the same techniques used in previous stability assessments at Dikulushi. The orientated structural data was analysed using the following software packages: Dips (Rocscience, 2010); and Swedge (Rocscience, 2010). “Dips” was used to evaluate multiple structural measurements with stereographic projections, and “Swedge” was used to evaluate wedge stability.

16.3.2. GEOTECHNICAL DOMAINS

Geotechnical domains were defined during previous studies based on rock type, rockmass strength,

pit orientation, and bedding dip and orientation. The domain boundaries were adjusted in the most

recent study (February 2011) with the additional data and modified slightly to suit changes in

bedding orientation and pit designs. The stability analyses were undertaken for each geotechnical

domains (Figure 16.5).


P a g e | 85

Figure 16.5 Pit slope design domains

16.3.3. SLOPE GUIDELINES

The domain data was analysed using Dips (Rocscience, 2010) to determine representative dip and

direction values for the major structural sets. This was used to provide an indication of the potential

for wedge and toppling failures. No areas were determined to be susceptible to toppling failure but

a number of wedges were observed.

The Swedge software (Rocscience, 2010) was used to further evaluate the potential for wedge

instability in the domains. A joint water content of 25% was used for all joints, together with

cohesion of 100kPa and friction of 30°.

The most recent slope design guidelines are summarised in Table 16.2. These guidelines assume

good quality blasting, drained slopes and no additional joint sets or faults, or major changes to

bedding dip and dip direction. Changes to any of these conditions will require an additional analysis

of data to check the continued suitability of designs.

The pit designs (110729base3008.dtm) complied with these pit design guidelines apart from 4m

benches instead of 6m on the 830mRL in the eastern half of the pit. Catch fences should be installed

on these benches to ensure rockfalls can still be controlled.

The additional information obtained from the 2010 geotechnical drill holes has permitted the bench

height to be increased to 20m in Domains A-East, D, E and F, whereas the bench height for the A-


P a g e | 86

West Domain has been maintained at 10m. Pre-splitting, good quality production blasting and

drained slopes are essential for slopes with these designs to remain stable.

Table 16.2 Slope design guidelines

Domain Location Inter-Ramp

Slope Angle

Bench Face

Angle

Bench

Width

Bench

Height

Weathered To 30m depth 40° 50° 5m 15m

A-West Northwest 42° 55° 4m 10m

A-East Northeast 50° 60° 5m 20m

B C and D West 50° 60° 5m 20m

E South and southeast 60° 75° 6m 20m

F East 60° 75° 6m 20m

WEATHERED DOMAIN

Previous design guidelines have used a depth of 30m for the depth of weathering and change in

bench face angles. The depth to fresh rock in the 2010 series of holes shows the significant variation

in the depth to fresh rock around the Dikulushi pit (Table 16.3). Previous personal experience of

Mike Turner, the geotechnical consultant, with Dikulushi has shown that the main indicators of

weathering below 30m, such as the change in rock colour and weathering along bedding planes only

have a relatively minor impact on rockmass strength. The impact of weaker bedding planes can

become significant however if blasting quality is poor, hence the planned use of pre-splitting for all

walls.

Table 16.3 Weathering depth from new holes

Hole Number Depth to Fresh Rock

1009DK003 57.8m

1009DK102 44.6m

1009DK103 78.9m

1009DK104 36m

1009DK106 >23m (hole stopped prematurely)

1009DK121 4.8m

1009DK122 54.4m

1009DK123 8.2m

1009DK124 18.5m

Indicative values of rock strength from the logged values of the recent series of drillholes indicate

the rock at depth only reaches the equivalent maximum of 25MPa for most of the holes. This is a

significant under-estimate compared to previous laboratory test results, ranging from 66 to 197 MPa

(Turner, May 2010), and will be due to failure along weak bedding planes. Rock strength and

bedding failure should be monitored as the cut-back progresses to enable optimisation of blast

designs.


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For the purposes of the pit design work the depth to the base of weathering has been maintained at

30m for slope design after reviewing the logging. The actual depth and resultant mining in the

weathered material will be guided by inspection of the rock mass as mining of the cut-back

progresses. As a result of this there should be sufficient area allowed around the pit to adjust the pit

edge accordingly.

A-WEST DOMAIN

The bedding angle in the A-West domain is 45°/099° (into the pit) and this will result in major large-

scale instability if the inter-ramp angle is cut steeper than bedding.

The slope design guidelines for this domain include the flattest slopes in the pit for this reason, with

a 55° bench face angle, 10m bench heights and 4m bench widths.

A-EAST DOMAIN

The additional data obtained for Domain A-East enabled a more detailed analysis for a potential cut-

back. The recommended design uses a 60° bench face angle, 20m bench height and 5m bench width,

for an inter-ramp angle of 50°.

The data and previous experience in this section of the pit indicates variable bedding and joint

orientations, with large-scale undulations. These features will result in bench-scale instabilities,

irrespective of bench face angle. Mesh might be required to stabilise bench faces above haulage

ramps if located along this wall. This is an acceptable, and well established, support technique used

in the stability control of open pit walls.

D DOMAIN

The Swedge analysis for Domain D was undertaken using a 60° bench face angle and 20m bench

heights and the analysis indicated no major wedges. Previous experience and the variability of the

rockmass and structural orientations indicates that this will lead to material falling off from bench

faces. The 5m bench widths have previously performed satisfactorily with regards to controlling this

type of fall material.

E DOMAIN

The Swedge analysis of Domain E showed that a combination of a 75° bench face angle, 20m face

height and 5m bench width would produce a factor of Safety of 1.58 assuming a joint-water content

of 25%.

The sensitivity analysis shows the critical impact of wall orientation. The water content of joints is

also critical, reducing the Factor of Safety from 1.59 for the general 25% water content to 1.2 for a

water content of 82%. This emphasises the need for good wall drainage of groundwater. The

previous failure in the south of the pit was associated with both poor mining and water inflows into

the joints (water flowing down the haulage ramp).

The results of the Swedge analysis can be seen in Figure 16.6 and the factor of safety sensitivity in

Figure 16.7.


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Figure 16.6 Domain E Wedge Potential


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Figure 16.7 Factor of Safety Sensitivity Analysis, Domain E, 75° Bench Face Angle

F DOMAIN

The Swedge analysis for Domain F was undertaken using a 75° bench face angle and 20m bench

height. The analysis indicated a potential for wedges only with a rare combination of joint sets. This

wall has previously been shown to be stable but is susceptible to blast damage, as with all other

walls.

16.3.4. POTENTIAL FAILURES

There have been several wall failures in the pit during the previous mining activities. It was

determined that a previous south wall failure at Dikulushi during the operations under Anvil was

partly caused by over mining of a lower bench and poor wall blasting aggravated by rainfall run-off

flowing into the rockmass via the haulroad. A north wall failure occurred on a previously unmapped

fault, also aggravated by run-off flowing down the haulroad.

It is essential that good mining practices and blasting techniques are enforced. The other factors

affecting stability described in the May 2010 report are still valid and need to be taken into account

during the design and operational stages.

The factors that could affect stability in the pit include:

blast-damage to walls

over-mining of benches

elevated groundwater levels, such as perched water levels behind structures or poor

drainage through clay-rich zones

unexpected structures and joint combinations

1.52

1.54

1.56

1.58

1.6

1.62

0 10 20 30 40 50 60 70 80 90 100

Fact

or

of

Safe

ty

Percent Change (%)

Percent Change (%) vs. Factor of Safety

Slope Dip (60°-90°) Slope Dip Direction (312°-382°)

Slope Height (10 to 30m) Water Percent Filled (0 to 50%)


P a g e | 90

exposure of wide shear zones adjacent to the orebody

severe rainfall surface run-off

long-term water inflow into rockmass

failure of benches and pit floor into existing underground excavations.

Some of these could impact on stability around the pit, and the major risks are discussed below.

NORTHERN WALL

The Northern wall suffered a major failure in 2006 following exposure of a persistent planar fault

behind the wall. Exposure of similar combinations of faults and bedding will require adjustments to

slope designs to prevent failures. The current designs (110729base3008.dtm) include leaving in-situ

a portion of ground between the Northwall fault and the northern wall of the pit. This block of

ground does not daylight and is therefore not technically a wedge. Failure of the block will still be

possible due to stepped-path failure and therefore an intensive cable bolt reinforcement

programme is planned to stabilise the area as the wall is cut-back.

NORTHEAST WALL

The northeast wall is susceptible to bench-scale wedge failures and adherence to slope design

guidelines is necessary to minimise such failures.

EAST WALL

The northern half of the east wall suffered a combination circular/toppling slumping failure in a fault

related zone during the wet season in the first quarter of 2005. The failed material was removed as

part of a cut-back and has shown no sign or similar movement. Drainage has improved since the

failure and the width of the poor ground zone has narrowed significantly further to the east. No

similar instabilities are expected if groundwater levels are controlled by pumping from underground.

SOUTHEAST WALL

The south/southeast wall suffered a large wedge failure in 2006 due to a combination of issues,

including poor run-off control down the haulage ramp, blocked weep holes and very poor blasting

techniques at the toe of the south wall. Improved groundwater drainage, run-off controls and good

blasting techniques will significantly reduce the risk of similar failures and the Factor of Safety is over

1.5 for the design with 75° bench face angles (Figure 16.6). A sensitivity analysis using Swedge for

the 75° bench face angle (Figure 16.7) shows the most critical item impacting on stability is the water

content of joints.

WEST WALL

The west wall was the site of very difficult mining conditions early in the life of the pit, due to the

mélange of carbonates, clay and other fault-breccia related material in combination with saturated

ground conditions. The deeper exposures on the west wall show less intense weathering, reduced

water content and significantly more competent rock. The proposed weep holes and run-off

controls will assist in maintaining stable walls.


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16.3.5. OTHER FACTORS AFFECTING STABILITY

MINING PRACTICES

Slope stability can be seriously affected by sub-standard mining practices and Mawson West at

Dikulushi is planning to dramatically improve mining practices relative to those previously employed.

Blasting, dewatering and surface run-off should be improved and cable bolt support and catch-

fences should be used in critical sections.

BLASTING

The stability of existing bench faces at Dikulushi was affected by the quality of wall perimeter

blasting. The walls were damaged by poor blasting and stability was compromised, leading to

numerous small batter-scale wedge failures. Pre-split blasting was only undertaken along some of

the bench faces, and the improvement in conditions was very noticeable.

Pre-split blasting and improved buffer blasting should be used to reduce blast damage to the walls.

Spare drilling capacity and additional supervision should be made available to ensure a high standard

of drill and blast practices are employed

Pre-splitting or well-designed buffer-blasting will prevent or limit failure to bedding planes, which

are weakly cemented. Failure of bedding planes leads to wedge failures and potentially could lead to

toppling failure in the south-east wall of the pit. Wall-control blasting should incorporate angled

holes parallel to the bench face angle where possible.

The slope design guidelines included in this report assume good quality wall control blasting.

DEWATERING

Instability due to groundwater is not expected for the planned cutback.

The rockmass close to the existing Dikulushi pit is planned to primarily be dewatered via drainage

into the existing voids. Deeper sections of the cutback should incorporate weep holes drilled into

the walls to ensure the rockmass within 10m of the wall is drained. This drainage method has

worked successfully in the past at Dikulushi.

The presence of an existing underground mine will have a significant dewatering impact, especially

on the northern wall where most of the underground development is located.

Conditions in the western wall of the pit have deteriorated in the past due to saturated conditions.

Existing dewatering boreholes to the west of the pit were installed to reduce groundwater levels in

the area and reduce inflows to the weak western rocks. These dewatering boreholes should be re-

commissioned ahead of mining if still serviceable. The planned surface run-off controls should limit

the inflow and recharge of groundwater via haul roads, which was previously a major issue.


P a g e | 92

SURFACE RUN-OFF

Poor surface run-off control was implicated in all previous major slope failures at Dikulushi and in a

number of batter-scale failures. High intensity tropical storm events cause significant flow down the

haulage ramps and along bench.

Previous run-off control down the haulage ramp was inadequate, with rainfall, dust suppression and

seepage water flowing along the road and into the rockmass.

In-pit surface run-off has been addressed by planned excavation of sumps at the lower end of

drainage domains (Figure 16.8). Water is planned to be piped from these sumps to the surface,

either directly with pumps or gravity fed in pipes to a central pump and from there to surface. Run-

off control improvements can be made where the haulage ramp exits the pit and down the haulage

ramp.

Figure 16.8 Run-off control domains

EXISTING UNDERGROUND EXCAVATIONS

The recent pit design has taken into account existing underground excavations and has been

modified to minimise the impact of such excavations on wall stability. 19 excavations will be


P a g e | 93

intersected on the northern wall and 5 on the southern wall. Ore drives and cross-cuts will be

intersected in the base of the pit on every underground level, spaced 20m apart. Some of these

underground excavations collapsed or suffered overbreak which should be taken into account when

the open pit approaches any excavation.

The mine should use the guidelines covering mitigation of this risk published in the Open Pit Mining

Through Underground Workings document issued by the DoIR (2000) in Western Australia.

Measures included marking off zones of potential impact with caution tape, probe drilling and

modified drill and blast patterns and techniques.

Probe holes should be drilled when approaching potential voids, holes should be fired to collapse

the rock around the voids and old voids should be filled where the pillars between the pit and the

voids are potentially unstable.

EARTHQUAKES

The seismic hazard data for the Dikulushi area has been assessed from the Global Seismic Hazard

Assessment Program data (GSHAP, 1999). The data indicated a 10% probability of exceeding

between 0.4 and 0.8 m/s2 peak ground acceleration over 50 years (based on a 475 year return

period). This falls in the low-hazard category and increased acceleration has not been considered in

the stability analyses.

16.3.6. MAPPING, MONITORING AND ADDITIONAL DATA

Additional measures and data are required to ensure a high degree of confidence in designs and

stability. These include ongoing geological mapping, monitoring prisms and additional diamond

drillholes.

Fresh exposures of rock should be mapped for rocktype, structures and cracks as the cut-back

progresses deeper and should be stored on hardcopy and digitally.

16.4. IN-PIT SUPPORT REQUIREMENTS

Due to the interaction of the pit with the Northwall fault, existing underground stopes and

underground development it was necessary to review the stability of the pit walls in more detail.

Underground stope and development blasting may have adversely affected the integrity of the

wallrock adjacent to the 850E stope between 870m RL and 850m RL, and rockbolt and mesh support

should be installed on walls either side of the stope void intersection. The lateral and vertical extent

of such support depends on the condition of the rockmass, which should be evaluated during the

cutback.

Underground development will have affected the integrity of the rock adjacent to excavations and

additional support may be required at various points around the pit to control loose rocks, especially

at the 19 holing points on the northern wall and 5 on the southern wall.

There will be a need for monitoring of the pit slope where there is potential for movement due to

the Northwall fault, and areas close to underground excavations. Regular documented visual


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examinations will suffice unless cracks and deformations are observed and then prisms should be

installed.

The reviewed pit designs (110729base3008.dtm) have followed the most recent geotechnical slope

design guidelines. The lowest berm (830m RL) is only 4 m instead of 6 m on the eastern end of the

north and south walls and catch fences are planned to control rockfalls on these berms. Sufficient

funds and equipment are available for the catch fences and the support and reinforcement work

mentioned below.

As the pit approaches the 870m RL the effects of the interaction with the underground stope will

need to be monitored carefully, especially the state of the wall rock. Geotechnical reviews should be

undertaken as the pit progresses below the 870m level. Modifications to the design and support

regime may be required as a result of these reviews.

It is considered that mining down to the 850 bench can be effectively managed, and possibly below

to the 840m RL with no major ore loss. However, should the ground conditions be significantly

different to those considered then the steeper pit walls may not be practicable and would render

mining of material below this point impractical and unsafe due to the narrowness of the pit below

840m RL.

16.4.1. EXISTING UNDERGROUND EXCAVATIONS

Cutback designs have taken into account existing underground excavations. Footwall drives and ore

drives are the main high-risk excavations as they run parallel to the pit slopes and designs have been

adjusted to avoid any impact from these excavations on slope stability. There are a number of other

drives that the pit will intersect and these have been highlighted in Figure 16.9 and Figure 16.10 for

the North and South walls respectively.

The mine will be following the guidelines covering mitigation of this risk issued by the DoIR (2000) in

Western Australia (Open Pit Mining Through Underground Workings).

Measures included marking off zones of potential void intersection with caution tape followed by

probe drilling. No-entry tapes and modified drill and blast patterns and techniques would follow

once the void location has been confirmed.

Deepening of the pit floor above old stopes and drives will be an area requiring extra care, especially

as the pit passes through the old 810 ore drive. Drilling of probe holes, and firing holes to collapse

the rock around the voids or for filling of old voids should be undertaken where required.

NORTH WALL

A total of 16 holing points have been highlighted in the North Wall from analysis of the existing

Surpac files obtained from Anvil Mining Figure 16.9.


P a g e | 95

Figure 16.9 Underground development holings in the North Wall of the pit design

SOUTH WALL

5 holing points were highlighted in the South wall Figure 16.10. There is a possibility of collapsed

and disturbed ground above the 810 drives and probe drilling will be important, possibly followed by

filling of any voids that are exposed.

Figure 16.10 Underground development holings in the South Wall of the pit design

850E STOPE

The eastern end of the 850 East stope cuts into the pit wall for an estimated distance of 10m from

870m RL to 850m RL. The stability of the open stope void and of the pit adjacent to the stope has

been assessed and no indications of major failure were indicated. Rockbolts (grouted bars or split

sets) and mesh should be installed on the north and south corners of the 870 and 850 berms either

side of the stope to prevent unravelling failures Figure 16.11 and Figure 16.12.

DRAWING REFERENCE

DATE

NTS July 2011

SCALE

NOT FOR CONSTRUCTIONMAWSON WEST LIMITEDDikulushi Open Pit -North Wall Underground Holing Points

Dikulushi Open Pit Review

DRAWING REFERENCE

DATE

NTS July 2011

SCALE

NOT FOR CONSTRUCTIONMAWSON WEST LIMITEDDikulushi Open Pit -South Wall Underground Holing Points

potential for collapsedground above drives



P a g e | 96

Figure 16.11 Stope Intersection Indicating Rockbolt support.

Figure 16.12 Area of pit wall requiring rockbolt support to prevent unravelling

870 TO 830 VENTILATION RISE

The top of the 870m RL to 850m RL Ventilation rise will be located on the middle of the 870m RL

berm. There will only be around 6m between the south side of the rise and the pit wall. This is a

potential area of instability and cable bolts should be installed on the pit wall opposite the rise

Figure 16.13.

DRAWING REFERENCE

DATE

NTS July 2011

SCALE

NOT FOR CONSTRUCTIONMAWSON WEST LIMITEDDikulushi Open Pit -Rockbolts required near old 850 Stope

10mPotential for wall damage Rockbolts required, 870 berm

Rockbolts required, 850 berm

10m


DRAWING REFERENCE

DATE

NTS July 2011

SCALE

NOT FOR CONSTRUCTIONMAWSON WEST LIMITEDDikulushi Open Pit - East end stability issues due to old stope

4m10m

Potential for wall damage

Rockbolts required on 850 Level

Rockbolts required on 850 Level

10m

10m



P a g e | 97

Figure 16.13 Plan view indicating development associated with the 870 to 830 Ventilation Rise

16.4.2. MAJOR STRUCTURES

The only major structure that required modified designs and support is the Northwall fault which

previously caused a large failure on the north wall. The current design does not open a free lower

failure surface into the pit but previous experience has shown that a stepped-path failure would still

be possible in the ground between the fault and the pit wall.

The pit designs have minimised the ground between the wall and the fault, and the remaining

ground will be reinforced with cable bolts, straps and mesh as illustrated in Figure 16.14 and Figure

16.15

DRAWING REFERENCE

DATE

NTS July 2011

SCALE

MAWSON WEST LIMITEDDikulushi Open Pit - 870 Ventilation rise cable bolts

870

Cable bolts below 870mRL(from 860) to support rockbetween ventilation rise and pit

10m 6m



P a g e | 98

Figure 16.14 North wall cable bolts and catch fence

Figure 16.15 North wall bolting patterns.

16.4.3. CABLE BOLTS AND CATCH FENCES

The proposed pit should utilise cables bolts, other support and catch fences to improve stability

compared to the old Dikulushi pit which used none of these methods.

DRAWING REFERENCE

DATE

NTS July 2011

SCALE

MAWSON WEST LIMITEDDikulushi Open Pit - North Wall Cable bolts and catch fence

850

870

890

810

830

2m catch fence required on 830 berm (60m)

Cable bolts for ground betweennorthern fault and north wall

Cable bolts below 870mRLfor ventilation rise

830

840

850

860

870

880

890


DRAWING REFERENCE

DATE

NTS July 2011

SCALE

MAWSON WEST LIMITEDDikulushi Open Pit - Northwall Fault cable bolts plans

850 and 860

870 and 880

890

Cable bolts for ground betweennorthern fault and north wall

10m20m

25m

830 and 840



P a g e | 99

Cable bolts, mesh and straps should be installed to ensure stability of the material to the south of the Northwall Fault (Figure 16.14 and Figure 16.15) adjacent to the ventilation rise between 870 and 860 berms (Figure 16.15).

Rockbolts should be used either side of the 850 East stope holing point.

Catch fences should be used on 830 berms on the eastern ends of both the north and south walls where the berm width is only 4m (Figure 16.16).

Figure 16.16 South wall catch fence on 830m RL

16.5. ROM PAD DESIGN

The existing ROM (run of mine) pad adjacent to the processing plant will be used for the remainder

of the mine life and no adjustments to this area are proposed or have been made.

DRAWING REFERENCE

DATE

NTS July 2011

SCALE

MAWSON WEST LIMITEDDikulushi Open Pit - South Wall catch-fence requirements

850

8102m catch fence required on 830 berm(95m)

830



P a g e | 100

16.6. WASTE DUMP DESIGN

Figure 16.17 Location of waste dump relative to expanded pit

The waste dump design parameters used for the Cutback dumps are:

Face slope 20º

Bench height 10m

Berm width 5m

Overall slope 15º

The waste dump capacities have been based on a swell factor of 30%. The top of the waste dump is

at 1,020mRL. The pit cutback will generate 20Mt of waste material with the height of the dumps

limited to 20m. The dump height is limited to 20m in order to keep the waste stripping costs to a

minimum. The dump footprint covers 82.5Ha with 71.5Ha in the south dump and 11Ha in the north

dump as per Figure 16.17.

The waste dumps were designed in Surpac using the dump design module. The waste dump

positions have been determined by taking into account geologically prospective ground (where

sterilisation drilling is still to be carried out), the existing drainage patterns, waste haulage profiles

and the space and infrastructure issues required for the planned operations. A grid of dump blocks

was overlayed on the dumping areas. Mining waste was scheduled into these blocks progressively

away from the pit until the total volume of waste is accounted for in the waste blocks. Additional


P a g e | 101

waste capacity is available at the Dikulushi site by either extending the length of the dumps or by

raising the height.

Some existing road re-alignments are required to accommodate the dumps. The waste mined

during the pre-production period will be utilised to provide road base material and other

infrastructure items, such as haul roads, as required.

16.7. SURFACE WATER MANAGEMENT

The surface water hydrology at Dikulushi is well understood. The mine has been operating for 9

years and current surface water management practices have been successful in controlling runoff in

the mine and plant area during this time.

In order to accommodate the pit expansion the eastern surface water management diversion drain

requires relocating. This diversion channel to the south of the existing pit was previously installed as

a result of the initial mining operations.

The new drain has been designed to accommodate as a minimum requirement a one in one hundred

year rainfall event based on historical rainfall records for the area. The gentle undulating

topography and rainfall records indicate that relocation of the surface water management drain

does not pose a significant flood risk to the expanded pit design.

The ESIA report previously submitted and approved for the Dikulushi mine includes an

Environmental Management Plan (EMP). The EMP includes water monitoring of the local waterways

and annual environmental reports that are submitted to the DRC govt. Water sampling in the

Dikulushi Mine area clearly indicates there is not an acid rock drainage problem associated with the

Dikulushi minesite. The water sampling records for the Dikulushi minesite have to be submitted

annually to the DRC government for review and have consistently been compliant with the

requirements of the DRC regulations. The waste and ore material that will be mined in the proposed

pit cutback at Dikulushi is the same as previously mined and therefore no ARD drainage problems

are envisaged based on the current dumps being in place for over 9 years with no ARD issues to

date.

A study of the mine water flows and mine water requirements was previously completed by SRK

consulting of South Africa in 2007 (report “Water Balance for Dikulushi July 2007”). The current and

future requirements for process water were reviewed by SRK consulting to ensure future mine water

requirements will be met and SRK issued an updated report on the Dikulushi mine water

requirements in July 2011 (“Water balance for Dikulushi Mine 2011 update”). The current water

sources are sufficient for the current operation although additional water will have to be drawn from

Lake Newton or boreholes in the driest month. The report contained the following

recommendations;

Recording of mine water flows from the installed meters and on site staff training to ensure

this is done correctly.


P a g e | 102

Any water discharged into Lake Newton is tested before discharge to ensure environmental

guidelines are met.

Installation of an A-pan evaporator to record evaporation rates at the mine site.

Figure 16.18 Surface water management – general arrangement


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17. RECOVERY METHODS

17.1.1. PLANT FLOWSHEET

The Plant and associated infrastructure had been dormant since December, 2008 and was

refurbished prior to start up in June, 2010.

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 (Figure 17.1).

Figure 17.1 Dikulushi Plant flow diagram

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


P a g e | 104

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%.

Each bag is weighed ready for despatch by truck to the Kilwa port.

17.1.2. TAILINGS STORAGE FACILITIES (TSF)

The first TSF for HMS tailings 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 (previous owners and operators of Dikulushi Mine). Mawson West has recovered and

processed approximately 15,000 t of coarse sand fraction tailings and fine material from this TSF.

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 side-hill impoundment and provides the area required

to limit the rise rate of tailings to acceptable norms.

Supernatant water from the tailings slurry is reclaimed for use in the processing plant via a gravity

decant comprising outfall pipe with three stacked ring penstock inlets.

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

from the processing of the low grade ore stockpile. The study concluded that a 2m embankment

raise would be required during 2011.


P a g e | 105

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 tailings. 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 expected to commence during

July, 2011. During this period, TD2 will be temporarily recommissioned for a period of

approximately 3 months, after which deposition will resume on TD3.

This will provide a tailings storage facility capable of supporting the open pit cut back mining

operation.

17.1.3. 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.

Table 17.1 Dikulushi Processing Summary relevant to ore to be mined in the pit cutback

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.

MWL PROCESSING

Mawson West has been blending material from surface stockpiles and the HMS Tails through the

plant to maximise copper output since June 2010. The recoveries from this activity are 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 will be associated more with those from the previous open pit


P a g e | 106

and underground mining operations carried out by Anvil. Processing statistics for the LG material

completed by MWL are shown in Table 17.2.

Table 17.2 Processing statistics for the LG material completed by MWL

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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18. PROJECT INFRASTRUCTURE

The Dikulushi operation was previously an operating mine and the infrastructure remains in place. It

has been used and maintained by MWL since they took over the project site. The infrastructure is

considered adequate for the resumption of open pit activities at the designed mining rates.

18.1. SURFACE FACILITIES

The existing surface facilities (Figure 18.1) remaining from the underground mining operation will be

suitable for use by the open pit mining personnel.

Figure 18.1 On-site office 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: 1x 2.0Megawat

Caterpillar, 1x 1.6 Megawatt Caterpillar, 4x 0.8Megawatt Mirrlees for a total capacity of

6.8Megawatts. The current power demand is in the order of 1.8MW and only the 2.0MW Cat,

1.6MW Cat and 2x Mirrlees are being utilised on rotation. This is sufficient to supply the extra

demand of 0.6-1.0 Megawatts for dewatering purposes during the cut-back project.









18.3. PROCESS WATER SUPPLY

Lake Newton (Figure 18.2) on the perennial Dikulushi stream provides storage for dewatering and

serves as a reservoir for the supply of process water.

Figure 18.2 Lake Newton


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Meteorological data has been collected between August 2005 and July 2011, 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 since the 2007 report due to changes in the mine ownership and as a

result the water balance was reviewed in July 2011 by SRK.

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. The recent update carried out a full water balance for June and July which are amongst

the driest months in the year. The study arrived at the following recommendations:

The water balance should be updated monthly so the variation in water use can be

measured and incorporated into the water balance. Measuring of meters has recently

commenced and the seasonal variation is unknown at this stage;

The water that is discharged to the Lake Newton will need to be monitored and only

released to the environment when the quality meets the discharge requirements

Record maximum and minimum temperatures, wet and dry bulb temperatures, rainfall,

wind speed, pressure and humidity from the weather station on site;

Install an A- Pan evaporator to record the evaporation at the mine,

A training session is recommended with mine personnel to train and handover the water

balance model.

A monthly water balance was prepared for Dikulushi Mine and is simulating an average monthly

water balance for June and July 2011. The model has been set up so that it can be updated on a

monthly basis. However, the meter data has only been collected for the past two months and

therefore little can be deduced from the model at this stage. The hydrology has however been

incorporated to try and simulate the impact of weather changes. An average water balance is

presented in Figure 18.3


P a g e | 110

Figure 18.3 Average water balance

This was also calculated for both a dry month and a wet month. From the water balance review the

following is 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.


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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.

CMCC currently has a contract to sell all the copper concentrate produced from the low grade

stockpile to smelters in India and China using LN Metals as its agent. CMCC has not yet committed

any of the concentrate that will be produced from the Project cut-back operation.

19.2. CONTRACTS

CMCC currently has a contract to sell all the copper concentrate produced from the low grade

stockpile to smelters in India and China using LN Metals as its agent and this agreement can be

extended at CMCC’s discretion. CMCC has not yet committed any of the concentrate that will be

produced from the Project cut-back operation.

There are various contracts either already in place or required to be entered into for the following

major areas:

Mining

Diesel supply

Transport

Reagents

Spares

A mining services agreement has been entered into with Mining Company Katanga SPRL, which will

be ratified in the coming weeks.









The current revenue estimates include the concentrate being transported to a smelter in China

where it will be converted to metal and sold to market.

Copper concentrate will be sold either directly to smelters, or via an agent or directly to metal

trading companies.


P a g e | 112

The current cost for transport to the smelter is $335.00/wet metric tonne (wmt) of concentrate with

a treatment charge of $56.00/wmt of concentrate and refining charges of $0.06/lb copper and

$0.40/oz silver.

The resultant net smelter return (NSR) for copper is 96.75% and the NSR for silver is 91%. The

estimated moisture content is 10%.

The study has used a copper price of $7,716/tonne copper ($3.50/lb. copper) and a silver price of

$0.96/g silver ($30.00/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 marketing is

estimated to be $1,201 per tonne of copper metal sold.

Commodity price projections have not been evaluated due to the short life of the cut back and

processing operations being less than 24 months.


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20. ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL

OR COMMUNITY IMPACT

CMCC submitted an Environmental Management Plan (EMP) which was accepted and adopted as

one of the permitting documents. MWL is required to provide annual environmental reports and

demonstrate that it is in compliance with the EMP. Mine remediation is one of the compliance items

in the EMP.

CMCC has lodged an environmental bond of $368,410. 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 is yet to be lodged. This

review did not find anything out of requirements.


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21. CAPITAL AND OPERATING COSTS

21.1. CAPITAL COST ESTIMATE

Table 21.1 Capital cost estimates

The majority of the above capital is spent over the first 6 months of the project (Table 21.1). An

additional $2,000,000 on top of the above itemised capital has been allowed for sustaining capital

and is allocated to be spent over the last 12 months of the operation. The total capital cost of the

project is thus $8,910,000.

21.2. OPERATING COST ESTIMATE

The Dikulushi operation is planned as a contractor mining operation with a number of permanent

Mawson West staff based at the site. The permanent staff will cover the following functions:

Site management

Contractor management

Mine management

Drilling (AC supplied) and blasting activities (AEL supplied)

ITEM Allocation USD

Consultants Fees

Mine planning & control 50,000

Geotechnical & water 60,000

D&B 30,000

Environmental 10,000

Preparing contracts 10,000

Capital Purchases

Office equipment Furniture 20,000

Technical peripherals Plotter / scanner / photocopier/printre 30,000

Computers & UPS ets 50,000

Dewatering system Both OP & UG, 12 x 80kW mounted pumps, 2 submersibles1,000,000

Radio communications upgrade 20,000

Survey gear upgrade 90,000

Software systems upgrade

Geology/planning/survey 180,000

Ditchwitch & accessories For grade control and in pit drainage works 80,000

Drill fleet 4 x Atlas Copco L6-30 rigs 2,200,000

LV Fleet 10 LV's + 1 Service vehicle for drill fleet 575,000

Ancillary equipment Cherry picker and accessories, to severe steelworks from UG reinforcements/ Rockbolting work100,000

Exiiting Equipment upgrades Axera 6 Jumbos, LM75 UG drill and LHD. 435,000

Electricity Infrastructure Post sub-station cabling for dewatering systems and LM 25 diamond rig for depressurization and probe drilling70,000

Site readiness Accommodation and East side drainage culvert 550,000

Recruitment Consultants / advertising / travels 50,000

Sterilization Sterilisation of South dump foot print 120,000

Mobilisation MCK Fleet mobilisation 1,100,000

Training Surpac Training 80,000

Total Project Capital Budget 6,910,000


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Dewatering

Ore Processing

Resource definition

Geological and grade control

Mine planning

Environmental monitoring and management

Community relations

Commercial, procurement and logistics

The mining contractor will be responsible for the following functions:

Clearing and topsoil removal

Truck loading

Haulage from the pit face to a waste dump or ROM pad adjacent to the pit or processing

plant

Roadway maintenance

Maintenance of all mining equipment

Supply of necessary ancillary mining equipment such as graders, dozers and lighting plant

21.2.1. MINING OPERATING COST

The mining operating costs have been based on a schedule of rates tendered by the mining

contractor who will carry out the mining activity. Adjustments have been made to these costs in

areas where additional difficulty in mining is expected and an hourly rate has been calculated. These

adjustments increase the mining costs in these areas and are considered to be very conservative.

DRILL AND BLAST COSTS

Drill and blast costs have been separately estimated based on a schedule of rates tendered by a

drilling contractor and explosive costs by the supplier for this project.

Table 21.2 Drill and blast unit costs

Drill and Blast Costs US$/t rock

Weathered 0.00

Transitional / Fresh 0.96

LOAD AND HAUL COSTS

Costs have been tendered by the mining contractor engaged to carry out the mining of this cut-back.

The overall load and haul costs have been based on the haulage profiles for the various benches in

the pit design. These profiles cover the haulage of the ore or waste from the respective mining

bench, to surface by the pit ramp system, thence to either the designated waste dump or the ROM

pad for feeding to the mill.

Adjustments have been made to the loading and hauling of material from below the 830m RL. This

adjustment is based on the perceived difficulty in mining operations below this horizon where the

pit steepens and narrows for the final ore removal. Whilst the incremental factors are arbitrarily


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arrived at it is considered that they err on the conservative side, and as such can be considered

appropriate for this study.

Table 21.3 tabulates the Load and Haulage costs as used. The specific gravity used to convert bank

cubic metres (BCM) to tonnes in the below calculation is a nominal value of 2.6. In the cost model,

the costs have been generated based on tonnes as derived from the density field in the resource

model which changes according to the specific material type. Tonnage, rather than volume will be

the controlling haulage factor. The base diesel cost used for these estimates is $1.25 / l . The costs

of ancillary equipment (dozers, graders, water carts and light vehicles) are included in the load and

haul cost estimates.

Table 21.3 Load and Haul unit costs

Haulage Depth

(mRL)

L & H

$ / t

% Hourly

Rate

1010 1.65 0%

1000 1.73

0%

990 1.57

0%

980 1.61 0%

970 1.71 0%

960 1.74 0%

950 1.99 0%

940 1.91 0%

930 2.05 13%

920 2.14 28%

910 2.27 27%

900 2.45 34%

890 2.63 38%

880 2.78 57%

870 2.95 58%

860 2.95 46%

850 3.25 61%

840 3.55 100%

830 3.59 100%

820 5.44 100%

810 7.35 100%

800 11.17 100%

OTHER MINING COSTS

Mobilisation and demobilisation cost estimates have been based on the schedule as tendered by the

mining contractor. These are $1,100,000 for mobilisation, which has been accounted for in the

capital expenditure table. The demobilisation cost of the contractor’s fleet and equipment,

estimated at $500,000, has not been allowed for as it is believed that this cost will be “considered”

as a mobilisation cost of the mining fleet to the Kapulo Project.


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The contractors (MCK, AC; AEL) fixed monthly fee amount ($408,311) has been converted into a unit

cost of $0.36 / tonne ($0.93 / BCM). MWL’s monthly Management and labour costs are estimated at

$370,208 ($0.84 / BCM). Miscellaneous earthworks have been costed at $0.19 / BCM. Additional

fuel excess costs have been estimated at $0.25 / BCM; Rehabilitation costs at $0.03 / BCM and

dewatering, geotechnical and grade control costs at $0.33 / BCM.

21.3. PROCESSING OPERATING COSTS

The Dikulushi treatment plant has been operated by CMCC since June 2010 treating low grade ore

and has average processing costs of $37.00 per tonne of ore. The TCRC’s of the current low grade

ore are approximately USD$1200.00 per tonne of copper metal after treatment and refining charges

and silver credits have been applied. It is anticipated that similar processing costs per tonne of ore

will be achievable when processing the proposed cutback ore. Due to the changes in ore grades and

recoveries for both copper and silver, it is estimated that the treatment cost will be approximately

$992.35 per tonne of copper metal after treatment and refining charges and silver credits have been

applied. The figure assumes a concentrate grade of 50% copper and 10% moisture.

21.3.1. OVERHEAD OPERATING COST

Site overheads and administration costs were based on an estimated a total amount of $4.2 Million

as supplied by MWL that includes permanent staff, travel, site maintenance, environmental

monitoring, transport, logistics, local taxes, levies and community costs.

Mining Contactor overheads are covered in the fixed monthly rate of $408,311 based on a schedule

of rates as supplied by the mining contractor (MCK); Drilling contractor (AC) and explosive supplier

(AEL) and accounted for in the mining costs.

21.3.2. CAPITAL EXPENDITURE

The cost of refurbishment of the plant has been repaid for by the processing of low grade ore so

there are no further capital costs for this project other than a small quantity of sustaining capital,

which has been allowed for in the sustaining capital estimate of $2.0 Million.

21.3.3. OPERATING COSTS

The process and administration operating costs have been estimated at $37.00/t of ore processed,

based on current operating parameters.

Table 21.4 Operating costs

Operating Cost Unit Cost

Management @ Admin $ / t $6.00

Labour $ / t $7.00

Operating Consumables $ / t $6.00

Maintenance Consumables $ / t $2.00

Power $ / t $16.00

TOTAL $ / t Processed $37.00


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22. ECONOMIC ANALYSIS

22.1. MINING SUMMARY

Table 22.1 below provides a summary of the mining and cashflow performance of the Project mine.

The total material moved is 20.64 Mt of material of which 538,969 t is ore at a diluted and recovered

grade of 6.12% copper. The average life of mine stripping ratio is 38.0 on a waste tonne for ore

tonne basis.

The average mining cost $3.95 per total tonne and $151.27 per ore tonne. The processing cost is $37

per ore tonne processed.

Capital costs are relatively low as the installations are already in place. A total capital cost $8.91M

has been allowed for of which $6.91M is start-up capital and $2.0M of sustaining capital is

estimated.

The mine life is 18 monthswith the majority of the ore mined in the last year. Initial ore production

occurs after 9 months of mining. Processing and copper production and sales occurs after month 10.

Power and dewatering costs have been included in the water management cost category as a unit

cost. Ancillary equipment has been included into the contract unit rates used in the operating cost

estimate.

The processing recovery for copper used for this estimate was 90% for the transitional and fresh

material.

The total cash cost of the operation is $92.7M. A total of 28.7 Kt of copper is sold along with

2.6 Moz of silver to produce revenue of $299.47M. This relates to an NPV12 of $115.78M at a

discount rate of 10%.

Table 22.1 Dikulushi Mining and Financial Summary

Year 1 Year 2 Total

Physical Schedule Total material mined K tonnes 17,485 3,157 20,642

Waste Mined K tonnes 17,437 2,666 20,103

Ore Mined tonnes 48,087 490,883 538,970

Copper mined grade copper% 3.71 6.35 6.12

Silver mined grade silver g/t 88.76 191.44 182.28

Copper mined t 1,783 31,188 32,971

Silver mined oz 137,222 3,021,321 3,158,544

Stripping Ratio 363:1 5.4:1 38:1

Costs Mining


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Year 1 Year 2 Total

Drill & Blast $ 13,849,387 3,373,844 17,223,231

Grade Control $ 161,079 280,235 441,314

Load & Haul $ 33,628,523 10,319,937 43,948,460

Rehab & Geotech $ 903,456 1,302,776 2,206,232

Contractor Overheads $ 6,252,878 1,116,811 7,369,689

General & Administration $ 5,647,761 1,008,732 6,656,493

Dewatering $ 105,998 82,567 188,565

Dayworks $ 1,310,574 226,383 1,536,957

Fuel Excess $ 1,541,961 419,575 1,961,536

Total Mining Opex $ 63,401,617 18,130,860 81,532,477

Total Mining Opex $/ore tonne mined 1,318.48 36.94 151.27

$/total tonne mined 3.63 5.74 3.95

Processing

Processing $ 2,202,298 16,561,203 18,763,501

$/ore tonne milled 45.80 33.69 37

Management & Admin

Administration $ 955,185 3,282,399 4,237,584

$/ore tonne milled 19.86 6.68 7.85

Total Operating Costs $ 66,559,100 37,974,462 104,533,562

Total Capital Costs $ 6,590,000 2,320,000 8,910,000

Revenue

Metal in conc copper t 1,605 28,102 29,707

silver oz 123,500 2,722,608 2,846,108

Metal sold copper t 1,553 27,188 28,741

silver oz 112,385 2,477,573 2,589,958

Sales & Transport Costs $ 2,465,036 32,048,551 34,513,587

Duties and Taxes 954,705 7,489,121 8,443,826

Copper NSR $ 11,981,283 209,788,550 221,769,833

$/t mined 249.16 427,37 411.47

Silver NSR $ 3,371,542 74,327,201 77,698,743

$/oz mined 70.11 151.42 144.16

Total Revenue


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Year 1 Year 2 Total

Revenue from Sales $ 15,352,825 284,115,751 299,468,576

Royalties $ 0 0 0

Net Revenue $ 15,352,825 284,115,751 299,468,576

Cashflow from Operations $ -61,216,014 204,283,616 143,067,602

NPV 10% $115,771,838 IRR 128%

22.1.1. SENSITIVITY ANALYSIS

MWL has carried out a sensitivity analysis on the cash flow forecasts and this is provided in Tables

22.2 to 7. It can be seen that the project is profitable at even modest copper and silver prices.

Tables 22.2 to 7 Sensitivity analysis on the cash flow forecast for the open pit cutback and treatment at Dikulushi

Table 22-2

Dikulushi Copper Project

Project Sensitivity to a Change in copper Price

Discount Rate

NPV (US$ million)

Change in copper Price

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

8% 83 102 121 140 159

10% 79 97 116 134 153

12% 75 93 111 129 147

Table 22-3


Project Sensitivity to a Change in silver Price

Discount Rate

NPV (US$ million)

Change in silver Price

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

8% 107 114 121 127 134

10% 103 109 116 122 127

12% 98 105 111 117 123

Table 22-4


Project Sensitivity to a Change in Operating Costs

Discount Rate

NPV (US$ million)

Change in Operating Costs


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

8% 124 122 121 119 117

10% 119 117 116 114 112

12% 114 113 111 109 107

Table 22-5


Project Sensitivity to a Change in Capital Costs

Discount Rate

NPV (US$ million)

Change in Capital Costs

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

8% 123 122 121 120 119

10% 118 117 116 115 114

12% 113 112 111 110 109

Table 22-6


Project Sensitivity to a Change in Fuel on Mining Costs

Discount Rate

NPV (US$ million)

Change in Fuel on Mining Costs

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

8% 123.8 122.3 121 119.3 117.7

10% 118.8 117.3 116 114.3 112.7

12% 113.9 112.4 111 109.5 108

Table 22-7


Project Sensitivity to a Change in Metal Transport Costs

Discount Rate

NPV (US$ million)

Change in Metal Transport Costs

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

8% 124.6 122.7 121 118.9 117

10% 119.5 117.6 116 113.9 112

12% 114.5 112.7 111 109.2 107.4

22.2. PAYBACK

As discussed the refurbishment cost of the mill has already been covered by the revenues from the

LG treatment, thus there is no formal capital payback period. The development of the cutback is to

be fully funded out of MWL’s current existing cash reserves. The total negative cashflow (including

capital costs) is -$61.22M (end of month 12) and which is back in positive territory by the end of

month 16.


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22.3. MINE LIFE

As stated, the mine life is 18 months for the Open pit cutback and treatment operations. MWL fully

expects to extend the mine life once other satellite projects within 50km of Dikulushi have been fully

evaluated.

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 nine

and half 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.

The mining convention also provides 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 Project.


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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 in April 2010 and work started immediately on

refurbishment of the plant, which was completed in June 2010. Since June 2010, MWL has been

producing copper-silver concentrate from a feed of blended HMS tails and reclamation of the LG

stockpile.


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25. INTERPRETATION AND CONCLUSIONS

The Dikulushi Project is a producing and developing property. Current processing of the Dikulushi LG

stockpile reserves has provided MWL with a robust cash flow and results demonstrate reliable

grades of remaining stocks when compared with RC drilling results and estimates.

The Dikulushi deposit has a long 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 pit cut-back volume. The pre-feasibility study and resulting

Mineral Reserves from the open pit cut-back further strengthens MWL’s production life from the

Dikulushi Project. MWL has approved, and already has, the funding for the open pit cut-back.

Successful mining will result from good mining practice providing clean pre-split of the final walls,

maintenance of the mining design, safe handling of the interaction with any underground

development and good dewatering of the pit walls. MWL intends to continue processing the LG

stockpile during the build up phase to production from the pit cut-back.

MWL’s strategy is to continue to develop satellite deposits to Dikulushi, such as Kazumbula, in

addition to the remaining Dikulushi Mineral Resources located below the planned pit cut-back. The

recent exploration drilling at Kazumbula has provided good quality 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 50km

of Dikulushi are currently being drilled by MWL.


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26. RECOMMENDATIONS

It is recommended that MWL continues with the planned Project cut-back. The key aspects for the

success of the cut-back are those of maintaining good mining practices, including very good pre-split,

ensuring no undercutting, keeping water away from the pit and maintaining the pit draining systems.

Annual reviews will be required for the environmental approvals and to ensure the integrity of the

tailings dam is 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.


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SRK Consulting: Project No: 436159 Water Balance for Dikulushi Mine – 2011 Update


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28. CERTIFICATES

OPTIRO PTY LTD

CERTIFICATE OF QUALIFIED PERSON

As the lead author of the report entitled “Technical Report on the Dikulushi Open Pit Project,

Democratic Republic of Congo” (the Study) dated 16 September 2011, on the Project cut back, of

Mawson West Ltd , I hereby state:

1. My name is David Gray 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 member of the AusIMM (303226) and registered member of

The South African Council for Natural Scientific Professions (PrSciNat, 400018/04).

3. I am a graduate of Rhodes University in South Africa with a BSc (Hons) in Geology in 1988

4. I have practiced my profession continuously since 1990.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of

Disclosure for Mineral Projects) (the “Instrument”).

6. I visited the Dikulushi Project property and surrounding areas for two days in November 2010. I

have performed consulting services and reviewed files and data associated with Dikulushi

between May 2009 and the present.

7. I am responsible for all the Sections of the Study and have contributed to Sections 17.1 and 17.3

and the associated text in the summary, conclusions and recommendations.

8. As of the date of this certificate, 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.

9. I am independent of Mawson West Ltd pursuant to section 1.4 of the Instrument.

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 22 September 2011.

David Gray

Principal Consultant (Optiro Pty Ltd)

BSc (Hons) (Geology), MAusIMM, PrSciNat


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OPTIRO PTY LTD

CERTIFICATE OF QUALIFIED PERSON

As a Qualified Person and author of the report entitled “Technical Report on the Dikulushi Open Pit

Project, Democratic Republic of Congo” (the Study) dated 16 September 2011, on the Project cut

back, of Mawson West Ltd , 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 a “qualified person” as that term is 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 Dikulushi from

August 2011 to the present.

7. Based on the information provided by Mawson West Ltd and reviewed by myself I contributed

to Sections 15, 16, 19, 20, 21, 22, 24, 25, and 26.

8. As of the date of this certificate, 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.

9. 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.

10. 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 22 September 2011.

Andrew Law

Director - Mining (Optiro Pty Ltd)

FAusIMM


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TURNER MINING AND GEOTECHNICAL PTY LTD


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SEDGMAN LTD

mawson west limited

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