Kapulo Copper Project, DRC National Instrument 43-101 Technical Report
Prepared by Coffey Mining Pty Ltd on behalf of: Mawson West Ltd
Effective Date: 30 June 2011
Qualified Persons: Harry Warries - MSc Mine Engineering, MAusIMM Steve Le Brun - BSc(Hons), MSc, MAusIMM, MMICA Chris Johns - MSc, P.Eng.(Alberta), MIEAust Peter Hayward - Dip Metallurgy, MAusIMM Aaron Massey - Dip Metallurgy, MAusIMM Chris Orr - MSc, MAusIMM
MINEWPER00482AC Coffey Mining Pty Ltd DOCUMENT INFORMATION
Author(s): Harry Warries Manager Mining - West Perth MSc Mine Engineering, MAusMM Steve Le Brun Principal Resource Geologist BSc(Hons), MSc, MAusIMM, MMICA Chris Johns Associate Geotechnical/Environmental Engineer MSc, P.Eng.(Alberta), MIEAust Peter Hayward Senior Process Engineer (Sedgman Ltd) Dip Metallurgy, MAusIMM Aaron Massey Senior Process Engineer (Sedgman Ltd) Dip Metallurgy, MAusIMM Chris Orr Associate Engineering Geologist MSc, MAusIMM, MAIG
Date: 30 June 2011
Project Number: MINEWPER00482AC
Version / Status: Final
Path & File Name: F:\MINE\Projects\Mawson West Ltd\MINEWPER00482AC_Mawson West_Kapulo Mine Eng\Report\43.101_June2011\CMWPr_482AC_MawsonWest_KapuloMineEng_43-101_30June2011_SEDAR.docx
Print Date: Thursday, 30 June 2011
Copies: Mawson West Ltd (2) Coffey Mining – Perth (1)
Document Change Control
Version Description (section(s) amended) Author(s) Date Final HW, JH 30 June 2011
Document Review and Sign Off
[Signed] [Signed] Primary Author Supervising Principal Harry Warries John Hearne
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Table of Contents
1 Summary ...... 1 1.1 Introduction ...... 1 1.2 Location ...... 1 1.3 Ownership ...... 1 1.4 Geology, Mineralisation ...... 2 1.5 Exploration Concept ...... 3 1.6 Exploration Status ...... 3 1.7 Metallurgical Assessment ...... 3 1.8 Resources ...... 3 1.9 Mineral Reserves ...... 4 1.10 Development and Operations ...... 5 1.11 Financial Summary ...... 5 1.12 Conclusions and Recommendations ...... 7
2 Introduction ...... 8 2.1 Scope of the Report ...... 8 2.2 Site Visit ...... 8 2.3 Principal Sources of Information ...... 8 2.4 Participants, Qualifications and Experience ...... 8 2.5 Independence ...... 10 2.6 Abbreviations ...... 10
3 Reliance on Other Experts ...... 12
4 Property Description and Location ...... 13 4.1 Background Information on Democratic Republic of Congo (DRC) ...... 13 4.1.1 Demographics and Geographic Setting ...... 13 4.1.2 History and Political Status ...... 13 4.1.3 Infrastructure ...... 13 4.1.4 Industry ...... 14 4.1.5 Mining ...... 14 4.2 Mineral Tenure ...... 14
5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ...... 17 5.1 Access ...... 17 5.2 Climate, Physiography & Local Resources and Infrastructure ...... 17
6 History ...... 18 6.1 Introduction ...... 18 6.2 Belgian Exploration ...... 18 6.3 Current Exploration ...... 18 6.4 Resource History ...... 19 6.5 Production History ...... 19
7 Geological Setting ...... 20 7.1 Regional Setting ...... 20 7.2 Project Geology ...... 22
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
8 Deposit Types ...... 24
9 Mineralization ...... 25
10 Exploration ...... 31 10.1 Belgian Exploration ...... 31 10.2 Falconbridge-Seremi Exploration ...... 31 10.3 Mawson West Exploration ...... 32
11 Drilling ...... 34 11.1 Introduction ...... 34 11.2 Mawson West Drilling ...... 34 11.3 Survey Control ...... 35 11.4 Collar Pickups ...... 36 11.5 Topography and Pit Surveys ...... 36 11.6 Drilling Orientation ...... 36 11.7 Oxidation ...... 36 11.8 Relationship Between Sample Length and Mineralised Thickness ...... 36
12 Sampling Method and Approach ...... 37 12.1 Soil Sampling Procedures and Sample Transportation ...... 37 12.2 Diamond Core Sampling ...... 37 12.3 Diamond Core Recovery ...... 38 12.4 Diamond Core Logging ...... 38 12.5 RC Drilling ...... 38 12.6 RC Sampling and Logging ...... 38 12.7 Sample Quality ...... 39
13 Sample Preparation, Analyses and Security ...... 40 13.1 Sample Security ...... 40 13.2 Analytical Laboratories ...... 40 13.3 Sample Preparation and Analytical Procedure ...... 40 13.3.1 Drillcore Analyses ...... 40 13.3.2 RC Sampling and Analyses ...... 41 13.3.3 Soil Sampling Analyses ...... 41 13.4 Bulk Density Determinations ...... 41 13.5 Adequacy of Procedures ...... 41
14 Data Verification ...... 42 14.1 Standards and Blanks – Mawson West ...... 42 14.1.1 Umpire Assay Checks ...... 43 14.1.2 Drillhole Twinning ...... 43 14.1.3 Laboratory Blanks and Standards ...... 45 14.2 Duplicates ...... 47 14.3 Data Quality Summary ...... 48
15 Adjacent Properties ...... 49
16 Mineral Processing and Metallurgical Testing ...... 50 16.1 Introduction ...... 50 16.2 Metallurgical Samples ...... 50 16.2.1 Head Assays and Mineralogy ...... 51
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
16.2.2 Comminution ...... 52 16.2.3 Oxide Flotation ...... 52 16.2.4 Sulphide Flotation ...... 53 16.2.5 Detailed Concentrate Assays ...... 54 16.2.6 Thickening Testwork ...... 54 16.2.7 Filtration Testwork ...... 55 16.3 Process Design ...... 55
17 Mineral Resource and Mineral Reserve Estimates ...... 57 17.1 Shaba Resource Estimate ...... 57 17.1.1 Resource Database and Validation ...... 57 17.1.2 Geological Interpretation and Modelling ...... 58 17.1.3 Statistical Analysis of Composites and Top Cuts...... 60 17.1.4 Bulk Density Data ...... 63 17.1.5 Variography ...... 66 17.1.6 Block Model ...... 71 17.1.7 Grade Estimation ...... 73 17.1.8 Resource Reporting and Classification ...... 76 17.1.9 Comments and Recommendations ...... 78 17.2 Safari Resource Estimate ...... 78 17.2.1 Resource Database and Validation ...... 79 17.2.2 Geological Interpretation and Modelling ...... 80 17.2.3 Statistical Analysis of Composites and Top Cuts...... 82 17.2.4 Bulk Density Data ...... 84 17.2.5 Variography ...... 85 17.2.6 Block Model ...... 90 17.2.7 Grade Estimation ...... 91 17.2.8 Resource Reporting and Classification ...... 94 17.2.9 Comments and Recommendations ...... 97 17.3 Mineral Reserve Estimates ...... 97
18 Other Relevant Data and Information, ...... 99 18.1 Mining ...... 99 18.1.1 Mining Approach...... 99 18.1.2 Geotechnical Input ...... 99 18.1.3 Hydrogeology and Hydrology Input ...... 102 18.1.4 Mine Waste Geochemical Testwork Results ...... 103 18.1.5 Contract Mining ...... 105 18.1.6 Pit Optimisation ...... 105 18.1.7 Mine Design ...... 111 18.1.8 Mine Production Schedule ...... 114 18.2 Proposed Processing Operations ...... 116 18.2.1 Crushing Circuit ...... 118 18.2.2 Crushed Ore Stockpiles ...... 119 18.2.3 Milling ...... 119 18.2.4 Flotation ...... 120 18.2.5 Concentrate Handling ...... 121 18.2.6 Reagents ...... 121 18.2.7 Tailings ...... 123 18.2.8 Air Supply ...... 123 18.2.9 Water Services ...... 123 18.3 Tailings Storage and Raw Water Dam ...... 124
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
18.4 Infrastructure ...... 125 18.5 Personnel ...... 126 18.6 Markets ...... 126 18.7 Contracts ...... 127 18.8 Environmental ...... 127 18.8.1 Legislative Background ...... 128 18.8.2 Baseline Study...... 132 18.9 Taxes, Duties and Royalties ...... 144 18.10 Capital Cost Estimates ...... 144 18.10.1 Initial Fixed Capital Costs ...... 144 18.10.2 First Fill and Working Capital ...... 144 18.10.3 Deferred Capital Costs ...... 145 18.11 Operating Costs ...... 145 18.12 Economic Analysis ...... 147 18.12.1 Introduction ...... 147 18.12.2 Project Economics ...... 147 18.12.3 Sensitivity Analysis ...... 148 18.13 Project Implementation ...... 150 18.14 Risk Assessment ...... 151
19 Interpretation and Conclusions ...... 153
20 Recommendations ...... 154
21 References ...... 155
22 Certificates of Qualified Persons ...... 157
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
List of Tables
Table 1.8_1 – Kapulo Resource Statement June 2010 4 Table 1.9_1 – Shaba Reserve Statement - 22 June 2011 4 Table 1.11_1 – Life of Mine Operating Costs 6 Table 1.11_2 – Life of Mine Project Cash Flows 6 Table 2.6_1 – List of Abbreviations 11 Table 4.2_1 – Tenement Schedule 16 Table 6.1_1 – Historical Exploration Work Summary 18 Table 9_1 – Kapulo Significant Intercepts (Relevant Samples) 27 Table 11.2_1 – Kapulo Project Drilling Summary 34 Table 13.3.3_1 – Analytical Summary Soil Samples 41 Table 14.1_1 – Mean Values of Geostats Certified Reference Materials Standards 42 Table 14.1_2 – Certified Standards Summary 43 Table 14.1.2_1 – List of Drillholes Used For Twin Hole Comparisons 45 Table 14.1.3_1 – ALS Chemex Internal QAQC Checks 45 Table 14.2_1 – Field Duplicates Summary 47 Table 16.2_1 – Samples Submitted for Metallurgical Testwork 50 Table 16.2.1_1 – Head Grades 51 Table 16.2.2_1 – Comminution Test Results 52 Table 16.2.5_1 – Final Concentrate Assays Summary 54 Table 17.1.3_1 – Shaba – Summary of Raw Statistics for Copper and Silver by Mineralised Zone 61 Table 17.1.3_2 – Shaba - Summary Copper Statistics- No High-Grade Cuts 61 Table 17.1.3_3 – Shaba - Summary Silver Statistics- No High-Grade Cuts 61 Table 17.1.4_1 – Summary of Bulk Density Measurements by Lithology and Weathering Profile at Shaba 64 Table 17.1.4_2 – Shaba Mineralisation Bulk Density Statistics 65 Table 17.1.5_1 – Shaba - Variogram Parameters for the Combined Mineralised Zones 67 Table 17.1.6_1 – Shaba - Block Model Summary 72 Table 17.1.7_1 – Shaba - Sample Search Parameters for OK Estimate 73 Table 17.1.8_1 – Shaba - Confidence Levels of Key Categorisation Criteria 76 Table 17.1.8_2 – Shaba Deposit – Resource Estimate by Mineralisation Domain 77 Table 17.1.8_3 – Shaba Deposit – Resource Estimate by Oxidation 78 Table 17.2.3_1 – Safari - Summary of Raw Statistics for Copper and Silver by Mineralised Zone 82 Table 17.2.3_2 – Safari - Summary Statistics - 2m Composite Copper (No High Grade Cuts Applied) 83 Table 17.2.3_3 – Safari - Summary Statistics - 2m Composite Silver (No High Grade Cuts Applied) 83 Table 17.2.4_1 – Safari - Summary Bulk Density Measurements By Lithology and Weathering Profile 85 Table 17.2.4_2 – Safari Mineralisation Bulk Density Statistics 85 Table 17.2.5_1 – Safari - Variogram Parameters for the Combined Mineralised Zones 89 Table 17.2.6_1 – Safari - Block Model Summary 90 Table 17.2.7_1 – Safari - Sample Search Parameters for OK Estimate 92 Table 17.2.8_1 – Safari - Confidence Levels of Key Categorisation Criteria 95 Table 17.2.8_2 – Safari North Deposit – Resource Estimate 96 Table 17.2.8_3 – Safari South Deposit – Resource Estimate 96 Table 17.2.8_4 – Safari North & South Deposits– Resource Estimate by Oxidation 97 Table 17.3_1 – Kapulo Mineral Reserves 98 Table 18.1.2_1 – Summary Pit Wall Slope Parameters 101 Table 18.1.5_1 – Summary Contract Mining Costs 105 Table 18.1.6_1 – Source of Main Input Parameters 106
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Table 18.1.6_2 – Summary Processing Recoveries 106 Table 18.1.6_3 – Summary Whittle Four-X Input Parameters 107 Table 18.1.6_4 – Detailed Pit Optimisation Results based on Indicated Resources 110 Table 18.1.7_1 – Summary Material Breakdown by Pit Design 111 Table 18.1.8_1 – Summary Mine Production Schedule 115 Table 18.2_1 – Summary Process Design Criteria 117 Table 18.2.3_1 – Kapulo Grinding Circuit Key Parameters 120 Table 18.2.6_1 – Reagents and Consumables 122 Table 18.6_1 – Copper Price Projection 126
Table 18.8.2_1 – Ambient SO 2 Quality Results 138 Table 18.8.2_2 – Noise Quality Results 139 Table 18.8.2_3 – Surface Water Monitoring Sites 141 Table 18.8.2_4 – Lunkinda Upstream – Surface Water Quality 142 Table 18.8.2_5 – Lunkinda Downstream – Surface Water Quality 143 Table 18.10.1_1 – Summary Initial Capital Cost Estimate 145 Table 18.10.3_1 – Summary Deferred Capital Cost Estimate 145 Table 18.11_1 – Summary Operating Cost Derivation 146 Table 18.11_2 – Summary Mining Costs 146 Table 18.11_3 – Summary Processing Cost Estimate 146 Table 18.11_4 – Life of Mine Operating Costs 147 Table 18.12.2_1 – Life of Mine Project Cash Flows 147 Table 18.12.3_1 – Project Sensitivity to a Change Cu Price 148 Table 18.12.3_2 – Project Sensitivity to a Change Operating Costs 148 Table 18.12.3_3 – Project Sensitivity to a Change Capital Costs 149 Table 18.13_1 – Summary Key Project Milestones 151 Table 20_1 – Future Work Summary 154
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
List of Figures
Figure 1.2_1 – Kapulo Project Location Map 1 Figure 1.3_1 – Mineral Licenses of the Kapulo Copper Project 2 Figure 1.3_1 – Mineral Licenses of the Kapulo Project 2 Figure 4.2_1 –Mineral Licenses of the Kapulo Copper Project 15 Figure 7.1_1 – Regional Geology of Mawson’s convention area in the D.R.C 20 Figure 7.1_2 – Stratigraphy of Kapulo Region with known Styles of Mineralization 21 Figure 7.2_1 – Typical Cross Section (Shaba Deposit) 23 Figure 7.2_2 – Local Geology of the Kapulo Area 23 Figure 9_1 – Bornite Dominant Matrix in Granite Breccia – Shaba Deposit 26 Figure 9_2 – Shaba Grade-Thickness Projection 29 Figure 9_3 – Safari North Grade-Thickness Projection 30 Figure 14.1.1_1 – ALS vs. Genalysis; Cu% & Ag ppm Umpire Assays Comparisons 43 Figure 14.1.2_1 – Twin Hole Comparison RC versus DD Drillhole at Shaba 44 Figure 14.1.3_1 – Shaba Standard GBM303-2 46 Figure 14.1.3_2 – Safari North Standard GBM309-16 46 Figure 14.2_1 – Shaba Field Duplicate Scatterplot 47 Figure 14.2_2 – Safari North Field Duplicate Scatterplot 48 Figure 16.2_1 – Section View Showing Metallurgical Master Composite Location 51 Figure 17.1.2_1 – Shaba - Geological Cross-Section at 19,885mN 59 Figure 17.1.2_2 – Shaba – Mineralisation Cross-Section at 19,885mN 59 Figure 17.1.2_3 – Shaba –Weathering Cross-Section at 19,885mN 60 Figure 17.1.3_1 – Shaba - Histogram Plot for Cu% - Zone 100 (Hangingwall) 62 Figure 17.1.3_2 – Shaba - Histogram Plot for Cu% - Zone 110 (Footwall) 62 Figure 17.1.4_1 – Shaba - Bulk Density Regression Statistics for Hangingwall Domain (100) Fresh Material 63 Figure 17.1.4_2 – Shaba - Weathering Cross-Section of Final Block Model at 19,885mN 65 Figure 17.1.4_3 – Shaba - Geology Cross-Section of Final Block Model at 19,885mN 65 Figure 17.1.7_1 – Shaba - Copper (Combined Zones) – Gaussian Transformed Variography 68 Figure 17.1.5_2 – Shaba - Copper (Combined Zones) – Back Transformed Variography (to Original Variance) 69 Figure 17.1.5_3 – Shaba - Silver (Combined Zones) – Gaussian Transformed Variography 70 Figure 17.1.5_4 – Shaba - Silver (Combined Zones) – Back Transformed Variography (to Original Variance) 71 Figure 17.1.7_1 – Shaba Block Model Grades 74 Figure 17.1.7_2 – Shaba - Examples of Comparison Plots of Block Grades and Informing Composite Grades 75 Figure 17.1.8_1 – Shaba Classified Block Model 77 Figure 17.2.2_1 – Safari - Geological Cross-Section at 19,885mN 81 Figure 17.2.2_2 – Safari –Weathering Cross-Section at 19,885mN 82 Figure 17.2.3_1 – Safari - Histogram Plot for Cu% - Zone 200 (Hangingwall) 84 Figure 17.2.3_2 – Safari - Histogram Plot for Cu% - Zone 210 (Footwall) 84 Figure 17.2.5_1 – Safari - Copper Correlogram (Combined Zones) 86 Figure 17.2.5_2 – Safari - Silver Correlogram (Combined Zones) 87 Figure 17.2.5_3 – Safari - Density (Combined Zones) – Traditional Variogram 88 Figure 17.2.7_1 – Safari - NS Long Section Safari North Block Model Grades & Composites 93 Figure 17.2.7_2 – Safari - Examples of Comparison Plots of Cu Block Grades and Informing Composite Grades 93 Figure 17.2.7_3 – Safari - Examples of Comparison Plots of Ag Block Grades and Informing Composite Grades 94 Figure 17.2.8_1 – Safari North Classified Block Model 95 Figure 18.1.2_1 – Schematic Simplified Geological Cross-section of the Shaba Deposit 100 Figure 18.1.3_1 – Surface Water Management - General Arrangement 104
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Figure 18.1.6_1 – Commodity Price vs Mill Feed - Based on Indicated Resources only 108 Figure 18.1.6_2 – Summary Pit Optimisation Results - Based on Indicated Resources only 109 Figure 18.1.7_1 – Plan View of Starter Pit and Final Pit Design 112 Figure 18.1.7_1 – Mine Infrastructure Layout 113 Figure 18.1.8_1 – Summary Mine Production Schedule 115 Figure 18.1.8_2 – Summary Processing Schedule 116 Figure 18.8.2_1 – Proposed Kapulo Exploitation Permit Inside Current PR 133 Figure 18.8.2_2 – Satellite Imagery showing the Kapulo Project Area 133 Figure 18.8.2_3 – Seismic Risk Map of Southern Africa 135 Figure 18.8.2_4 – Kapulo Monthly Rainfall and Temperature 136 Figure 18.8.2_5 – Dikulushi Monthly Rainfall and Temperature 136 Figure 18.8.2_6 – Kawambwa Monthly Rainfall and Temperature 137 Figure 18.8.2_7 – Noise Contour Map of Kapulo Permit 139 Figure 18.8.2_8 – Kapulo Surface Water Drainage 141 Figure 18.12.3_1 – Summary Project Sensitivity Analysis 149
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
1 SUMMARY
1.1 Introduction
Mawson West has been developing the Kapulo Project (the Project) since April 2006. The Project is located in the Katanga Province of the Democratic Republic of Congo and consists of three deposits, namely Shaba, Safari North and Safari South. A Definitive Feasibility Study (DFS) was completed for the Shaba deposit during Q2 2011.
1.2 Location
The Project is located in the southeastern corner of the Democratic Republic of Congo (DRC), approximately 125km northeast of Dikulushi and 45km east-northeast of the town of Pweto, a small town at the northern edge of Lake Mweru, at a latitude of 08°:18:10S and longitude of 29°:14:03E and (Figure 1.2_1). The principle area of exploration is situated along the Kapulo Fault in the Proterozoic Kundelungu Group sediments in the southeast corner of the country.
Figure 1.2_1 Kapulo Project Location Map
1.3 Ownership
The Project is located in the Haut Katanga District, Katanga Province of the Democratic Republic of the Congo (DRC). The Project 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.
Kapulo Copper Project, DRC – MINEWPER00482AC Page: 1 National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
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 (“Mawson West”), hold 90% of the issued capital of CMCC, the remaining 10% is held by the Dikulushi – Kapulo Foundation NPO.
Figure 1.3_1 Mineral Licenses of the Kapulo Copper Project
1.4 Geology, Mineralisation
The Kapulo copper deposits are interpreted to be hypogene, fault controlled deposits, comprising disseminated/massive/brecciated chalcopyrite-bornite mineralisation with massive to nodular cuprite and supergene azurite and malachite. The mineralisation is litho- structurally controlled, and is hosted in shales and sandstones of the Kundelungu Group and in the footwall granite.
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1.5 Exploration Concept
The company has completed soil sampling over the prospective Kapulo Fault and over the Kanke Fault which has resulted in a number of Cu in soil targets. Exploration work is currently focused on testing these soil anomalies with trenching and then drilling. Additional areas have been stream sediment sampled and these areas will also be followed up. The company has recently completed a CSAMT/IP geophysical survey along the Kapulo Fault and has identified a number of targets which require drill testing in 2011.
1.6 Exploration Status
Exploration is at an advanced stage at Shaba and Safari North with the resource at Shaba sufficiently defined by drilling for an open pit resource. Further drilling is required to define the depth extents of both resources for potential underground extensions. Safari South has been drilled on a nominal 40m by 30m grid and additional drilling is required to fully define this resource which may prove to join up with Safari North.
1.7 Metallurgical Assessment
Metallurgical work has focussed on the Shaba deposit to date with samples collected and stored at AMDEL Perth for future work on the Safari North oxide ore. A series of batch flotation tests were completed on the Shaba sulphide mineralisation. The composite was spatially and geologically representative of the Shaba Deposit. Based on the batch tests a locked cycle test was also completed on this composite which gave a Cu recovery of 93.2% (Ag recovery of 90%) at a grade of 34% Cu and 90g/t Ag in the concentrate. ICP analysis of a suite of elements showed that there were no penalty elements present in significant quantities to attract a penalty in the Shaba sulphide concentrate. A spatially representative master composite for the Shaba oxide was also sent to AMDEL and a series of batch flotation tests were used to optimise the flotation conditions which lead to a locked cycle test which gave a recovery of 71.6% Cu for a grade of 48% Cu in the concentrate. A total of 10 new Diamond Drill Holes (DDH) were completed during May-August 2010 to provide samples for the sulphide metallurgical variability testwork. The variability testwork confirmed the validity of the previous Shaba sulfide flotation testwork results.
1.8 Resources
A lithological based interpretation of the mineralized zones and the resource block model for Shaba was completed in June 2010 and Safari North and Safari South were completed in December 2010 by Steve Le Brun of Coffey Mining Pty Ltd of Perth, Western Australia a geological and mining consulting company listed on the ASX and Adam Anderson of Mawson West Ltd of Perth, Western Australia. Modelling was completed in Datamine mining software. The results of the resource estimation are shown in Table 1.8_1 below.
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Table 1.8_1 Kapulo Copper Project Kapulo Resource Statement June 2010 Reported Above 0.3% Cu Cutoff Ordinary Kriged Estimate Using 2m Cut Cu% Composites Parent Cell Dimensions of 10mEW by 20mNS by 5mRL Depleted for Artisanal Surface Mining
Grade Cu Metal Grade Ag Metal Classification Zone Mt (Cu%) (Tonnes) (Ag ppm) (MOz) Shaba Deposit HW 1.710 7.7 131,650 23 1.24 Indicated FW 3.280 1.4 45,650 3 0.30 Subtotal 4.990 3.6 177,300 10 1.54 HW 0.195 7.4 14,400 30 0.19 Inferred FW 0.829 1.1 8,800 5 0.12 Subtotal 1.024 2.3 23,200 10 0.31 Safari North Deposit HW 0.509 4.8 24,360 10 0.16 Indicated FW 0.451 1.8 7,950 3 0.05 Subtotal 0.959 3.4 32,310 7 0.21 HW 0.313 4.1 12,875 8 0.08 Inferred FW 0.693 1.3 9,075 3 0.06 Subtotal 1.006 2.2 21,950 4 0.14 Safari South Deposit HW Indicated FW Subtotal HW 0.098 4.2 4,150 1 0.01 Inferred FW 0.294 1.1 3,095 1 0.01 Subtotal 0.392 1.8 7,245 1 0.01 Note: Figures have been rounded
1.9 Mineral Reserves
The following Mineral Reserves were determined for the Shaba Deposit. At this stage no Reserves were determined for either Safari North or Safari South.
Table 1.9_1 Kapulo Copper Project Shaba Reserve Statement 22 June 2011 Reported Above 0.3% Cu Cutoff 5% Dilution applied at zero grade
In-situ In-situ Classification Tonnes Cu Grade Ag Grade Cu Metal Ag Metal (Mt) (%) (kt) (Ag ppm) (kOz) Proven - - - - - Probable 3.6 3.6 128.0 8.3 954.8 Total 3.6 3.6 128.0 8.3 954.8
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This reserve estimate has been determined and reported in accordance with Canadian National Instrument 43-101, ‘Standards of Disclosure for Mineral Projects’ of December 2005 (the Instrument) and the classifications adopted by CIM Council in November 2010.
1.10 Development and Operations
The DFS that was undertaken for the Shaba deposit indicated that the annual production for Shaba will be, on average, 15.8kt of copper and 78koz of silver, over a 7.5 year mine life. The plant is expected to treat 0.5Mtpa of mill feed.
All ore and waste will be mined via conventional, open pit mining methods using a mining contractor.
The operation is planned to utilise selective mining techniques to separate ore and waste. The main mine production equipment that is considered to be appropriate for the Project includes 60t to 100t excavators and haul trucks with a payload of between 40t and 60t.
Provision has been made for drilling and blasting from surface.
Mill feed will be processed in a conventional single stage ball mill with two ball mills operating in parallel.
Testwork indicates that expected copper recoveries for the Shaba deposit, based on the selected treatment route, will be 93.2% and 71.6% for sulphide and oxide material respectively.
The Project will employ approximately 400 people throughout the operating phase of the project. Initially selected posts requiring specific skills or experience will be filled by expatriates. In addition to performing their job function, expatriate personnel will be expected to transfer knowledge and expertise in order to develop the capabilities of the national staff. In the longer term, it is anticipated that nationals of the DRC will fill most operating and management positions within the company. In addition, the mining contractor will employ approximately 180 people for a total for the Project of 580 people.
Power will be supplied by way of 4 x 1000mW diesel run generators.
The water supply will be from three local streams and pumped to the plant. Total water requirement will be 150m³/hr.
It is expected that the Project construction will take 17 month starting in July 2011, with commissioning of the process plant scheduled for November 2012.
1.11 Financial Summary
Mawson West has prepared a financial model to evaluate the economics of the Project. The model is presented as an equity model assuming 100% equity financing. No allowance has been made in the model for the effects and levels of debt financing available or required.
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Operating costs are anticipated to be $1.97/lb copper as shown in Table 1.11_1.
Table 1.11_1 Kapulo Copper Project Life of Mine Operating Costs
Value Unit Cost Item [M$] [$/lb copper] Mining Costs 121.4 0.47 Treatment Costs 128.5 0.50 On Site General and Administration Costs 43.8 0.17 Sub Total 293.7 1.14 Marketing Duties and Taxes 46.5 0.18 Cu Sales and Transport Costs 166.1 0.65 Sub Total 212.6 0.83 Total 506.3 1.97
The results of the financial model are presented in Table 1.11_2 on the basis of the assumptions applied as follows:
Weighted average copper price over the life of mine: US$7,975/t (US$3.62/lb).
No silver credits applied.
Discount rate of 10%
Table 1.11_2 Kapulo Copper Project Life of Mine Project Cash Flows
Item Units Value Net Revenue US$M 735 Capital Expenditure US$M (90) Operating Expenses US$M (294) Marketing Duties and Indirect Taxes US$M (46) Free Cash Flow US$M 305 Net Present Value (@ 10%) US$M 157 Mawson West portion (90%) US$M 141 Internal Rate of Return % 61
DRC’s corporate income taxes are applicable once the sunk exploration costs, acquisition costs and capital costs have been recovered by the Project. In addition, under the Dikulushi Mining Convention the project is exempt from taxes for the first five years. As such, DRC taxes are only payable for the last two and a half years of the Project life.
The principal results of the financial evaluation are as follows:
Internal Rate of Return 61%
Net Present Value at 10% US$157 million
Capital Investment US$69.5 million
Free Cash Flow US$305 million
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Sensitivity studies have been undertaken on the financial model for the following scenarios:
Copper Price ± 20%, in 10% increments.
Operating Costs ± 20% in 10% increments.
Capital Costs ± 20% in 10% increments.
Discount Rate ± 2% in absolute terms.
The sensitivity analysis showed that the Project is most sensitive to a change in Cu price. This is a common phenomenon where the parameters that affect the revenue will have the greatest impact on a project, with a change in the Cu grade or processing recovery exhibiting a similar result as shown for the copper price. A 10% change in the Cu price resulted in an approximate 34% change in the Net Present Value (NPV).
1.12 Conclusions and Recommendations
The results of Shaba DFS indicate the economic viability of exploiting the Shaba deposit.
It is recommended that Mawson West proceeds with:
The construction of the process plant and other infrastructure.
Contract negotiations with a mining contractor.
Additional metallurgical testwork on the Safari North oxide and sulfide material to assess processing recoveries.
Further geochemical testwork on tailings and waste rock samples to further assess the potential of acid rock drainage.
Development of detailed TSF and RWD construction specifications and scopes of work.
Assess raw water dam water balance and filling rate schedule.
Further infill and extensional drilling to raise the level of confidence and extend the Inferred Resources at both Shaba and Safari North.
Detailed ground water assessment during the first six months of mining to confirm observations made during the exploration program.
Further resource definition drilling as the geophysics completed during 2010 has provided a number of targets for drilling in 2011 and, as such, there is reasonable potential to increase resources within trucking distance of the Project.
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2 INTRODUCTION
2.1 Scope of the Report
In November, 2009 Coffey Mining Pty Ltd (Coffey Mining) was commissioned by Mawson West Limited (Mawson West) to prepare a NI 43-101 compliant Technical Report that details the Definitive Feasibility Study undertaken for the Shaba deposit, which forms part of the greater Kapulo Copper Project (the Project), located within the northern Katanga region of the Democratic Republic of Congo (DRC).
This report is to comply with disclosure and reporting requirements set forth in National Instrument 43-101, Companion Policy 43-101CP, and Form 43-101F1.
The report complies with Canadian National Instrument 43-101 for the ‘Standards of Disclosure for Mineral Projects’ of December 2005 (the Instrument) and the resource and reserve classifications adopted by CIM Council in November 2010. The report is also consistent with the ‘Australasian Code for Reporting of Mineral Resources and Ore Reserves’ of December 2004 (the Code) as 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).
2.2 Site Visit
Steve Le Brun visited the Project and surrounding areas for two days in April 2010 to assess the project, available data, and the data collection protocols
2.3 Principal Sources of Information
In addition to site visits undertaken to the DRC copper projects in 2010 by Steve Le Brun, this report has relied extensively on information provided by Mawson West, extensive discussions with Mawson West technical personnel, various consultants and contractors. The principal sources of information used to compile this report comprise supplied digital data and some published information relevant to the Project area and the region in general. A full listing of the principal sources of information is included in Section 21 of this report and in Section 3.
Coffey Mining has made all reasonable enquiries to establish the completeness and authenticity of the information provided and identified, and a final draft of this report was provided to Mawson West along with a written request to identify any material errors or omissions prior to lodgement.
2.4 Participants, Qualifications and Experience
Coffey Mining was responsible for the preparation of all portions of this report apart from Sections 4 to 11 and the associated text in the summary, conclusions and discussion.
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Coffey Mining is an integrated Australian-based consulting firm, which has been providing services and advice to the international mineral industry and financial institutions since 1987. In September 2006, Coffey International Limited acquired RSG Global. Coffey International Limited is a highly respected Australian-based international consulting firm specialising in the areas of geotechnical engineering, hydrogeology, hydrology, tailings disposal, environmental science and social and physical infrastructure.
The primary author of this report is Mr Harry Warries of Coffey Mining who is responsible for the entire document as the qualified person and principal author. Mr Harry Warries is a professional mining engineer with 20 years experience in the mining industry. Mr Warries, a Principal Mining Consultant with Coffey Mining and a Member of the Australasian Institute of Institute of Mining and Metallurgy (AUSIMM), has the appropriate relevant qualifications, experience and independence to generally be considered as a “Qualified Person” as defined in Canadian National Instrument 43-101. Mr Warries has not visited the Project.
Mr Steve Le Brun is a professional geologist with 20 years experience in exploration and mining geology. Mr Le Brun, a Principal Resource Consultant with Coffey Mining and a Member of the Australasian Institute of Institute of Mining and Metallurgy (AUSIMM) and a member of the Mineral Industry Consultants Association (MICA), has the appropriate relevant qualifications, experience and independence to generally be considered as a “Qualified Person” as defined in Canadian National Instrument 43-101. Mr Le Brun visited the Project in April 2010.
Mr Adam Anderson is a professional geologist with 17 years experience in exploration and mining geology and has contributed to Sections 4 to 11 of the report. Mr Anderson is the GM Resource Development with Mawson West Ltd. and is a Member of the Australasian Institute of Institute of Mining and Metallurgy (AUSIMM). Mr Anderson has visited the Project on at least 12 occasions, most recently in Nov 2010.
Mr Peter Hayward is a professional process engineer with 36 years of experience in metallurgical process design. Mr Peter Hayward is a Senior Process Engineer with Sedgman Ltd and a member of the Australasian Institute of Institute of Mining and Metallurgy (AUSIMM) and has the appropriate relevant qualifications, experience and independence to generally be considered as a “Qualified Person” as defined in Canadian National Instrument 43-101.
Mr Aaron Massey is a professional process engineer with 16 years of experience in metallurgical process design. Mr Peter Hayward is a Senior Process Engineer with Sedgman Ltd and a member of the Australasian Institute of Institute of Mining and Metallurgy (AUSIMM) and has the appropriate relevant qualifications, experience and independence to generally be considered as a “Qualified Person” as defined in Canadian National Instrument 43-101.
Mr Christopher Martin Orr is an engineering geologist with 38 years experience in civil and mining geotechnics. Mr Orr, Principal and Director of George, Orr and Associates (Australia) Pty Ltd and a Member of the Australasian Institute of Mining and Metallurgy (AusIMM) and The Australian Institute of Geoscientists (AIG), has the appropriate relevant qualifications, experience and independence to generally be considered as a “Qualified Person” as defined in Canadian National Instrument 43-101. Mr Orr visited the Project in November 2008.
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Mr Chris Johns of Coffey Mining is responsible for the Kapulo Project tailings storage and water storage facility design report. Mr Chris Johns is a professional geological engineer with 15 years experience in geotechnical consulting. Mr Johns, an Associate Consultant with Coffey Mining and a Member of the Institution of Engineers Australia (MIEAust) and a registered Professional Engineer with the Association of Professional Engineers, Geologists, and Geophysicists of Alberta (APEGGA), has the appropriate relevant qualifications, experience and independence to generally be considered as a “Qualified Person” as defined in Canadian National Instrument 43-101. Mr Johns has not visited the Project.
Certificates for the Qualified Persons are located in Section 22.
2.5 Independence
Coffey Mining is part of Coffey International Limited (CIL), a highly respected Australian-based international consulting firm specialising in the areas of exploration, geology, mining, metallurgy, geotechnical engineering, hydrogeology, hydrology, tailings disposal, environmental science and social and physical infrastructure.
Neither Coffey Mining, nor the authors of this report, except for Mr Chris Orr, have or have had previously any material interest in Mawson West or related entities or interests. Its relationship with Mawson West is solely one of professional association between client and independent consultant. This report is 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.
Mr Chris Orr owns Ordinary Fully Paid Shares in Mawson West Ltd, issued in March 2009 as part payment of consulting fees. Apart from these shares, Mt Orr does not have nor does he expect to receive a direct or indirect interest in the Kapulo property of Mawson West Ltd and does not own, directly or indirectly, any other securities of Mawson West Ltd or any associate or affiliate of such companies.
Mr Anderson is a full-time employee of Mawson West and owns Ordinary Fully Paid Shares in Mawson West Ltd.
2.6 Abbreviations
A listing of abbreviations used in this report is provided in Table 2.6_1 below.
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Table 2.6_1 List of Abbreviations
Description Description -75µ 75 microns l/hr/m² litres per hour per square metre $ United States of America dollars M million “ inches m metres µm microns Ma thousand years 3D three dimensional Mg Magnesium AAS atomic absorption spectrometry ml millilitre Ag Silver mm millimetres bcm bank cubic metres Mtpa million tonnes per annum CC correlation coefficient N (Y) northing cm centimetre Ni nickel Co cobalt NPV net present value
CRM certified reference material or certified standard NQ 2 size of diamond drill rod/bit/core Cu copper ºC degrees centigrade CV coefficient of variation OK Ordinary Kriging
DDH diamond drillhole P80 80% passing DTM digital terrain model Pd palladium E (X) easting ppb parts per billion EDM electronic distance measuring ppm parts per million Fe iron psi pounds per square inch g gram PVC poly vinyl chloride g/m³ gram per cubic metre QC quality control GPR Ground Penetrating Radar QQ quantile-quantile g/t gram per tonne of gold RAB Rotary Air Blast HARD half the absolute relative difference RC reverse circulation HDPE high density poly ethylene RL (Z) reduced level NQ size of diamond drill rod/bit/core ROM run of mine hr hours RQD rock quality designation HRD half relative difference SD standard deviation ICP-AES inductivity coupled plasma atomic emission spectroscopy SG Specific gravity ICP-MS inductivity coupled plasma mass spectroscopy Si silicon ISO International Standards Organisation SMU selective mining unit kg kilogram t tonnes kg/t kilogram per tonne t/m³ tonnes per cubic metre km kilometres tpa tonnes per annum km² square kilometres w:o waste to ore ratio kW kilowatts kWhr/t kilowatt hours per tonne
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3 RELIANCE ON OTHER EXPERTS
Neither Coffey Mining nor the authors of this report are qualified to provide extensive comment on legal issues, including status of tenure, and taxation associated with the Kapulo property referred to in this report. Assessment of these aspects has relied heavily on information provided by Mawson West, which has not been independently verified by Coffey Mining, and this report has been prepared on the understanding that the property is, or will be, lawfully accessible for evaluation, development, mining and processing.
Similarly, Coffey Mining nor the authors of this report are qualified to provide extensive comment on hydrogeological, environmental and social issues. The hydrogeological issues and its descriptions are based on studies issued by Mawson West and its hydrogeological consultants Groundwater Resource Management Pty Ltd (GRM). Coffey Mining has relied on the report ‘Kapulo Project, Democratic Republic Of Congo, Stage 2 Hydrogeological Investigation’ dated November 2010 provided to Coffey by Mawson West and prepared for Mawson West by GRM. The environmental and social issues and its descriptions are based on environmental studies issued by Mawson West and its environmental consultants. Coffey Mining has relied on “The Kapulo Copper Project, Environmental Impact Assessment” dated May 2011 provided to Coffey by Mawson West and prepared for Mawson West by African Mining Consultants Limited of Kitwe, Zambia.
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4 PROPERTY DESCRIPTION AND LOCATION
4.1 Background Information on Democratic Republic of Congo (DRC)
4.1.1 Demographics and Geographic Setting
The country has a population of 71 million people which are made up of over 200 African ethnic groups of which the majority are Bantu. The four largest tribes make up about 45% of the population. The main religion is Roman Catholic 50%, with the remainder being Protestant 20%, Kimbanguist 10%, Muslim 10% and other 10% (includes syncretic sects and indigenous beliefs). The official language is French with a number of African dialects also spoken. The majority of people live outside major towns (66%) with a subsistence type of lifestyle, with an urbanization rate of about five percent per year.
4.1.2 History and Political Status
Established as a Belgian colony in 1908, the Republic of the Congo gained its independence in 1960, but its early years were marred by political and social instability. Col. Joseph Mobutu seized power and declared himself president in a November 1965 coup. He subsequently changed his name - to Mobutu Sese Seko - as well as that of the country - to Zaire. Mobutu retained his position for 32 years. Ethnic strife and civil war, touched off by a massive inflow of refugees in 1994 from fighting in Rwanda and Burundi, led in May 1997 to the toppling of the Mobutu regime by a rebellion backed by Rwanda and Uganda and fronted by Laurent Kabila. He renamed the country the Democratic Republic of the Congo (DRC), but in August 1998 his regime was itself challenged by a second insurrection again backed by Rwanda and Uganda. Troops from Angola, Chad, Namibia, Sudan, and Zimbabwe intervened to support Kabila's regime. A cease-fire was signed in July 1999 by the DRC, Congolese armed rebel groups, Angola, Namibia, Rwanda, Uganda, and Zimbabwe but sporadic fighting continued. Laurent Kabila was assassinated in January 2001 and his son, Joseph Kabila, was named head of state. In October 2002, the new president was successful in negotiating the withdrawal of Rwandan forces occupying eastern Congo; two months later, the Pretoria Accord was signed by all remaining warring parties to end the fighting and establish a government of national unity. A transitional government was set up in July 2003. Joseph Kabila as president, and four vice presidents represented the former government, former rebel groups, the political opposition and civil society. The transitional government held a successful constitutional referendum in December 2005 and elections for the presidency, National Assembly, and provincial legislatures in 2006. The National Assembly was installed in September 2006 and Joseph Kabila was inaugurated president in December 2006. Provincial assemblies were constituted in early 2007, and elected governors and national senators in January 2007.
4.1.3 Infrastructure
The infrastructure is slowly being improved after many years of neglect due to civil unrest. The country has 154,000km of roads of which 3,000km are paved. The Katanga region has no paved roads and virtually no power supply. Cellular telephone coverage is improving with signal available near the town of Pweto on the northern shore of Lake Mweru. Power is available in Pweto from Zambia. No rail network exists near the Project area, although barge transport is possible on Lake Mweru.
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4.1.4 Industry
The main industries are mining (diamonds, gold, copper, cobalt, coltan, zinc), mineral processing, consumer products (including textiles, footwear, cigarettes, processed foods and beverages), cement and commercial ship repair. The country’s Gross Domestic product (GDP) estimate in 2009 was $11b with a per capita GDP estimate of $300 million.
Following an absence of approximately ten years, the International Monetary Fund and the World Bank have re-engaged the DRC and are assisting the development of coherent legislative and economic reforms, aimed at a reconstruction of the country. As part of this effort, in 2002 the government introduced a new Mining Code. Gécamines, a state owned mining company, holds significant quantities of Mineral Reserves and Mineral Resources of copper, cobalt, germanium and zinc in the Katanga province of DRC.
4.1.5 Mining
The economy of the DRC has historically been dominated by its resource sector. The Congo Copperbelt region of the country, in the southern province of Katanga, is renowned as one of the richest mineral regions of the world and until the mid 1980's enabled the country to be one of the largest producers of copper, with annual production exceeding 500,000 tonnes of copper. Adverse political events beginning in the early 1990s, together with military activity, have led to a dramatic reduction in national output. However, diamonds, copper and cobalt remain the principal foreign exchange earning exports for the country.
4.2 Mineral Tenure
Mawson West Limited, via 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 were first granted on the 22 May 2001 and were renewed for a further five years early this year as shown in Table 4.2_1 below.
In accordance with Article 49 of the DRC Mining Code, CMCC holds PR1697, which is extended until a decision regarding the exploitation permit application has been processed by the DRC government.
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 exploitation permit is issued once an Environmental Impact Study has been approved. 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.
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Figure 4.2_1 Mineral Licenses of the Kapulo Copper Project
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Table 4.2_1 Kapulo Copper Project Tenement Schedule
Project Group Entity Permit No. Area km² Type Granted Expiry Dikulushi AMC PE606 40.5 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 PR1685 399.0 Exploration 12-Apr-11 11-Apr-16 Kapulo AMC PR1686 395.7 Exploration 22-May-11 21-May-16 Kapulo AMC PR1688 398.8 Exploration 12-Apr-11 11-Apr-16 Kapulo AMC PR1689 398.8 Exploration 22-May-11 21-May-16 Kapulo AMC PR1690 398.9 Exploration 22-May-11 21-May-16 Dikulushi AMC PR1693 398.6 Exploration 12-Apr-11 11-Apr-16 Dikulushi AMC PR1694 398.5 Exploration 12-Apr-11 11-Apr-16 Kapulo AMC PR1697 398.7 Exploration Held Held Dikulushi AMC PR1700 398.4 Exploration 12-Apr-11 11-Apr-16 Dikulushi AMC PR1703 398.3 Exploration 22-May-11 21-May-16 Dikulushi AMC PR1705 237.0 Exploration 22-May-11 21-May-16 Dikulushi AMC PR1706 398.0 Exploration 22-May-11 21-May-16 Dikulushi AMC PR1707 397.7 Exploration 23-May-11 22-May-16 Dikulushi AMC PR1708 405.1 Exploration 22-May-11 21-May-16 Dikulushi AMC PR1709 345.0 Exploration 22-May-11 21-May-16 Dikulushi AMC PR1710 397.0 Exploration 22-May-11 21-May-16 Dikulushi AMC PR1711 396.9 Exploration 22-May-11 21-May-16 Total Area 7,283
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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY
5.1 Access
The Project is accessed from the Zambian side of Lake Mweru. There are sealed roads all the way through Zambia to Nchelenge (on Lake Mweru), followed by 80km of unsealed road from Nchelenge to Pweto and a further 50km from Pweto to Kapulo. The closest international airport is in Lubumbashi, 530km to the south and an all weather airstrip is located at Pweto, located 53km from the Kapulo Project.
Supplies for the Project will be trucked on sealed roads from South Africa via Botswana to Nchelenge and then on to Kapulo via Pweto.
5.2 Climate, Physiography & Local Resources and Infrastructure
The western part of the greater Kapulo area is characterized by a large 20km to 30km wide floodplain, which represents the northern extension of the Lake Mweru depression at approximately 1,000m above sea level. The plain is bordered to the west by the Kundelungu plateaus and to the east by the Marungu Plateau. The contact between the plain and the Marungu Plateau is characterized by a N-S escarpment of 650m height. The neighbouring area is almost entirely covered with natural woodland and forest, interrupted in the low areas by fields of cassava and by localized clearing. The Project is in the western foothills of the Bangulian Granite complex. The Kapulo Copper deposits outcrop in the steep N-S valley which forms the contact with the lower Makana and Kapona hills and the Marungu Plateau. The valley has an average elevation of 1,200m above sea level. Few rivers drain into Lake Mweru from the DRC. A minor ephemeral stream runs nearby the Kapulo site. Lake Mweru is fed by the Luapula River, which along with Lake Mweru forms the international boundary between Zambia and the DRC. The climate of the area is tropical with a distinct wet and dry season. The wet season begins towards the end of September and finishes at the end of April.
The average rainfall, as indicated by mission records, is 1,260mm, with a range of 800mm to 2,200mm. An Oregon Scientific weather station was installed at Kapulo in June 2009. Weather data collected at the Kapulo site shows a total of 715mm of rain over 12 months. Low average monthly temperatures of 12°C were recor ded from June to August and monthly average high temperatures varied from 25°C to 33°C. The wet season does inhibit exploration activities and access to some areas within the PR, with flooded roads and rivers and the terrain becoming difficult for vehicular access. Labour is sourced locally for camp and geological and drilling assistant type work.
The only infrastructure existing at Kapulo is the road from Pweto as no fuel or grid power is available at site. All fuel required is trucked in by the company. The company is currently upgrading the road access from Pweto to Kapulo to an all weather road. There is grid power and some trained labour available in Pweto, especially tradesman.
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6 HISTORY
6.1 Introduction
The development of exploration and mining at the Project and surrounds can be broken down into the following periods:
Historical - Belgian (Simikat) and Falconbridge Exploration Work.
Current - Mawson West.
Table 6.1_1 provides an overview of the historical exploration work.
Table 6.1_1 Kapulo Copper Project Historical Exploration Work Summary
Year Supervision Exploration Completed 1906-1954 Tanganyika/Simikat Adit sampling, Metallurgy 1970-1972 Falconbridge Adits, DDH, Soils Geochem, Geophysics, Metallurgy 1998 Anvil Mining Ltd Dikulushi Mining Agreement signed 1998-2002 Anvil Mining Ltd Civil War prevented work 2006 Mawson West JV Agreement between Anvil and Mawson West 2006-2007 “ Mapping and Reconnaissance sampling, Soil Geochemistry, Aeromagnetics 2008 “ Diamond drilling, Metallurgical Sampling 2009 “ Diamond drilling, Resource Estimate, Met Testwork 2010 “ Met testwork, DDH, RC Drilling, Resource Est, ESIA
6.2 Belgian Exploration
Copper mineralisation in the area was initially evaluated by the Belgians (Simikat) from 1910 until 1923. The two main areas of interest were Shaba and Safari North/South, which are about 2.4km apart. The Belgians drove numerous adits into the Safari and Shaba deposits in order to evaluate the copper potential of the area. Other gossanous outcrops were sampled by costeans and trenching works.
Falconbridge Exploration Limited obtained an interest in the Kapulo Deposits during the early 1970’s and completed adit sampling, diamond drilling, metallurgical testwork, soil geochemistry and geophysics. The Projects lay dormant until Anvil Mining pegged the ground in the late 1990’s and subsequently signed the JV agreement with Mawson West.
6.3 Current Exploration
Anvil Mining Limited obtained the Kapulo licences in 1998 and only completed reconnaissance type work due to civil war and unrest in the area until 2002. After 2002, Anvil concentrated on the successful development of the Dikulushi deposit.
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Mawson West has completed the following work on the Kapulo Licences since the JV agreement was signed in 2006:
Mapping and reconnaissance sampling completed by consultant Walter Witt in order to provide information on mineralisation controls.
A low level airborne geophysical survey was completed in 2007 by UTS.
Diamond and reverse circulation (RC) drilling at the Shaba and Safari north deposits to define a JORC compliant resource estimate.
Regional soil geochemical sampling program.
CSAMT/IP ground geophysical survey.
RTK GPS topographical survey of the site.
6.4 Resource History
No published, JORC or National Instrument 43-101 compliant resources were completed on the Project prior to the Mineral Resources that were reported in the Technical Report “Kapulo Project, DRC, National Instrument 43-101 Technical Report” dated 3 February 2011.
No update to the Mineral Resources as reported in the aforementioned Technical Report has been undertaken thus far.
6.5 Production History
No historic copper production has been recorded on any of the current project tenements.
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7 GEOLOGICAL SETTING
7.1 Regional Setting
Mawson West’s concession area contains an area of gently folded Katangan sediments which are bounded to the east by the Banguelo Granite. The contact between the sediments and the granite is the Kapulo Fault which is orientated approximately north-south along the eastern side of Lake Mweru to the south. The fault dips to the west at 60° and is considered a thrust fault some strike-slip movement is suggested. There are three main copper deposits known to date Shaba, Safari North and Safari South which are all hosted along the Kapulo Fault with Banguelan Granite to the east and Katangan Sediments to the west. The Katangan sediments are comprised of three main supergroups, the Roan, the lower Kundelungu and the upper Kundelungu forming a 10km thick sedimentary pile. Sedimentation took place over three main periods which corresponds to these supergroups. Figure 7.1_1 shows the regional geology of the project area.
Figure 7.1_1 Regional Geology of Mawson’s convention area in the D.R.C
The mineralization is bounded by hanging wall shale and is hosted in coarse grained sandstones near the fault, and brecciated granite in the footwall part of the mineralisation The sediments have not been mapped in detail to put them in a stratigraphic context, but it has been postulated they may be Mwashya (upper Roan) at the base overlain by Lower Kundelungu sediments in the project area (Figure 7.1_2).
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Figure 7.1_2 Stratigraphy of Kapulo Region with known Styles of Mineralization m-Approx Super Group Group Formation Unit Lithology Mineralization thickness Plateaux Kilunga Lupili Pink Arkose <200 Ks-3 Arkose
Interbedded red argillaceous Kapenga Schists 350 sandstone and red argillates
Disseminated chalcocite and Sonta Sandstone Red sandstone and quartzite 100 malachite at base of Sonta Sandstone
Interbedded red argillaceous Sampwe schists 300 sandstone and red argillates Kiubo Ks - 2.2 Ks - 2 Kyafwama - Red sandstone 50 Kombo Sandstone
Interbedded red argillaceous Lufila schists 150 sandstone and red argillates
Kiubo Sandstone Red sandstone 50
Argillaceous red sandstones Fault controlled Cu-Ag Upper Kundelungu - Ks Ks - 1.3 Mongwa Schistsinterbedded with argillates. <100 mineralisation at the Minor Sandstone dolomite/ sandstone contact
Interbedded pink, cross Lubudi bedded dol-arenites,with olitic <50 Dolomites caps
Kanianga Sandstone - argillaceous and <50 Kalule Sandstone weakly calcareous Ks - 1 Kiaka Carbonates Ks - Pink intramicrite flakestones <50 1.2
Lusele Pink Interbedded white/pink Dolomites <50 carbonate muds and arenites
Zn-Pb mineralization at Basal dolomite (BD) 4 contact with diamictite Le Petit Ks - 1.1 Diamictite <100 Conglomerate
Stratiform copper Kinkumbi, Monwesi Monwesi Fluvio glacial sandstones <50m Lufukwe anticline Lower Le Grand Likasi Ki-1 Ki -0 1.1 Diamictite 500 Kundelungu Ki Conglomerate Mwashya Dolomitic shales and Roan R R - 4.2 ? R-4 sandstones
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A prominent sedimentary unit has been mapped within granite east of the Kapulo Fault. First recognised as greywacke during 2006 mapping, the unit appears to extend almost 25km in a north-south direction. East of Shimumba, the sedimentary unit comprises a thin basal arkose unit overlain by a thick succession of well-exposed polymictic conglomerate. A basal unconformity is well exposed and bedding dips gently to the east.
An inlier of mafic gneissic basement is exposed in the Kasama area NW of the Kapulo copper deposits. The gneiss contains more felsic zones and is cut by a SW trending shear zone which has introduced retrogressive chlorite and sericite alteration along the shear. The gneiss shows strong foliation along the shear zone and quartz veining has been emplaced along the shear zone and is prospective for gold, silver, copper, lead and zinc. The gneiss is overlain by upper Kundelungu sediments
7.2 Project Geology
The geology of the area surrounding the Kapulo deposits comprises the following key rock types:
Kundelungu Group Sediments including sandstones, mudstones, siltstones and quartzites.
Banguelan granite to the east
Polymictic breccias and mafic intrusions
The Banguelo Granite is a coarse grained biotite rapakivi textured monzogranite which is separated from the Kundelungu sediments to the west by the Kapulo Fault Zone. Where the fault is exposed in artisanal workings, it dips steeply to the west. The Kundelungu sediments are exposed as a ridge to the west of the Kapulo Fault. The dominant rock type is poorly bedded, sorted quartz-rich feldspathic sandstone. Minor occurrences of granite in the sediments suggest the granite sediment contact to be unconformable. A typical section through the Shaba deposit is presented below in Figure 7.2_1.
At Shaba, the copper mineralisation is controlled by a NNW fault that dips 55° to the west and sits at the contact between a black shale unit and a buff-brown hangingwall silicified sandstone unit. The hangingwall sandstone is not mineralised and surrounded by variably brecciated and epidote altered porphyritic granite.
The granite groundmass typically shows moderate penetrative deformation. Below the shale, the shale has acted as an aquatard during the mineralising event and no mineralisation has been noted above the shale in the hangingwall sandstone or granite. Copper mineralisation occurs as a band of cuprite on the footwall contact of the black shale with a broad zone of supergene malachite and azurite near surface. At depth, the copper lode is characterised by disseminated chalcopyrite (and minor chalcocite) with increasing bornite (and silver) as the lode plunges to the south. The felsic intrusive at Shaba can be mineralised and represent finer grained apophyses of the footwall alkali granite. A fine grained mafic intrusive also occurs at Shaba, which may be silica-altered lamprophyre or basalt, and is mineralised if intruded near the fault below the shale unit. The mafic unit is typically on the footwall contact of the greywacke.
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Figure 7.2_1 Typical Cross Section (Shaba Deposit)
At Safari (North and South), the copper mineralisation is hosted in similar structural position along the Kapulo Fault (as Shaba) in the coarse grained sediments with the sandstones and shale as a hanging wall to the mineralization, and the mineralisation hosted in the grey sandstone below the shale unit. The chalcopyrite/bornite mineralisation then continues into the brecciated footwall alkali granite. As the granite becomes less fractured and more brecciated, pyrite becomes the more dominant sulphide and the chalcopyrite decreases into the massive granite.
Figure 7.2_2 Local Geology of the Kapulo Area
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8 DEPOSIT TYPES
The Kapulo copper deposits are interpreted to be hypogene, fault controlled deposits, with disseminated chalcopyrite-bornite mineralisation in the sulfide zone and massive to nodular cuprite and supergene azurite and malachite near surface. The mineralization is litho- structurally controlled, and is hosted in shales and sandstones of the Kundelungu Group and in the footwall granite. Both the sediments and the granite show varying degrees of brecciation with the high grade zone having a bornite-chalcopyrite filled matrix. The upper parts of the deposits are hosted in the black shale and sandstone whilst the lower parts tend to be fault controlled granite breccias.
The Dikulushi copper-silver mine to the southwest of Kapulo, is interpreted to be a low temperature, hypogene vein of massive to semi massive sulphide. Both deposits are not typical of the copper deposits of the central African copperbelt.
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9 MINERALIZATION
The Kapulo fault has acted as a fluid conduit allowing deeper Cu-rich fluids to migrate up the fault until coming into contact with the black shale unit whereby the resulting redox reaction has precipitated the copper in the shale/sandstone units and the footwall granite. The granite shows variable brecciation associated with the mineralising fluids with breccia matrix of chalcopyrite and bornite mineralisation.
The Kapulo copper deposits contain both supergene or near surface and primary or sulphide mineralization. Secondary enrichment in the oxidised zone in the top 30m at the Kapulo Deposits has resulted in the deposition of malachite, azurite, cuprite and rarer chalcocite in the near surface parts of Shaba and Safari North. The malachite occurs in two forms with the most abundant being a light green disseminated malachite in the shale and sandstone on the footwall side of the fault and the lesser form as darker green bands of fracture fill, typically in the shales. The azurite occurs as nodules and layers in the shales and sandstones at both Shaba and Safari North. Cuprite mineralization “takes the form of massive and nodular cuprite located within the black shale unit and on its contacts, particularly the footwall contact” (Witt, 2006). Malachite, azurite and cuprite mineralization is easily observed in the artisanal workings at both Shaba and Safari North.
The main primary sulphide mineral at both Shaba and Safari North is chalcopyrite, with lesser amounts of bornite and chalcocite. The chalcopyrite occurs as fine grained disseminated sulphide and as more massive blebs up to 10mm. In some holes (08KTDH037) the chalcopyrite forms replacement layers in the sandstone. With increased depth (towards the east) in all holes chalcopyrite becomes less dominant and pyrite becomes the dominant sulphide. The mineral assemblages suggest that there is hypogene and supergene bornite and chalcocite, and that bornite and chalcocite of the two parageneses are often difficult to distinguish. Evidence that chalcopyrite mineralisation is early-stage, followed by bornite, then chalcocite, is quite common even in more probably hypogene assemblages. When an early-stage, usually weakly pyritic, probably carbonaceous or hydrocarbon-bearing assemblage (in either sandstone or brecciated granite) first reacts with the Cu-bearing fluid, the assemblage chalcopyrite-pyrite forms. Where more Cu becomes available, reducing species and S are still available from the carbonaceous matter/hydrocarbons, but Fe begins to run out. All pyrite is consumed, after which .there is no
longer enough Fe for chalcopyrite (CuFeS 2) to crystallise, and the Cu-bearing fluid reacts with
reducing, S-bearing hydrocarbons to form bornite (Cu 5FeS 4). Finally, there is not enough Fe to
form bornite and chalcocite (Cu 2S) becomes the precipitating Cu phase (England, 2008). Chalcocite is probably the main source of Ag (Haest et al., 2007).
Bornite does increase down plunge to the south at Shaba and native silver has been noted in the drill core. A higher grade core to the mineralized zone at Shaba plunges to the south at - 50 degrees and bornite is the dominant sulphide in this zone. Bornite is typical in the sandstone and footwall granite contact which is typically brecciated and occurs as breccia fill in this zone (Figure 9_1 below).
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Figure 9_1 Bornite Dominant Matrix in Granite Breccia – Shaba Deposit
The Kapulo Fault locally trends 343° magnetic but o verall trends close to north. This kink in the fault combined with sinistral movement combined with a degree of normal movement has provided an area of dilation for the fluid to be focussed and thus creating the high grade plunge to the mineralization as shown in Figure 9_2 for Shaba and Figure 9_3 for Safari North.
Gangue minerals typically associated with the mineralization at Kapulo are quartz, feldspar, pyrite and clays.
A list of relevant significant intercepts based on the metre sampling is presented below in Table 9_1.
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Table 9_1 Kapulo Copper Project
Kapulo Significant Intercepts (Relevant Samples)
Hole_ID mFrom mTo Width SampleType Cu_pct Ag_ppm 07SNDH001 40 50.6 10.6 DDH 5.2 54 07SNDH002 53 71 18 DDH 6.6 16 07SNDH003 86 92 6 DDH 3.8 7 07SNDH007 28 33 5 DDH 1.1 8 08SNDH001B 40 65 25 DDH 5.5 14 08SNDH001B 70 82 12 DDH 2.7 20 08SNDH008 34 47 13 DDH 3.5 8 08SNDH009 47 55 8 DDH 6.6 17 08SNDH009B 50 61 11 DDH 6.0 12 08SNDH010 78 90 12 DDH 5.9 22 08SNDH011 97 102 5 DDH 1.5 3 08SNDH011 113 120 7 DDH 1.2 15 08SNDH013 80 87 7 DDH 2.6 4 08SNDH014 91 99 8 DDH 3.7 4 08SNDH016 111 135 24 DDH 4.7 11 08SNDH017 135 141 6 DDH 4.3 13 08SNDH018 84 89 5 DDH 3.4 6 08SNDH019 40 45 5 DDH 4.6 9 08SNDH021 122 130 8 DDH 2.0 6 08SNDH022 97 108 11 DDH 2.1 4 08SNDH023 131 159 28 DDH 3.1 5 08SNDH024 147 162 15 DDH 2.3 6 08SNDH025 133 141 8 DDH 2.0 3 09SNDH026 0 22 22 DDH 7.5 17 09SNDH027 0 16.4 16.4 DDH 5.1 5 09SNDH028 0 18 18 DDH 3.9 2 09SNDH029 0 10 10 DDH 1.8 1 09SNDH031 0 13 13 DDH 6.3 16 09SNWE002 77 84 7 RC 3.0 4 07KTDH002 86 94 8 DDH 2.1 2 07KTDH002 96 101 5 DDH 1.0 1 07KTDH002 108 116 8 DDH 1.3 2 07KTDH003 106 113 7 DDH 2.0 2 07KTDH007 85 95 10 DDH 3.2 7 07KTDH007 108 115 7 DDH 4.1 5 07KTDH008 90.7 97.33 6.63 DDH 3.4 11 07KTDH008 101 108 7 DDH 1.3 3 07KTDH009 114.9 130.9 16 DDH 3.1 14 07KTDH010 83 111 28 DDH 4.0 6 07KTDH011 15 26 11 DDH 3.9 5 07KTDH012 24 29 5 DDH 3.0 11 07KTDH013 23 43 20 DDH 2.9 4 07KTDH014 26 41 15 DDH 3.0 9 07KTDH015 15 43 28 DDH 6.9 13 07KTDH015 57 66 9 DDH 1.4 1 07KTDH016 22 52 30 DDH 6.4 20 07KTDH016 73 81 8 DDH 1.1 1 07KTDH016 96 102 6 DDH 1.1 1 08KTDH017 87 116 29 DDH 6.4 12 08KTDH018 99 125 26 DDH 7.4 22 08KTDH019B 144 155.6 11.6 DDH 6.1 8 08KTDH021 140 149 9 DDH 5.1 2 08KTDH021 159 166 7 DDH 1.8 3
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Hole_ID mFrom mTo Width SampleType Cu_pct Ag_ppm 08KTDH021 188 198 10 DDH 1.1 1 08KTDH022 85 107 22 DDH 3.7 4 08KTDH023 137 177 40 DDH 8.5 27 08KTDH023 191 200.4 9.4 DDH 2.5 3 08KTDH023W1 138 177 39 DDH 7.8 21 08KTDH023W1 189 203 14 DDH 2.5 6 08KTDH024 96 123 27 DDH 8.3 23 08KTDH024 134 141 7 DDH 1.8 3 08KTDH025 102 124 22 DDH 6.2 11 08KTDH026 163 177 14 DDH 8.2 20 08KTDH027 43 76 33 DDH 5.7 7 08KTDH027 81 89 8 DDH 2.6 1 08KTDH028 38 68 30 DDH 4.3 8 08KTDH028B 38 68 30 DDH 5.9 13 08KTDH029 50 77 27 DDH 6.2 34 08KTDH030 188 221 33 DDH 6.8 26 08KTDH034 18 58 40 DDH 4.9 11 08KTDH035 144 176 32 DDH 4.3 14 08KTDH035 181 189 8 DDH 1.5 1 08KTDH036 178 200 22 DDH 8.0 22 08KTDH037 99 117 18 DDH 5.9 13 08KTDH039 39 60 21 DDH 1.6 3 08KTDH042 61 72 11 DDH 2.1 6 08KTDH042 97 103 6 DDH 1.6 13 08KTDH043 175 182 7 DDH 3.5 12 08KTDH044 346 355 9 DDH 1.7 7 08KTDH045 311 318 7 DDH 1.4 5 08KTDH046 241 266 25 DDH 5.8 43 08KTDH047B 172 177 5 DDH 7.2 12 08KTDH047B 184 207 23 DDH 6.3 16 08KTDH049 20 25 5 DDH 2.1 5 08KTDH050 133 152 19 DDH 3.6 14 08KTDH050 164 171 7 DDH 1.6 5 08KTDH051 109 140 31 DDH 3.9 8 08KTDH051 144 149 5 DDH 1.2 2 09KTDH052 7 15 8 DDH 2.9 3 09KTDH052 19 24 5 DDH 1.2 2 09KTDH053 9 18 9 DDH 1.6 1 09SBDH054 14 19 5 DDH 1.5 1 09SBDH055 8 20 12 DDH 4.6 2 09SBDH056 0 24 24 DDH 10.9 3 09SBDH060 18 41 23 DDH 3.4 3 09SBWE001 129 147 18 RC 7.7 26 09SBWE001 151 158 7 RC 1.6 3 09SBWE003 88 119 31 RC 6.0 10 09SBWE004A 38 71 33 RC 8.3 14 09SBWE004B 36 68 32 RC 6.6 11 09SBWE005 144 189 45 RC 3.5 6
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Figure 9_2 Shaba Grade-Thickness Projection
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Figure 9_3 Safari North Grade-Thickness Projection
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10 EXPLORATION
The Project has been explored by three main groups over the last 100 years. The three main periods of exploration are:
Belgian Exploration (1910-1954)
Falconbridge Exploration (1970-1974)
Mawson West Exploration (2006 to present)
10.1 Belgian Exploration
The Shaba and Safari deposits were discovered by Tanganyika Concession Ltd in 1906 and then transferred to BAKAT (Mining Research Company of Lower Katanga) from 1911 to 1914 and were subsequently taken over by SIMIKAT in 1915. The work was completed by the Société de Recherche Minière and included several small trenches and a 20m gallery.
In 1915 SIMKAT drove 23 adits into Shaba and 33 into the Safari deposits. All the adits and shafts were above the water table and covered a vertical distance of approximately 50m. This work include 328m of drilling, 386m of shafts, 924m of trenching and 4,168m of adits.
In 1925 Simkat completed verification sampling in adits 8, 9 13 and 14 to check the previously completed work.
In 1956-57 Sermikat re-opened an unknown adit and produced 3 tonnes of sampling for metallurgical testwork. The sample was taken from three material types, namely direct ship, oxide and sulphide, with the sample returning 5.37% Copper of which 2.03% was oxide copper.
SERMIKAT also completed an unpublished historical grade tonnage estimate for all the Kapulo deposits which resulted in a total of 40,252t of copper based on 1.1Mt @ 6% Cu using a 3% cutoff to a depth of 60m (27,734t at Shaba, 11,318t at Safari North and 1,200t at Safari South). The company noted the deposits were all open at depth and along strike. Historical information is presented for informational purposes only. The resource estimate was not completed in accordance with National Instrument 43-101 and therefore should not be relied upon.
During the period 1932 to 1959 geological studies were pursued and a reassessment of the deposits using a lower copper cutoff grade of 1% Cu resulted in a contained metal tonnage of 53,000t of copper.
10.2 Falconbridge-Seremi Exploration
During the early 1970’s Falconbridge explored the copper prospects with regional soil geochemistry, photo geological mapping and completed two DDH into Shaba for a total advance of 176m.
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In 1970, Falconbridge Exploration Limited obtained an interest in the property and completed the following exploration during 1971 to 1974, in conjunction with Seremi:
Seven adits were re-opened in the Shaba deposit and channel sampled for a total of seven bulk samples which ranged from 10m to 30m in channel length and the samples were sent for metallurgical testwork and beneficiation testing. Check assaying of the above adits showed a slight decrease on previously reported Falconbridge and SIMIKAT results.
Falconbridge drilled two diamond holes for a total of 176m of which one intersected the main ore zone with a grade of 15m @ 4% Cu from a vertical depth of 125m.
A further seven drillholes were completed during 1974 by SEREMI for an advance of 1703m using a Japanese EP1 type drilling rig by Nikko Exploration and Development Company a branch of Nippon Mining. All drillholes were vertical and hit the mineralization around 200m with minimal Cu grades, and in some cases missed the mineralization altogether. Sampling was generally on a 1m basis but could extend up to 1.5m.
Completed 44 thin sections from drill core and surface samples and 25 polished sections from the drill core
Completed soil geochemical sampling on a 30m by 100m grid over the strike extent of the Kapulo fault in the immediate Shaba – Safari area for 14.4km.
Photo geological interpretation was completed at 1:40,000 and reconnaissance geological mapping at 1:20,000 scales were completed. Topographic mapping and RL checks with a Thommen altimeter.
Completed a grid base Misse a la Masse and Self Potential (SP) ground geophysical surveys which confirmed the known extent of the mineralized zones.
Prepared reports on the above work
10.3 Mawson West Exploration
Mawson West has completed the following work on the Project.
Mapping and reconnaissance sampling completed by consultant Walter Witt in order to provide information on mineralisation controls to assist with drillhole planning and to assess the suitability of the area for soil geochemistry. The report provided a number of suggestions for future exploration work such as collection of geophysical data and follow up geochemical work to assist in defining additional exploration targets.
A low-level airborne geophysical survey was completed in 2007 by UTS and the raw data subsequently reprocessed by Southern Geoscience Consultants in 2007. Several magnetic images including Total Magnetic Intensity (TMI), and 1 st and 2 nd vertical derivatives were produced and interpreted to improve the understanding of the structural framework of the area and specifically the Kapulo deposits.
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Mawson has completed a total of 204 drillholes for a total of 21,586m of drilling at the Kapulo Project including 16,408 m of NQ/HQ, and 5,178m of RC between 2007 and 2010. At the Shaba Prospect 93 drillholes for 13,623m have been completed and at Safari a total of 80 drillholes for 5,957m have been completed. To date drilling has defined the mineralization to the north and a high grade south plunging copper shoot. The mineralization remains open to the south and at depth at both Shaba and Safari North and South.
Regional soil geochemical sampling program was designed and supervised by Stan Harrison, a consultant geochemist, based in Perth, Western Australia. A total of 3,373 soil geochemical samples were collected over the northern and central regions of Kapulo with nominally 50m spaced samples on 400m spaced lines. The samples were sieved using 80# mesh during collection and bagged into numbered geochemical sample bags. Samples were then sent to Genalysis in Johannesburg for sample preparation and then submitted to Genalysis Laboratories in Perth for multi-element analysis by ICP-MS for Au, Ag, As, Bi, Mo, Sb, U, W and ICP_OES for Ba, Cu, Ni, Pb and Zn. Results highlighted several high order coincident multi-element geochemical anomalies along the Kapulo fault zone.
CSAMT ground geophysics survey to test the Kapulo Fault, located north and south of the known deposits. A number of targets have been identified from this work.
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11 DRILLING
11.1 Introduction
Prior to the JV agreement between Mawson West and Anvil Mining Ltd several small phases of diamond drilling were completed at the Kapulo deposits by several companies described above in the previous 60 years. As the majority of this work was completed prior to 1974 none of the data has been available for any of the prior drilling and no check assaying was able to be completed and as such none of the data has been used in the current resource estimate.
The Simikat and Falconbridge drilling was not supplied and is not included in the resource estimate that is described in this report.
11.2 Mawson West Drilling
Mawson West has completed a total of 204 drillholes for a total of 21,586m of drilling at the Project including 16,408 of NQ/HQ and 5,178m of RC between 2007 and 2010. In 2007 the drilling was completed by Drill Africa who was replaced in late 2007 by Chantete Drilling of Zambia. The objective of this work was to define the extent and tenor of the mineralization at the Project and to define an indicated and Inferred Resource. The average diamond core recovery at the Shaba Prospect is 86.4%. A summary of the drilling completed at the Project is presented in Table 11.2_1 below.
Table 11.2_1 Kapulo Copper Project Drilling Summary
Company Prospect Period Type No. of Holes Metes SIMIKAT Katanga 1910-1957 ? ? 328 Falconbridge Katanga 1973-1974 PQ/HQ 7 1,880 2007 NQ/HQ 17 2,286 2008 NQ/HQ 40 7,066 Shaba RC 7 992 2009 NQ/HQ 15 1,481 2010 NQ/HQ 14 1798 2007 NQ/HQ 6 508 Mawson West 2008 NQ/HQ 20 2,656 2009 NQ/HQ 16 523 Safari RC 13 855 RC 24 1,325 2010 NQ/HQ 1 90 Shaba Expl. 2009 RC 31 1,376 Total 211 + 23,794 Note: Figures have been rounded
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11.3 Survey Control
Original Survey Control
Survey control was established at the Project by Photomap Int surveyors from Nairobi, Kenya. A total of 10 control points were added using a MAGELLAN GPS PROMARK X-CM unit with processing completed in M-STAR Professional GPS software. The survey was based on the established control station CAR1. GPS fixes were carried out over several days from 14 th - 18 th August 2007 at CAR1 to obtain coordinates in the UTM Zone 35S, WGS84 system. The height is expressed as above the ellipsoid. Coordinates for CAR1 are shown below:
CAR1 9,084,000.871 745,291.374 1,046.186m
Control points were surveyed in UTM using the WGS84 datum and converted to local grid via a two point transformation in MicroMine software. The traverse was measured by GPS method. A stationary differential method was employed where one receiver was placed on a known station and others on remote stations. The mask angle was set to 15°. On each survey station, raw data was generally collected for 60-90 minutes at one-second interval. Raw survey data was processed with the M-STAR professional GPS software. All heights are based on point CAR1 and were carried forward by double run spirit levelling to connect to control points thus obtaining a good overall height reference in the network.
Once the control points were established a 50m by 50m grid was marked out using compass and tape and optical squares to provide a grid over each of the main prospects at the Project. Drillhole collars were then marked out using tape and compass using the 50m spaced grid.
2010 Survey Control
During 2010 surveyors from Cardno Spectrum (Cardno) in Perth, Western Australia spent approximately six weeks completing a site topographical survey using an RTK GPS to produce a DTM of the entire Project area and surrounds accurate to +-30mm. The survey control data was post processed using Topcon tools software. The digital terrain model (DTM) was generated in Surpac software and provided as a 3D DTM wireframe.
During this process the surveyor checked the WGS 3D co-ordinates at Kapulo, geodetically and found the previous surveys had been completed to height above ellipsoid rather than mean sea level. In the horizontal, the difference from true UTM is small (around 0.8m). The vertical difference however, is out by about 16 metres from the current “above EGM96 geoid” height. This is the elevation which best approximates mean sea level (MSL).
To obtain elevations approximating MSL a variable was applied (N separation), which is the difference between the earth’s ellipse and it’s gravimetric geoidal model. At Kapulo, the N separation is approximately 11.6 metres. Therefore, there is an offset of around 16 metres between the Kapulo heights and MSL. The DTM model produced by Cardno was accurate to 30mm for each spot height collected. All the Kapulo data was then subsequently transformed to the new coordinates using the shift below given by Cardno.
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The block shift is:
Northing +0.89m
Easting -0.713m
RL +16.590m
Check coordinates at CAR4 are:
Original: 9081475.602mN 745926.686mE 1155.009mRL
New 9081476.492mN 745925.973mE 1171.601mRL.
11.4 Collar Pickups
Drill collars have been marked out using a Geodimeter 620 total station and then the final completed collar position surveyed in LMG and converted to UTM using CAR1 for a simple two point transformation in MicroMine software.
11.5 Topography and Pit Surveys
During 2010 surveyors from Cardno Spectrum (Cardno) in Perth, Western Australia spent approximately six weeks completing a site topographical survey using RTK GPS to produce a DTM of the entire Kapulo Project area and surrounds accurate to ±30mm. The survey control data was post processed using Topcon tools software. The DTM was generated in Surpac software and provided as a 3D DTM wireframe.
11.6 Drilling Orientation
Down hole surveys of drillholes were completed at 30m downhole intervals. The surveys were measured using an Eastman downhole electronic single-shot camera which records the dip and azimuth to a chip and the results are then tabulated for each hole on a report card which was checked by the site geologists for accuracy. The camera has a stated accuracy of 1° in dip and azimuth. No magnetic minerals have b een noted in the logging of the drill core at the Project and thus the recorded azimuths are considered to be correct.
11.7 Oxidation
All drillholes were logged for oxidation and thus an oxidation boundary wireframe surface was constructed from these oxidation points in the drillholes and by using the core photography as a guide. The oxidation boundary was then used to code the block model for oxidation. At Kapulo, the surface of the oxide resource has been depleted by the illegal artisanal mining.. The artisanal mining was stopped in June 2007.
11.8 Relationship Between Sample Length and Mineralised Thickness
There is no relationship between the sample length and the true thickness of the mineralization. The mineralization encountered to date in the drilling is significantly thicker than the average sample length.
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12 SAMPLING METHOD AND APPROACH
The sampling procedures described below have been used for all diamond core drilled by Mawson West at the Project.
12.1 Soil Sampling Procedures and Sample Transportation
Soil sampling at the Project was completed under the direction of Stan Harrison, an independent geochemistry consultant from Perth, Western Australia. Soil samples were taken in areas of residual soils (rather than transported). Approximately 1kg of sample were sieved on the sample site using -80mesh sieves to collect approximately 100g of sample. Sieves were brushed and knocked clean between samples to avoid cross sample contamination. The sieved fraction was then placed into numbered paper geochemistry bags. The geochemical bags were then packaged in boxes and a submission form completed by the supervising geologist. The samples were then driven to the Zambian border for customs clearance and then on to Lusaka, Zambia. Once in Lusaka the sample boxes were cleared with Zambian customs and then transported by Pine Transport of Zambia in a locked weatherproof truck under customs seal to Genalysis Laboratories in Johannesburg, South Africa. Sample preparation was completed at Genalysis in Johannesburg including drying and pulverizing to 85% passing -75µm. The sample pulps were split to approximately 20gr and sent by commercial airfreight to Genalysis Laboratories in Perth, Western Australia. Genalysis Perth has customs clearance directly for prepared samples from outside of Australia.
12.2 Diamond Core Sampling
Drill core was cut in half using a water lubricated diamond blade saw which was cleaned frequently using a brick to prevent across metre contamination (especially in zones of massive sulphide). The half core was then retained for sampling. All drill core was sampled at the Project on a metre basis from the start to the end of a drillhole. Each metre sample of half drill core was collected from the southern (or western half) of the drill core tray and placed into a sequentially numbered calico bag. Sample log books were also used to record the drillhole number and sample number and the ‘from to’ sample depth as a backup to entering the data into a “toughbook” laptop computer.
Calico bags were then tied up and placed into larger plastic bags for transport which were labelled with sample number, from and to and an id number for the large bag. Submission forms for the laboratory were completed and put into a small plastic bag inside one of the large bags containing the calicos.
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12.3 Diamond Core Recovery
At the end of a core run, the drillers attached a water hose and pumped the HQ/NQ core out of the barrel into an angle iron to disturb the core as least as possible and the driller then wrote out a core block with the total depth and core run length, also noting any core loss and placed this into the core tray at the end of the run. The site geologist regularly checked the depths provided by the drillers with the core in the trays during rig visits. Any handling core breaks were marked with a cross. Once a core tray was full, the tray was labelled with from, to depths and hole number and moved to the core logging/storage area of the Kapulo campsite.
12.4 Diamond Core Logging
The drill core was washed to remove any residual cuttings, metre marks down hole were then made by the geologist and written on the northern (or eastern) half of the core. The core was then made wet and photographed using a digital camera before cutting. Labels showing hole number, tray number and from and to depths were placed in the photo frame for each core tray during photography. Core recovery, RQD, geology, alteration and mineralization were logged into a “toughbook” laptop. The files were then emailed to Mawson’s Perth office and loaded into a central database. The above orientation line mark was used as a guide for cutting the core in half and across the core at the metre marks as all core was sampled in one metre samples except for the end of hole where a longer sample may be collected, i.e. if the hole ends at 104.4m the last sample would be 0.4m long. Likewise if the drillhole ends at 108.8m the last sample would be 0.8m in length.
12.5 RC Drilling
RC drilling was completed by Titan drilling of Lubumbashi (DRC) during late 2009 and 2010 for the groundwater investigation holes and Safari North resource definition holes using a Samil truck mounted RC rig with 900cfm and 350psi compressor.
12.6 RC Sampling and Logging
RC samples were routinely collected at 1m intervals into a cyclone with an RC type green UV degradeable plastic bag to catch the cuttings. The cuttings were then split with a Jones riffle 3 tier splitter, which splits 12.5% to 87.5%. The calico was placed on the front of the splitter and collects the 12.5% as a split sample and the residue went back into the green UV plastic bag. Field duplicates were taken via the splitter every 20 samples with the sample split twice to provide a field duplicate. The bags of cuttings were routinely weighed prior to taking the sub-sample via the Jones riffle splitter. The calico bags were then weighed and put into a labelled polywovens bag, which was sealed with a cable tie. The samples were taken back to camp and stored in a dry area until collection.
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The quality control procedures are listed below:
Use of face sampling hammers.
Routine blowbacks (i.e. each metre is cleared into the sample bag prior to moving on to the next metre of advance).
Routine weighing of collected RC cuttings.
The cyclone is cleaned at each rod change.
On a rod change any material in the hole is cleared before the first new metre sample is collected.
The riffle splitter is cleaned with brushes between each sample.
Every 20m a duplicate field sample is collected by riffle splitting the bag of cuttings once more.
On the extremely rare occasions that moisture is encountered in the hole (please note the emphasis on moisture rather than wet drilling conditions), the hole was ‘conditioned’ or dried prior to advancing further.
Additional compressed air boosters were routinely used in order to ensure dry samples below the water table.
12.7 Sample Quality
The sampling procedures adopted for drilling are consistent with current industry best practise. Samples collected by diamond coring within the highly weathered zones are of moderate quality, with the remainder being high. Sample recoveries and quality for the RC drilling are high with no wet drilling encountered for the RC.
Dedicated drillhole twinning of the DC drilling and RC drilling was completed by Mawson West and is discussed in Section 14.1.2. The twin analysis did not show any bias from the DDH to the RC drilling.
RC field duplicate samples were routinely collected every 20m of RC to allow assessment of the field sampling error (or bias) once the laboratory error, determined from analysis of pulp duplicates, was subtracted. Acceptable reproducibility was identified during an assessment of RC field duplicate data (Section 14.2) generated and no distinct bias was evident
Diamond core sampling was done by cutting the core in half and always sampling the same side of the core in the tray. The core cutting machine was cleaned between each metre to reduce any possibility of contamination. Metre samples were then bagged into numbered calico bags which, in turn, were bagged into polywoven bags and sealed with cable ties for transport to the laboratory.
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13 SAMPLE PREPARATION, ANALYSES AND SECURITY
13.1 Sample Security
The close scrutiny of sample submission procedures by Mawson technical staff, and the rapid submission of samples from drilling for analysis, provides little opportunity for sample tampering. Equally, given the umpire assaying via an external international laboratory and the regular ‘blind’ submission of international standards to both the primary and umpire assay facilities, any misleading analytical data would be readily recognised and investigated.
Current Kapulo sampling procedures require samples to be collected in staple closed bags once taken from the rig. They are then transported to the Kapulo camp to be picked up by the laboratory truck. The laboratory truck then takes them to the laboratory directly.
Reference material is retained and stored at the Kapulo exploration camp, as well as chips derived from RC drilling, half-core and photographs generated by Diamond drilling, and duplicate pulps and residues of all submitted samples. Assessment of the data indicates that the assay results are generally consistent with the logged alteration and mineralisation, and are entirely consistent with the anticipated tenor of mineralisation which is typically visible in the core photography.
13.2 Analytical Laboratories
Preparation and assaying of samples from the Kapulo Project has been carried out at two independent laboratories:
Genalysis (RSA & Australia) (from Jan 2007).
ALS-Chemex (RSA) (from Jan 2008).
During 2007 the core samples from Kapulo were shipped to Genalysis Laboratories in Johannesburg, South Africa for sample preparation and a split of the pulp sent to Genalysis, Perth, Western Australia for analysis. During 2008 all diamond core samples from the Kapulo Copper Project have been shipped to ALS-Chemex Laboratories in Johannesburg, South Africa for analysis.
13.3 Sample Preparation and Analytical Procedure
13.3.1 Drillcore Analyses
Samples sent to Genalysis Johannesburg, South Africa were processed and analysed using the following methods.
All samples were weighed then dried at 110° for eig ht hours and then crushed to a nominal 10mm in a conventional jaw crusher. The entire sample was then pulverised to a nominal 85% passing -75µm in an LM-5 mixer-mill. A scoop of the sample was then digested by the AX method, which is a modified (for higher precision) four acid digestion (for base metals and analysed by AAS for Copper (0.01% DL) and Inductively Coupled Plasma Mass Spectrometry (ICP-MS) for Ag (1ppm DL), As (10ppm DL), Co (1ppm DL) and U (0.1ppm DL). Results were reported via email and a hard copy report was mailed to the Perth office.
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Diamond core samples sent to ALS Chemex in Johannesburg, South Africa were prepared and analysed by the following procedures.
Samples were weighed and then dried for 8 hours at 110° for eight hours and then fine crushed to 2mm with a 250gram split of the sample taken for pulverising to 85% passing - 75µm. The sample was then digested in a four acid mixture (HF, HNO3, HCLO4) and a HCL leach with analysis by ICP-AES method to determine which samples contained greater than 1% Cu or 100g/t Ag and these samples were then analysed by the AAS (method AA62) for Copper (1% to 40% reporting range). Other elements were analysed by the ICP-AES (optical emission method) with detection limits in brackets, Ag (0.5ppm), As (5ppm), Co (1ppm), U (10ppm) and Co (1ppm).
13.3.2 RC Sampling and Analyses
Samples from the RC drilling were analysed by ALS Chemex in Johannesburg using the same sample preparation and analytical procedure described above for the diamond core.
13.3.3 Soil Sampling Analyses
Soil samples were all submitted to Genalysis for multi-element analysis at their Perth laboratory. As the samples were sieved during collection the charge was digested directly using the Aqua Regia digest method and a suite of 14 elements were analysed using the ICP-MS or ICP_OES methods. Analysed elements, analytical method and detection limits are presented in Table 13.3.3_1 below.
Table 13.3.3_1 Kapulo Copper Project Analytical Summary Soil Samples
Element Au Ag As Ba Bi Co Cu Mo Ni Pb Sb U W Zn Method MS MS MS OES MS OES OES MS OES OES MS MS MS OES D.L.(ppm) 1 0.01 0.5 2 0.01 1 1 0.1 1 2 0.02 0.01 0.05 Q1
13.4 Bulk Density Determinations
Samples were collected from a representative selection from the transitional and primary zones and from un-mineralised areas. Diamond core samples were prepared by ‘squaring off’ the ends of approximately 10cm to 20cm billets of half core. A total of 1,294 bulk density (BD) measurements were made of dried half core to obtain the dry weight at Shaba. The same piece of core was then measured in water on a suspension cage below the same electronic scale.
BD = Dry Sample Weight / (Dry Sample Weight – Wet Sample Weight)
The average BD measurements and statistics for Shaba and Safari for mineralised and un- mineralised core are described in Section 17.
13.5 Adequacy of Procedures
Analytical procedures associated with data generated to date are consistent with current industry practise and are considered acceptable for the style of mineralisation identified at Kapulo. Quality control procedures are described in Section 14.
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14 DATA VERIFICATION
Mawson West routinely submits internationally recognised standards at a rate of 1:30 samples with all samples submitted to ALS Johannesburg. For RC samples a field duplicate is taken every 20 th sample. Up to 5% of sample pulps are submitted to another independent laboratory for umpire analysis. If a standard value is more than 2 standard deviations away from the excepted value the lab is alerted and re-assays the entire submission associated with that standard. No independent samples were taken by Steve Le Brun during his site visit. The outcropping mineralisation at both Shaba and Safari north is visual in nature.
14.1 Standards and Blanks – Mawson West
The performance of the reference standards analysis is good with all standards falling inside the ± 2 standard deviations of the expected value. In addition, the laboratories submit their own reference standards into the assay stream and this data is presented below in Table 14.1_1.
Table 14.1_1 Kapulo Copper Project Mean Values of Geostats Certified Reference Materials Standards
Standard Element Value Confidence Interval Cu 72,921ppm ±710.8 GBM303-2 Co 641ppm ±20.3 Ag 26.1ppm ±0.7 Cu 1,138ppm ±13.3 GBM305-4 Co 70ppm ±3.2 Ag 2.6ppm ±0.1 Cu 454ppm ±7.1 GBM398-3 Co 15ppm ±1.2 Ag 1.1ppm ±0.19 Cu 16,603ppm ±157.3 GBM900-3 Co 150ppm ±3.4 Ag 7.4ppm ±0.25 Cu 26,295ppm ±250 GBM902-5 Co 79ppm ±2.7 Ag 15.7ppm ±0.44 Cu 247ppm ±4.5 GBM906-1 Co 30ppm ±2.1 Ag 22.6ppm ±0.28 Cu 74ppm ±1.4 GBM908-1 Co 34ppm ±2.3 Ag 1.7ppm ±0.1 Other elements for each standard not shown
Certified reference material “standards” were inserted into the sample stream at a rate of 1 standard per 30 samples submitted. Standards were purchased from Geostats in O’Connor, Western Australia. Standards were also submitted with the rock chip and soil geochemistry samples with varying grades of nickel, copper, zinc, lead, arsenic, cobalt and silver.
All data gathered for quality control samples – blanks, duplicates and reference materials were automatically captured, sorted and retained in the QC Database.
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A total of 203 standards were submitted for Shaba and Safari (138 and 65 respectively). No systematic bias or significa nt issues have been observed in the QAQC results, minor deviation outside the expected values have been noted, but the majority of these are not significant.
Table 14.1_2 Kapulo Copper Project Certified Standards Summary
Project Area Hole Type Count Safari North Channel CH 6 Safari North Drill DDH 32 Safari South Drill DDH 3 Safari South Drill RC 24 Shaba Channel CH 38 Shaba Drill DDH 75 Shaba Drill RC 25 Total 203
All samples are considered to be representative of their respective interval and no bias has been introduced by selective sampling. Sample length was not based on width of mineralised interval, rock type or alteration, but is a standard one metre length. Sample lengths vary at th e end of holes but this is a drilling factor rather than a sampling factor. There are no other factors which could introduce significant bias into the sampling of the diamond core.
14.1.1 Umpire Assay Checks
A total of 222 pulps ; 169 for Shaba and 53 for Safari North, were split and sent to Genalysis as umpire assay checks. The performance of the umpire assay checks is shown in Figure 14.1.1_1 below. For both the copper and the silver assays, both comparisons fall within acceptable limits.
Figure 14.1.1_1 ALS vs. Genalysis; Cu% & Ag ppm Umpire Assays Comparisons
14.1.2 Drillhole Twinning
No specific twin holes have been drilled on the Kapulo deposits to date. There are none at Safari, there is one diamond hole pair, one RC hole pair and five RC -diamond hole pairs that are sufficiently close to each other within the mineralised zones at Shaba; within ±5m (up to 25m), to be used as an indication of twin hole representivity. Figure 14.1.2_1 shows a twin hol e comparison between a RC drill hole and a DD drillhole at Shaba.
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Figure 14.1.2_1 Twin Hole Comparison RC versus DD Drillhole at Shaba
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Table 14.1.2_1 provides a summary of the drillholes used for the twin hole comparison.
Table 14.1.2_1 Kapulo Copper Project List of Drillholes Used For Twin Hole Comparisons
Pair Type Hole 1 Hole 2 Separation DD-DD 08KTDH028B 08KTDH029 3m RC-DD 09SBWE003 08KTDH017 6m RC-DD 09SBWE003 08KTDH010 22m RC-DD 09SBWE003 08KTDH025 25m RC-DD 09SBWE005 08KTDH019B 4m RC-RC 09SBWE004A 09SBWE004B 5m
In all cases, the trace and statistical comparisons indicate that there is good correlation between the drillholes.
14.1.3 Laboratory Blanks and Standards
The laboratory performs internal QAQC checks to the following description form ALS Chemex:
“The Laboratory Information Management System (LIMS) inserts quality control samples (reference materials, blanks and duplicates) on each analytical run, based on the rack sizes associated with the method. The rack size is the number of sample including QC samples included in a batch. The blank is inserted at the beginning, standards are inserted at random intervals, and duplicates are analysed at the end of the batch. Quality control samples are inserted based on the following basis:”
Table 14.1.3_1 Kapulo Copper Project ALS Chemex Internal QAQC Checks
Sample Count Methods QAQC Sample Allocation 40 Regular AAS, ICP-AES and ICP-MS methods 2 standards, 1 duplicate, 1 blank
The laboratory staff analyses quality control samples at least at the frequency specified above. If necessary, laboratory staff may include additional quality control samples above the minimum specifications. Example QAQC standards charts are shown in Figures 14.1.3_1 and 14.1.3_2 below.
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Figure 14.1.3_1 Shaba Standard GBM303-2 MAWSON STANDARDS (ALS_JB - GBM303-2 - Cu)
Standard: GBM303-2 No of Analyses: 98 Element: Cu Minimum: 59,100.000 Units: ppm Maximum: 81,400.000 Detection Limit: Mean: 70,242.857 Expected Value (EV): 72,921.000 Std Deviation: 3,549.447 67,227.000 to E.V. Range: 78,615.000 % in Tolerance 79.592 % % Bias -3.673 % % RSD 5.053 %
Standard Control Plot (ALS_JB - GBM303-2 - Cu)
90000
80000
70000
value(PPM) 60000
50000 0-Apr-2 00810 -Aug-20 0822 -Oct-20 0813 29-Dec-2008 -Feb-200 912 009 -Mar-2 13 -Jun- 200929 28-Sep-2009 11-Jan-2010
DATE
value Expected Value = 72,921.000 EV Range (67,227.000 to 78,615.000) Mean of value = 70,242.857
Figure 14.1.3_2 Safari North Standard GBM3093-16
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14.2 Duplicates
No duplicates were completed on the diamond core samples. A total of 162 field duplicates (54 at Shaba and 108 at Safari) were submitted from the RC drilling and channel sampling, see Table 14.2_1, Figure 14.2_1 and Figure 14.2_2 below.
Table 14.2_1 Kapulo Copper Project Field Duplicates Summary
Project Area Hole Type Count Safari North Channel CH 22 Safari North Drill RC 3 Safari South Drill RC 33 Shaba Channel CH 38 Shaba Drill RC 16 Total 112
Figure 14.2_1 Shaba Original vs. Field Duplicates
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Figure 14.2_2 Safari North Original vs. Field Duplicates
14.3 Data Quality Summary
No significant issues have been observed in the QAQC results for the Shaba and Safari deposits. Any errors in standards results tend to be under-reporting results.
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15 ADJACENT PROPERTIES
The tenements to the north of Kapulo are adjacent but only have greenfields work completed on them at present, with no defined resources determined to date.
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16 MINERAL PROCESSING AND METALLURGICAL TESTING
16.1 Introduction
Mawson West is planning to develop the Project utilising a process plant with a rated capacity of 0.4Mtpa for oxide material and 0.5Mtpa for sulphide material.
The metallurgical testwork upon which the process design flowsheet was based is described in the following sections.
16.2 Metallurgical Samples
Three separate suites of samples were submitted for metallurgical testwork. Samples from 31 drillholes, from the Shaba sulphide ore zone, formed the first suite of samples, with 48 samples from the Shaba oxide ore zone and 37 samples from the Safari North oxide ore zone forming the second suite of samples and samples from nine drillholes, from the Shaba sulphide ore zone, forming the third suite of samples. No samples have been supplied from the Safari North sulphide ore zone. A total of nine composites were prepared for the testwork.
A summary of the composites, the head assays and the testwork programs is given in Table 16.2_1.
Table 16.2_1 Kapulo Copper Project Samples Submitted for Metallurgical Testwork
Testwork Head Assays Composite Ore Type Comments Program % Cu Ag g/t % S Master Sulphide 2847 4.37 9.4 4.36 Comminution, size by size, HLS & flotation Chalcopyrite Sulphide 2847 2.82 2.0 2.40 Flotation- 1 test only Bornite Sulphide 2847 7.92 22.0 6.22 Flotation- 1 test only Shaba Oxide 3035 5.41 11.0 0.86 Comminution, flotation, thickening &filtration Safari North Oxide 3035 5.39 ------Comminution & flotation Composite 1 Sulphide 3125 3.88 8.0 3.19 Comminution & flotation Composite 2 Sulphide 3125 4.31 8.8 3.67 Comminution & flotation Composite 3 Sulphide 3125 5.21 11.5 6.00 Comminution & flotation Blend Sulphide 3125 4.47 9.4 4.29 Flotation, thickening & filtration
Composites 1, Composite 2 and Composite 3 were prepared from samples so that the composites represented the projected plant feed for Year 1, Year 2 and Years 3 to 5 respectively.
The blend composite was made up of equal masses from Composite 1, Composite 2 and Composite 3. The head assays for the chalcopyrite, bornite and blend composites were calculated head assays from the testwork.
Figure 16.2_1 shows a section view of the locations of the metallurgical master composite location.
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Figure 16.2_1 Section View Showing Metallurgical Master Composite Location
16.2.1 Head Assays and Mineralogy
The head assays and mineralogy of the Kapulo metallurgical composites are summarised in Table 16.2.1_1. No specific mineralogy has been undertaken at this stage, however, the sequential copper assays give a good guide to the copper mineral species present in the mineralisation. Information from Mawson West geologists indicate that the copper minerals present in the Kapulo ore bodies are predominantly malachite, bornite and chalcopyrite with only minor or rare occurrences of other copper species.
Table 16.2.1_1 Kapulo Copper Project Head Grades
Shaba Sulphide Oxide Shaba Sulphide Analyte Units Master Shaba Safari North Comp 1 Comp 2 Comp 3
Cu acid soluble (malachite) 0.18 5.67 --- 0.20 0.12 0.12 %
Cu CN soluble (bornite) 1.97 0.37 --- 2.35 2.21 1.76 %
Cu residual (chalcopyrite) 2.24 0.06 --- 1.40 1.78 3.39 %
Cu sequential total 4.39 6.10 --- 3.96 4.11 5.27 %
Cu Total 4.37 5.41 5.39 3.88 4.31 5.21 % Ag 9 11 --- 8 9 12 ppm Fe 4.08 1.00 --- 2.53 2.97 5.81 %
S Total 4.36 0.86 --- 3.19 3.67 6.00 %
S Sulphide 3.90 ------%
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The sequential copper assays confirm that the Shaba oxide mineralisation was predominantly malachite with minor amounts of bornite and chalcopyrite. No sequential copper assay was conducted on the Safari North oxide sample as it has been assumed that would be similar to the Shaba oxide ore.
The sulphide composites were composed predominantly of bornite and chalcopyrite.
16.2.2 Comminution
The comminution testwork is summarised in Table 16.2.2_1.
The results show that the oxide mineralised material had low impact crushing work indices, low Bond rod mill work index and moderately high Bond ball mill work index.
The sulphide mineralised material had moderate Bond rod mill work indices and moderately high Bond ball mill work indices.
The abrasion indices were variable from very low to moderately high for the oxide mineralised material and from moderately high to very high for the sulphide mineralised material. The high abrasion indices will impact on operating costs due to high crushing circuit wear rates and high grinding media consumption rates.
Table 16.2.2_1 Kapulo Copper Project Comminution Test Results
Shaba Sulphide Shaba Safari North Shaba Sulphide Parameter Units Master Comp. Oxide Oxide Comp 1 Comp 2 Comp 3 Impact Work index --- 4.7 2.6 ------kWh/t 0.0337 Bond Abrasion Index 0.4108 0.2874 0.0360 0.3980 0.5723 0.2807 g 0.0107 Bond Rod Mill Work Index F80 10,157 5,062 --- 10,114 10,127 9,735 microns P80 791 703 --- 750 771 874 microns Work Index 13.1 6.8 --- 12.4 12.5 12.1 kWh/tonne Bond Ball Mill Work Index F80 2,360 1,417 --- 2,136 2,586 2,549 microns P80 84 87 --- 88 86 85 microns Work Index 17.4 16.9 --- 13.3 15.2 16.1 kWh/tonne
16.2.3 Oxide Flotation
A total of 29 flotation tests were conducted on the Shaba (23 tests) and Safari North (6 tests) oxide samples. The tests consisted of 21 rougher flotation tests, five cleaner flotation tests, one bulk cleaner flotation test and two locked cycle flotation tests.
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The conclusions drawn from the oxide ore flotation testwork were:
Optimum grind size for both Shaba and Safari North was a P 80 of 106µm.
No regrinding was required.
Optimum Eh, as measured by a Ag/AgCl electrode, was -50mV for the Shaba and - 150mV for the Safari North.
Conditioning for 15 minutes with NaHS, prior to rougher flotation, was found to be critical for the successful flotation of the oxide ores.
The optimum conditions were used to perform locked cycle tests, with two stages of cleaner flotation, on both the Shaba and Safari North oxide samples.
NaHS additions, based on the LCT results, were moderate for Shaba at 2.0kg/t and moderately high for Safari North at 6.4kg/t.
Collector additions, based on the LCT results, were high for Shaba at 750g/t PAX and 450g/t AP407 and Safari North at 850g/t PAX and 550g/t AP407.
The Shaba LCT achieved a final concentrate grade of 48.2% Cu at a recovery of 71.6%.
The Safari North LCT achieved a final concentrate grade of 33.2% Cu at a recovery of 44.7%.
Additional testwork including mineralogy is recommended to help define the reasons for the poorer flotation response of the Safari North ore compared to the Shaba ore.
16.2.4 Sulphide Flotation
A total of 17 flotation tests were conducted on Shaba sulphide ore samples. The tests consisted of three rougher flotation, tests, 12 cleaner flotation tests, one LCT test and one unit cell flotation test. The Shaba sulphide composites 1 to 3 are the variability testwork samples.
The conclusions drawn from the sulphide ore flotation testwork were:
Unit cell or flash flotation was not considered to be beneficial for inclusion in the Kapulo processing plant design.
Optimum grind size was a P 80 of 75µm.
No regrinding was required.
The chalcopyrite and bornite rich composites gave poorer flotation responses than the master composite. The reasons for the poorer flotation could not be investigated as there was insufficient sample available for any further testwork. These samples are not considered to be representative of any potential feed material for a plant as both ore types would have to be mined simultaneously.
The optimum conditions were used to perform a locked cycle test, with two stages of cleaner flotation, on the Shaba sulphide master composite.
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Collector addition, based on the LCT results, was moderately low 60g/t Aerophine 3418A.
The Shaba variability samples showed a similar flotation response to the the master composite.
The Shaba sulphide variability Composite 1 required higher Aerophine 3418A additions, i.e.120g/t, to achieve acceptable flotation response.
The LCT achieved a final concentrate grade of 34.4% Cu and 90g/t Ag at recoveries of 93.2% for copper and 91.7% for silver.
16.2.5 Detailed Concentrate Assays
Samples of selected final concentrate were submitted for detailed chemical analysis, the results of which are summarised in Table 16.2.5_1.
Table 16.2.5_1 Kapulo Copper Project Final Concentrate Assays Summary
Preliminary Shaba Sulphide Composites Shaba Shaba Blended Analyte Units Master Pyrite Bornite Oxide Sulphide Composite Cu 34.2 33.6 41.2 41.6 37.9 % Ag 61 18 89 63 46 ppm As 458 65 389 218 204 ppm Bi 860 342 1690 752 1550 ppm Co 466 14 168 20 95 ppm Fe 26.1 21.7 22.8 2.4 24.0 % Hg <0.01 <0.1 <0.1 ------ppm Pb 1220 950 1540 376 1305 ppm S 30.4 25.4 30.2 10.0 31.1 % Zn 0.54 0.12 1.26 0.007 1.08 %
This data shows that there are no significant deleterious elements in the final concentrates, although there were elevated levels of bismuth, lead and zinc.
16.2.6 Thickening Testwork
Thickening testwork gave the following results:
Final sulphide concentrate - feed flux rate 0.25t/(m².h), underflow density 73.1% solids.
Final oxide tailings - feed flux rate 1.00t/(m².h), underflow density 58.9% solids.
Final sulphide tailings - feed flux rate 0.80t/(m².h), underflow density 62.5% solids.
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16.2.7 Filtration Testwork
Filtration tests showed that the final oxide concentrate had a specific filtration rate of 224kgDS/m²h at a moisture content of 17.7%. The specific filtration rate for the final sulphide concentrate was found to be 1,079kgDS/m²h at a moisture content of 9.0%. The results for the oxide concentrate should be treated with caution as the testwork was restricted due to low available sample mass
16.3 Process Design
The flowsheet was based on conventional comminution, flotation and filtration processes. The plant design is based upon crusher feed grades of 5.40% Cu for the oxide material and 4.37% Cu for the sulphide material, these values being derived from the head assays of the completed metallurgical testing.
The general plant design was based on the philosophy of providing a cost effective processing solution, whilst maintaining high levels of reliability, operability and maintainability. In order to achieve this, the following strategies were adopted:
The crushing plant will operate on a 365 days/year, 12h/day operating cycle, with a design availability within this period of 75% for a nominal throughput of 117t/h on oxide ore and 152t/h on sulphide ore.
Downstream of the crushing circuit, the plant will operate on a 365 day/year, 24h/day operating cycle with a design availability of 91.3% for a nominal combined ore throughput of 48.1 dry t/h on oxide ore and 62.5 dry t/h on sulphide ore.
The design basis assumes a moderate level of instrumentation and automation to minimise the operator requirement without introducing undue complexity and expense.
Adherence to well-proven and conservative design practice appropriate to the copper flotation industry.
The processing facility design is based upon the supply and installation of new processing equipment and comprises the following areas:
Crushing circuit - including crushers, conveyors, screens, crushed ore stockpile, dust extraction and ancillary equipment.
Comminution circuit – including ball mill, hydrocyclones, ore reclamation from crushed ore stockpile, pumps and ancillary equipment.
Flotation circuit – including rougher flotation, two stages of cleaner flotation, in stream analyser, pumps and ancillary equipment.
Concentrate handling – including concentrate thickener, concentrate filtration, concentrate bagging, pumps and ancillary equipment.
Tailings disposal – including tailings thickener, pumps and pipelines to the tailings storage facility (TSF) discharge point.
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Reagents – including storage, mixing and distribution of reagents throughout the processing plant.
Services:
Water including raw water from the raw water dam outlet, process water, tailings return water, fire water, potable water
Air including high pressure air for instruments, filtration and general plant air and low pressure air for flotation.
Infrastructure – including power station, office block, laboratory, workshop buildings, change room and ablutions.
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17 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES
17.1 Shaba Resource Estimate
Coffey Mining has estimated the maiden Mineral Resource for the Shaba Deposit, Kapulo Copper Project as at 28 th of May 2010. All grade estimation was completed using Ordinary Kriging (‘OK’) for copper and silver. This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralisation, and the style of mineralisation. The estimation was constrained with geological and mineralisation interpretations.
The Qualified Person responsible for the resource estimate is Mr Steve Le Brun, who is an Associate Resource Geologist for Coffey Mining Ltd. The Qualified Person’s certificate for Mr Le Brun is included in Section 23. The details of the resource estimate are summarized in the following sections.
17.1.1 Resource Database and Validation
Database
The drillhole database in the vicinity of Shaba contains a total of 79 resource holes drilled by Mawson West between 2007 and 2009. The majority of the drillholes were drilled angled at between 40° to 80° towards east (WGS84_35S grid). The database contains 72 diamond holes for 10,833m and 7 RC holes for 992m. The database also contained 32 surface channel and trenches. These were not used in the estimation process due to survey pickup inaccuracies.
The database used in the estimation study contained copper and silver grade information from 7,568 intervals.
All of the drillholes in the database have DGPS collar pick up surveys accurate to ±10cm relative to the survey base station. All of the holes in the database have been surveyed for downhole deviation. The unsurveyed holes are a mixture of holes that were blocked and couldn’t be surveyed and holes that had only recently been drilled.
Validation
Prior to loading data into the database, the following checks were carried out:
Hole depths for the geology log, survey log and assay intervals don’t exceed the hole depth.
Dates are in the correct format and factually correct.
That set limits e.g. northing, easting, assay values etc. aren’t exceeded.
That sample IDs retuned from the laboratory match the IDs in the drill log from the field.
Deviation data is checked for “believability” and data spikes due to magnetic influence are removed.
That valid codes e.g. lithology, geotechnical log etc have been used.
Sampling intervals are checked for gaps and overlaps.
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After the data was loaded into the database, the following checks were carried out:
A visual check that collar locations are correct.
QAQC data for standards, blanks and field duplicates submitted with the drilling samples are checked and reported on a weekly basis.
169 samples from across Shaba were submitted to a second laboratory as an additional QC check on assays within the database.
Prior to modelling the various checks were performed on the database:
Comparison of design set out co-ordinates versus final DGPS collar pick up.
Visual checks of drillhole traces.
No significant validation errors were detected in the database.
17.1.2 Geological Interpretation and Modelling
Oxidation and mineralised domain boundaries have been interpreted based on grade information and geological observations and wireframes of these boundaries were modelled to constrain resource estimation for the Shaba deposit. Interpretation and digitising of all constraining boundaries has been undertaken on cross sections orientated at 090º W-E (drill line orientation). The resultant digitised boundaries have been used to construct wireframe surfaces or solids defining the three-dimensional geometry of each interpreted feature. The interpretation and wireframe models have been developed using the Datamine mine planning software package.
Geological Interpretation
The geological sequence at Shaba is a west dipping tabular zone of copper mineralisation immediately in the footwall of a regionally extensive fault structure. The hangingwall of the fault is dominated by variously brecciated undifferentiated granitoids and sandstones. The footwall of the fault is underlain by a wedge package of sediments, mostly greywacke and shales, underlain by k-feldspar granitoids and minor undifferentiated mafic rock types, possibly of intrusive origin. The fault structure is characterised by the presence of fault gouge. Figure 17.1.2_1, shows a cross-section of the geological interpretation.
Mineralisation Interpretation
Modelling of the Shaba mineralisation consisted of an initial geological model of the fault gouge, granitoid, sedimentary units and ultramafics. The mineralisation occurs within a west dipping tabular orebody with a sharp hangingwall contact and a gradational footwall contact. High grade mineralisation is preferentially located in proximity to the hangingwall. Two mineralised zones were interpreted:
A hangingwall (HW) high grade zone with a fault defining the HW contact and the footwall contact based on a nominal 4.0% Cu drillhole grade.
A footwall (FW) low grade zone contiguous with HW zone defining its upper contact and the lower contact based on a nominal 0.5% Cu drillhole grade.
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Figure 17.1.2_1 Geological Cross-Section of the Shaba Deposit at 19,885mN
Showing the Lithological Coding
The downhole thickness of mineralised intervals ranged from 2m to 30m, with an average of 14.6m for the hangingwall (HW) and from 0.8m to 57m, with an average of 14.6m for the footwall (FW). A total of 2 mineralised units were created that had strike extents up to 400m.
Extrapolation of the interpreted mineralisation was limited to 20m along strike from known drilling and no more than 40m down dip from drill intercepts. Mineralisation has not been properly closed off at depth or along strike by current drilling. Figure 17.1.2_2, shows a cross- section of the mineralisation interpretation.
Figure 17.1.2_2 Cross-Section of the Shaba Deposit at 19,885mN
Showing the Mineralisation Domains
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Surface Cover and Weathering Profile
Thin to non-existent alluvium covers most of the Project, which is heavily vegetated. Weathering profiles for completely oxidised, partially oxidised, transitional and fresh material were generated. The average drillhole thicknesses of the weathered material are shown below:
Oxide : from 2m to 65m with average of 25m
Upper Transitional: from 2m to 75m with average of 24m
Lower Transitional: from 1m to 243m with average of 41m
Fresh: from 14m to 230m with average of 70m
Figure 17.1.2_3, shows a cross-section of the weathering profile.
Figure 17.1.2_3 Cross-Section of the Shaba Deposit at 19,885mN
Showing the Weathering Profile
17.1.3 Statistical Analysis of Composites and Top Cuts
The lengths of the samples were statistically assessed prior to selecting an appropriate composite length for undertaking statistical analyses, variography and grade estimation. Summary statistics of the sample length indicates that 89.1% of the samples were collected at 1m intervals, 4.1% were collected at 2m intervals and the remainder (6.8%) were sampled at irregular intervals less than 1m or greater than 2m. The data captured within the mineralisation wireframes was composited to a regular 2m downhole composite length.
The global effect of the compositing is summarised by area in Table 17.1.3_2, and shows negligible effect to the copper mean (3.3% - HW & 0.01% - FW); similarly, for the silver assays. The drop in variance is a natural effect of compositing, as a result of the change of support. The 2m composites were used for all statistical, geostatistical and grade estimation studies.
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Table 17.1.3_1 Shaba Deposit, Kapulo Copper Project Summary of Raw Statistics for Copper and Silver by Mineralised Zone
Total Raw Data Zone Cu (%) Cu (%) Ag (ppm) Ag (ppm) Mean Variance Mean Variance 100 (HW) 8.11 38.74 19.34 574.09 110 (FW) 1.35 1.58 2.82 18.16
Table 17.1.3_2 Shaba Deposit, Kapulo Copper Project Summary Statistics - 2m Composite Copper (%) (No High Grade Cuts Applied)
Coeff. % Mean Zone Count Min. Max. Mean Std. Dev. Variance Var. Difference 100 (HW) 336 0.04 37.71 7.84 5.31 28.21 0.68 3.3% 110 (FW) 571 0.04 11.42 1.34 1.04 1.09 0.78 0.01%
Table 17.1.3_3 Shaba Deposit, Kapulo Copper Project Summary Statistics - 2m Composite Silver (ppm) (No High Grade Cuts Applied)
Coeff. % Mean Zone Count Min. Max. Mean Std. Dev. Variance Var. Difference 100 (HW) 336 0.10 120.89 20.39 21.63 467.96 1.06 5.4% 110 (FW) 571 0.10 41.56 2.87 3.92 15.38 1.37 1.8%
Statistical analysis was carried out on the composited data for each unit to determine appropriate top cuts to apply to the data. The method of top cutting was to look at grade distribution histograms and the change in coefficient of variation (CV) and Sichel mean with the progressive removal of high grades. Top-cuts result in the cutting back of a small number of outlier high grade samples that can have a disproportionately large effect on the estimated grade. Figures 17.1.3_1 and 17.1.3_2 show type examples of the graphs that were examined from each mineralised zone to assess the top cutting of outlier assays. Coffey Mining has determined that no top-cut is required for either domain at Shaba.
Figure 17.1.3_1 shows the histogram plot for the Hangingwall zone (100) copper composites illustrating the strong positive tail to the distribution.
Figure 17.1.3_2 shows the histogram plot for the Footwall zone (110) copper composites illustrating the weaker positive tail to the distribution.
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Figure 17.1.3_1 Shaba - Histogram Plot for Cu% - Zone 100 (Hangingwall)
Figure 17.1.3_2 Shaba - Histogram Plot for Cu% - Zone 110 (Footwall)
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17.1.4 Bulk Density Data
A total of 1,269 density readings were taken using the whole-core water immersion method from 57 diamond holes drilled at Shaba. These readings were grouped by lithology type and weathering profile and an average bulk density reading was calculated for each sub-domain, excluding the mineralised domains. Bulk densities were calculated by mineralisation domain and weathering profile using regression statistics (Figure 17.1.4_1) against copper grade. Sub-domains with insufficient readings to generate averaged or regression statistics were assigned a weighted value derived from the fresh measurements and the weathering profile of the undifferentiated granitoid. Table 17.1.4_1 summarises the density readings.
Figure 17.1.4_1 Bulk Density Regression Statistics for Hangingwall Domain (100) Fresh Material SG’s for Fresh Zone 100 4
3.5
3
2.5 y = 0.0262x + 2.628 R² = 0.424
2 SG
1.5
1
0.5
0 0 5 10 15 20 25 30 35 40 CU_PCT
Mineralisation domains were assigned a bulk density values based upon the regression statistics for the density and copper values, weighted by weathering profile. Table 17.1.4_2 summarises the density assignments.
Figure 17.1.4_2, shows a cross-section of the final model and the weathering coding used to assign bulk densities.
Figure 17.1.4_3, shows a cross-section of the final model and the lithological coding used to assign bulk densities.
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Table 17.1.4_1 Shaba Deposit, Kapulo Copper Project Summary Bulk Density Measurements By Lithology and Weathering Profile (Weighted values in red)
Lithology Undiff. Granitoid Sandstone Fault Gauge Sediment K-feldspar Granitoid Ultramafic Average Average Average Average Average Average “GROUND” Field No. of No. of No. of No. of No. of No. of Weathering Profile Bulk Density Bulk Density Bulk Density Bulk Density Bulk Density Bulk Density (in Final Model) Samples Samples Samples Samples Samples Samples (t/m³) (t/m³) (t/m³) (t/m³) (t/m³) (t/m³) Completely Oxidised 3 105 2.31 12 2.38 2 2.12 -- 2.28 -- 2.36 -- 2.48 Partially Oxidised 4 129 2.53 47 2.58 2 2.32 1 2.50 2 2.59 -- 2.72 Transitional 5 118 2.58 173 2.58 2.37 5 2.55 -- 2.64 -- 2.77 Fresh 6 65 2.59 21 2.59 8 2.38 209 2.56 375 2.65 22 2.78
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Table 17.1.4_2 Shaba Deposit, Kapulo Copper Project Shaba Mineralisation Bulk Density Statistics (Regression Equations)
Hangingwall (HW) Footwall (FW) Weathering Profile GROUND Zone 100 Zone 110 Completely Oxidised 3 0.0262*Cu% + 2.339 0.0449*Cu% + 2.334 Partially Oxidised 4 0.0262*Cu% + 2.549 0.0449*Cu% + 2.555 Transitional 5 0.0262*Cu% + 2.602 0.0449*Cu% + 2.6075 Fresh 6 0.0262*Cu% + 2.628 0.0449*Cu% + 2.6335
Figure 17.1.4_2 Cross-Section of the Shaba Block Model at 19,885mN Showing the Weathering Coding Used for Bulk Density Assignment
Figure 17.1.4_3 Cross-Section of the Shaba Block Model at 19,885mN Showing the Lithological Coding Used for Bulk Density Assignment
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17.1.5 Variography
The Isatis geostatistical software package was used to analyse the Shaba Deposit variography for Zones 100 and 110. Traditional semi-variograms were initially used to analyse the spatial variability of the copper and silver 2m composites. In order to obtain a better structured variography, the data was transformed into Gaussian space using a Hermite polynomial curve-fitting function within Isatis, and back-transformed using the point anamorphosis file, with the sill fitted to the original variance.
In this document, the term ‘variogram’ is used as a generic word to designate the function characterising the variability of variables versus the distance between two samples.
Variographic analysis using Isatis software on the separate domains resulted in poorly structured variograms, therefore variograms for combined domains were generated and which produced moderately structured variograms (see Figure 17.1.5_1). The rotations are tabulated as input into Isatis (geological convention), with X representing rotation around Z axis, Y representing rotation around Y` axis and Z representing rotation around X``. Dip and dip direction of major, semi-major and minor axes of continuity are also referred to in the text.
Search ellipsoids for the Ordinary Kriging (OK) estimate were oriented to lie parallel to the overall dip and dip direction of the modelled mineralisation and the ellipsoids were flattened in the plane of the mineralisation.
Table 17.1.5_1 summarises the resulting copper and silver variogram model parameters from the combined zones with sills adjusted to the domain variance. Figures 17.1.5_1 to 17.1.5_4 summarise display the resulting Gaussian transformed and back transformed variogram models from the study.
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Table 17.1.5_1 Shaba Deposit, Kapulo Copper Project Variogram Parameters for Mineralised Zones
Orientation Range 1 (m) Range 2 (m) Major Semi-Major Minor Zone C0 C1 Semi- C2 Semi- Major Minor Major Minor Dip Azi Dip Azi Dip Azi Major Major (°) (°) (°) (°) (°) (°) 100 0 0 60 90 -30 270 2.436 11.794 55 15 15 15.242 70 60 27.5 Cu 110 0 0 60 90 -30 270 0.138 1.666 55 15 15 0.861 70 60 27.5 100 0 0 60 90 -30 270 47.395 227.256 20 20 17.5 171.548 55 55 25 Ag 110 0 0 60 90 -30 270 1.532 7.346 20 20 17.5 5.546 55 55 25
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Figure 17.1.5_1 Shaba Deposit (Combined Zones) Copper – Gaussian Transformed Variography
upper left (Major) 0º 000º upper right (Semi-Major) 60º 090º, lower left (Minor) -30º 270º
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Figure 17.1.5_2 Shaba Deposit (Combined Zones) Copper – Back Transformed Variography (to Original Variance)
upper left (Major), upper right (Semi-Major), lower left (Minor)
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Figure 17.1.5_3 Shaba Deposit (Combined Zones) Silver – Gaussian Transformed Variography
upper left (Major) 0º 000º upper right (Semi-Major) 60º 090º, lower left (Minor) -30º 270º
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Figure 17.1.5_4 Shaba Deposit (Combined Zones) SIlver – Back Transformed Variography (to Original Variance)
upper left (Major), upper right (Semi-Major), lower left (Minor)
17.1.6 Block Model
Introduction
A three-dimensional block model was constructed for the Shaba deposit, covering all the interpreted mineralisation zon es and including suitable additional waste material to allow later pit optimisation studies.
Block Construction Parameters
A sub-block model was created in the local grid using Datamine mining software. Block coding was completed on the basis of the block centroid, wherein a centroid falling within any wireframe was coded with the wireframe solid attribute. The model extents, block sizes and attributes are summarized in Table 17.1.6_1 below.
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Table 17.1.6_1 Shaba Deposit, Kapulo Copper Project Block Model Summary
Type Y X Z Minimum Coordinates 19,300 9,550 640 Maximum Coordinates 20,100 10,470 1,340 User Block Size (m) 20 10 5 Minimum Block Size (m) 2.0 1.0 0.5 Rotation 0° 0° 0°
Total Blocks – Mineralisation 495,789 Total Blocks – Final Model 3,024,404
Attribute Name Type Decimals Background Description Classification (1=measured; 2=indicated; 3=inferred, RESCAT Integer - 4 4=unclassified) RCLASS Character - unknown Classification (Measured; Indicated; Inferred, Unclassified) DENSITY Real 2 2.65 Density Weathering Code (3=oxidised, 4=partially oxidised, GROUND Integer - - 5=Transitional, 6=fresh, 99=dumps, 199=openpit, 0=air OXIDN Character - - Weathering (OXID, TRN1, TRN2, PMRY, DUMP, AIR) ZONE Integer - x 2 domains; HW Zone=100, FW Zone=110 GEOL Character - unknown Rock types (GG, SST, FG, GA, SEDT, MM) MINED Integer - 0 Is Mined Flag (0=not mined, 1=mined openpit) CUc Real - 0.005 Estimated Copper from top-cut composites AGc Real - 0.005 Estimated Silver from top-cut composites
The parent block size was selected on the basis of the average drill spacing (40m section spacing) and the variogram models. A parent block size of 20mE x 10mN x 5mRL was selected as appropriate. Sub-blocking to a 2.0mE x 1.0mN x 0.5mRL size (1/10 th parent block) was completed to ensure adequate volume representation.
The attributes coded into the block models included the weathering and mineralisation models. A visual review of the wireframe solids and the block model indicates robust flagging of the block model.
Bulk density has been coded to the block model based on the weathering profile and lithology type. The average bulk density for each subdivision, as presented in Table 17.1.4_2, was coded via a block model script. A description of the bulk density measurement methodology can be found in Section 11.4.
DTMs/3DMs used to construct the model:
wf_SH_topo_tr/pt.dm Original pre-mined surface topography DTM
wf_pit_tr/pt.dm Surveyed extent of artisanal workings
wf_dumps_tr/pt.dm Surveyed extent of artisanal spoil dumps
wf_BOX_tr/pt.dm Base of complete oxidation
wf_TOT_tr/pt.dm Top of transitional material
wf_TOF_tr/pt.dm Top of fresh material
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wf_SST_tr/pt.dm Sandstone 3D model
wf_SH_UFG_tr/pt.dm Upper DTM surface for Fault Gouge
wf_LFG_tr/pt.dm Lower DTM surface for Fault Gouge
wf_SEDT_tr/pt.dm Sediment 3D model
wf_MM_tr/pt.dm Ultramafics 3D model
wf_SH_Z100_tr/pt.dm Hangingwall mineralisation – Zone 100
wf_SH_Z110_tr/pt.dm Footwall mineralisation – Zone 110
wf_SH_Z120_tr/pt.dm Lower Footwall mineralisation – Zone 110
17.1.7 Grade Estimation
Introduction
Resource estimation for the Shaba mineralisation was completed using OK within all Domains. Inverse Distance Squared estimates were also completed within these domains to allow comparison with the OK estimate.
OK Estimate
Copper and silver grades were estimated into the block model using OK after top cuts had been applied to the original 2m composite data.
No neighbourhood testing was carried out for this estimation (Table 17.1.7_1). Second and third passes with 2x and 3x multipliers for the search radii were applied, although the majority of the blocks were estimated in the first pass.
Table 17.1.7_1 Shaba Deposit, Kapulo Copper Project Sample Search Parameters for OK Estimate
Search Orientation Search Radii Number of Samples Pass Semi- Domain Unit Major Minor Max / No. Bearing Plunge Dip Major Min Max Axis (m) Axis (m) Hole Axis (m) 100 Copper 1 270 0 55 120 120 50 8 20 - 110 Copper 1 270 0 55 120 120 50 8 20 - 100 Silver 1 270 0 55 100 100 50 8 20 - 110 Silver 1 270 0 55 100 100 50 8 20 -
Validation
A variety of validation checks were done on the data prior to estimation to ensure that composite values and locations matched the original data in the database. After estimation was complete, validation checks on the block model included:
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Checks that the majority of blocks had filled with grade.
Correct assignment of density, classification, unit, domain and lithology information.
Volume comparison between the mineralisation wireframe and the mineralized units in the block model.
Visual inspection of estimated blocks against the informing composites and original drillhole data (Figure 17.1.7_1).
Comparison plots of average informing composite grade and average block model grade by mineralized unit (Figure 17.1.7_2).
Statistical comparison of 2m composites with de-clustered composites and block model grades.
Figure 17.1.7_1 Shaba Block Model Grades
Overall, there was a good comparison between the informing composite data input into the model and the resulting block grades. Blocks filled by direct assignment e.g. density, lithology and classification had all filled correctly.
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Figure 17.1.7_2 Examples of Comparison Plots of Block Grades and Informing Composite Grades for Individual Units by 50m Northing Windows
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17.1.8 Resource Reporting and Classification
Introduction
The resource estimate for the Shaba update has been classified in accordance with the criteria laid out in the JORC Code (2004) and Canadian National Instrument 43-101. Indicated and Inferred Resources were defined based on data quality, data density and geological and/or grade continuity after detailed consideration of the JORC and CNI43-101 guidelines.
Criteria for Resource Categorisation
The classification of the Shaba Resource was developed from the confidence levels of key criteria including drilling methods, geological understanding and interpretation, sampling, data density, data location, data quality, grade estimation and quality of the estimates. The current Shaba Resource is reported as a combination of Indicated and Inferred material, depleted for surface artisanal workings, the key criteria are listed in Table 17.1.8_1 below. Figure 17.1.8_1 illustrates the classification applied to the resource model.
Table 17.1.8_1 Shaba Deposit, Kapulo Copper Project Confidence Levels of Key Categorisation Criteria
Item Discussion Confidence Drilling Techniques RC/Diamond – standard industry approach. High Logging Standard nomenclature applied and apparent high quality. High Drill Sample Recovery Recorded as good. High Sub-sampling Techniques Industry standard for both RC and diamond. High and Sample Preparation Quality of Assay Data Good internal and external QAQC check data for majority of the High chemical assays. Verification of Sampling QAQC analysis is within industry acceptable standards High and Assaying Location of Sampling Drillhole collars surveyed by DGPS and all drillholes have been High Points surveyed downhole for deviation Data Density and Nominal 40 x 40m overall spacing over whole deposit. Moderate Distribution Audits or Reviews Site drilling and sampling procedures reviewed by Coffey Mining. High Database Integrity No material errors identified High Geological Interpretation Further infill drilling may change the mineralisation shapes and the Moderate geological interpretation. Estimation and Estimates based on statistical and geostatistical analysis. Moderate Modelling Techniques Cutoff Grades Range of cutoff grades reported. High Mining Factors or No ore loss or dilution factored in. The effect of emulating SMU N/A Assumptions (change of support) has not been investigated.
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Figure 17.1.8_1 Shaba Classified Block Model
Grade Tonnage Reporting
The resources for the Shaba prospect reported above a 0.3% Cu cutoff are reported below in Table 17.1.8_2 by mineralisation domain and by oxidation profile in Table 17.1.8_3 below.
Table 17.1.8_2 Shaba Deposit, DRC DRC – Indicated/Inferred Resource Estimate Reported Above 0.3% Cu Cutoff Ordinary Kriged Estimate Using 2m Cut Cu% Composites Parent Cell Dimensions of 10mEW by 20mNS by 5mRL Depleted for Artisanal Surface Mining Categorised according to JORC Code (2004)
Grade Cu Metal Grade Ag Metal Classification Zone Mt (Cu%) (Tonnes) (Ag ppm) (kOz) HW 1.711 7.7 131,650 23 1,241 Indicated FW 3.280 1.4 45,645 3 297.9 Subtotal 4.990 3.6 177,295 10 1,540 HW 0.195 7.4 14,400 30 189.3 Inferred FW 0.829 1.1 8,790 5 1245 Subtotal 1.024 2.3 23,190 10 313.8 Note: Figures have been rounded
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Table 17.1.8_3 Shaba Deposit, DRC Indicated/Inferred Resource Estimate by Oxidation Reported Above 0.3% Cu Cutoff Ordinary Kriged Estimate Using 2m Cut Cu% Composites Parent Cell Dimensions of 10mEW by 20mNS by 5mRL Depleted for Artisanal Surface Mining Categorised according to JORC Code (2004)
Grade Cu Metal Grade Ag Metal Classification Zone Mt (Cu%) (Tonnes) (Ag ppm) (kOz) Oxide 0.211 4.4 9,385 6 42.1 Upr Trans 0.170 3.8 6,400 7 38.1 Indicated Lwr Trans 0.205 5.5 11,335 17 109.3 Fresh 4.405 3.4 150,165 10 1,349.6 Subtotal 4.990 3.6 177,295 10 1,540 Oxide 0.007 1.7 100 6 1.9 Upr Trans 0.002 1.6 40 7 0.6 Inferred Lwr Trans 0.009 1.4 125 16 2.2 Fresh 1.006 2.3 22,900 10 309.1 Subtotal 1.024 2.3 23,190 10 313.8 Note: Figures have been rounded
Adjacent to the artisanal workings are approximately 31,000t of artisanal spoil dumps with an average grade of ~4% Cu and ~3ppm Ag, calculated from global average grades from three trenches across the dumps. These dumps are not classified and do not form part of the existing Resources. The potential quantity and grade of the dumps is conceptual in nature, and there has been insufficient exploration to define a Mineral Resource on the dumps and it is uncertain if further exploration will result in discovery of a Mineral Resource.
17.1.9 Comments and Recommendations
Drilling at the Shaba prospect has defined an Indicated and Inferred Resource as outlined in the above table.
Further infill and extensional drilling is required to raise the level of confidence and extend the Inferred Resources.
17.2 Safari Resource Estimate
Coffey Mining has estimated the maiden Mineral Resource for the Safari Deposit as at 15 th of December 2010. All grade estimation was completed using OK for copper and silver. This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralisation, and the style of mineralisation. The estimation was constrained with geological and mineralisation interpretations.
The Qualified Person responsible for the resource estimate is Mr Steve Le Brun, who is an Associate Resource Geologist for Coffey Mining. The Qualified Person’s certificate for Mr Le Brun is included in Section 23. The details of the resource estimate are summarized in the following sections.
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17.2.1 Resource Database and Validation
Database
The drillhole database in the vicinity of Safari (North and South) contains a total of 81 resource holes drilled by Mawson West between 2007 and 2010 (69 at Safari North and 12 at Safari South). The majority of the drillholes were drilled angled at between 40° to 80° towards east (WGS84_35S grid). The database contains 44 DDH for 3,545m and 37 RC holes for 2,180m. The database also contained 28 surface channel and trenches and 4 geotechnical DDH. The trenches and channels were not used in the estimation process due to survey pickup inaccuracies. The geotechnical holes were not used due to lack of assay and geological information. Three drillholes (07SNDH004, 005, 006) from the 2007 programme were included in the estimation database, but doubts regarding the reliability of their assay data, geologically inconsistent with neighbouring sections, means that blocks within the area of influence of these drillholes are restricted to Inferred status only until they can be redrilled and assayed during 2011.
The database used in the estimation study contained copper and silver grade information from 9,990 intervals.
All of the drillholes in the database have DGPS collar pick up surveys accurate to ±10cm relative to the survey base station. All of the holes in the database have been surveyed for downhole deviation. The unsurveyed holes are a mixture of holes that were blocked and couldn’t be surveyed and holes that had only recently been drilled.
Validation
Prior to loading data into the database, the following checks were carried out:
Hole depths for the geology log, survey log and assay intervals don’t exceed the hole depth.
Dates are in the correct format and factually correct.
That set limits e.g. northing, easting, assay values etc. aren’t exceeded.
That sample IDs retuned from the laboratory match the IDs in the drill log from the field.
Deviation data is checked for “believability” and data spikes due to magnetic influence are removed.
That valid codes e.g. lithology, geotech log etc have been used.
Sampling intervals are checked for gaps and overlaps.
After the data has been loaded into the database, the following checks are carried out:
A visual check that collar locations are correct.
QAQC data for standards, blanks and field duplicates submitted with the drilling samples are checked and reported on a weekly basis.
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A total of 53 samples from across Safari were submitted to a second laboratory as an additional QC check on assays within the database.
Prior to modelling, the various checks were performed on the database:
Comparison of design set out co-ordinates versus final DGPS collar pick up.
Visual checks of drillhole traces.
No significant validation errors were detected in the database.
17.2.2 Geological Interpretation and Modelling
Based on grade information and geological observations, oxidation and mineralised domain boundaries have been interpreted and wireframes modelled to constrain resource estimation for the Shaba deposit. Interpretation and digitising of all constraining boundaries has been undertaken on cross sections orientated at ~090º W-E (drill line orientation). The resultant digitised boundaries were used to construct wireframe surfaces or solids defining the three- dimensional geometry of each interpreted feature. The interpretation and wireframe models have been developed using the Datamine mine planning software package.
Geological Interpretation
The geological sequence at Safari is a west dipping tabular zone of copper mineralisation immediately in the footwall of a regionally extensive fault structure. The hangingwall of the fault is dominated by variously brecciated sandstones and minor undifferentiated granitoids. The footwall of the fault is underlain by a wedge package of sediments, mostly greywacke and shales, underlain by k-feldspar granitoids and minor undifferentiated ultramafic units, possibly of intrusive origin. The fault structure is characterised by the presence of fault gouge and varies in dip from ~60° in the southwest portion of Safari North to ~80° through the northern portions of the deposit.
Figure 17.2.2_1, shows a cross-section of the geological interpretation.
Mineralisation Interpretation
Modelling of the Safari mineralisation consisted of an initial geological model of the fault gouge, granitoid, sedimentary units and ultramafic units. The mineralisation occurs within a west dipping tabular orebody with a variously sharp hangingwall contact and a gradational footwall contact. High grade mineralisation is commonly located in proximity to the hangingwall. Two mineralised zones were interpreted:
A hangingwall (HW) high grade zone with a footwall contact based on a nominal 3.0% Cu drillhole grade. (Safari North – Main high grade - Zone 200; Footwall high grade - Zone 205; Safari South - Main high grade - Zone 300)
A footwall (FW) low grade zone contiguous with HW zone defining its upper contact and the lower contact based on a nominal 0.5% Cu drillhole grade. (Safari North – Main low grade - Zone 210; Footwall low grade – Zone 215; Safari South - Main low grade – Zone 310)
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Figure 17.2.2_1 Geological Cross-Section of the Safari Deposit at 17,400mN Showing the Lithological Coding
For Safari North, the downhole thickness of mineralised intervals ranged from 1m to 23m, with an average of 6.25m for the hangingwall and from 1m to 18m, with an average of 5.7m for the footwall. A total of 2 mineralised units were created that had strike extents up to 325m.
For Safari South, the downhole thickness of mineralised intervals ranged from 1m to 8m, with an average of 4.25m for the hangingwall and from 2m to 13m, with an average of 6.5m for the footwall. A total of two mineralised units were created that had strike extents up to 240m.
Extrapolation of the interpreted mineralisation was limited to 20m along strike from known drilling and no more than 40m down dip from drill intercepts. Mineralisation has not been properly closed off at depth or along strike by current drilling, although it does appear to be diminishing in tenor. The current drillhole coverage for Safari South indicates that the high- grade hangingwall material is pinching out down dip.
Surface Cover and Weathering Profile
Thin to non-existent alluvium covers most of the Project, which is heavily vegetated. Weathering profiles for completely oxidised, partially oxidised, transitional and fresh material were generated. The average drillhole thicknesses of the weathered material are shown below:
Oxide : from 2m to 65m with average of 25m
Upper Transitional: from 2m to 75m with average of 24m
Lower Transitional: from 1m to 243m with average of 41m
Fresh: from 14m to 230m with average of 70m
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Figure 17.2.2_2, shows a cross-section of the weathering profile.
Figure 17.2.2_2 Cross-Section of the Safari Deposit at 19,885mN Showing the Weathering Profile
17.2.3 Statistical Analysis of Composites and Top Cuts
The lengths of the samples were statistically assessed prior to selecting an appropriate composite length for undertaking statistical analyses, variography and grade estimation. Summary statistics of the sample length indicates that 98.3% of the samples were collected at 1m intervals; 0.5% were collected at 2m intervals and the remainder (1.2%) were sampled at irregular intervals less than 1m or greater than 2m.
The data captured within the mineralisation wireframes was composited to a regular 2m downhole composite length. Zones 200 and 205 and Zones 210 and 215 were amalgamated for statistical analysis into Zones 200 and 210 respectively.
Table 17.2.3_1 Safari Deposit, Kapulo Copper Project Summary of Raw Statistics for Copper (%) and Silver (ppm) by Mineralised Zone
Total Raw Data Zone Cu (%) Ag (ppm) Mean Variance Mean Variance 200 (HW) 4.90 9.216 10.4 338.48 210 (FW) 1.52 2.56 3.49 43.63 300 (HW) 4.23 2.42 0.48 0.12 310 (FW) 1.03 0.50 1.00 1.80
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The global effect of the compositing is summarised by area in Table 17.2.3_2 and Table 17.2.3_3, and shows negligible effect to the copper mean (±4% - HW & (±1.4% - FW); similarly, for the silver assays in the footwall, although the hangingwall shows larger reductions due to reducing the effect of high grade outlier intervals (max value of 317ppm Ag). The drop in variance is a natural effect of compositing, as a result of the change of support. The 2m composites were used for all statistical, geostatistical and grade estimation studies.
Table 17.2.3_2 Safari Deposit, Kapulo Copper Project Summary Statistics - 2m Composite Copper (%) (No High Grade Cuts Applied)
Std. Coeff. % Mean Zone Count Min. Max. Mean Variance Dev. Var. Difference 200 (HW) 188 0.05 12.7 4.71 2.89 8.37 0.61 4.0% 210 (FW) 242 0.05 7.81 1.54 1.51 2.28 0.98 1.4% 300 (HW) 19 1.55 5.98 4.20 1.15 1.32 0.27 0.7% 310 (FW) 43 0.14 2.95 1.04 0.62 0.39 0.60 1.0%
Table 17.2.3_3 Safari Deposit, Kapulo Copper Project Summary Statistics - 2m Composite Silver (ppm) (No High Grade Cuts Applied)
Std. Coeff. % Mean Zone Count Min. Max. Mean Variance Dev. Var. Difference 200 (HW) 184 0.1 179.76 9.34 14.61 213.65 1.57 11.4% 210 (FW) 238 0.1 55.75 3.45 6.24 38.92 1.81 1.1% 300 (HW) 19 0.15 1.20 0.48 0.26 0.07 0.54 0.0% 310 (FW) 43 0.10 4.40 0.97 1.18 1.39 1.21 3.0%
A statistical analysis was carried out on the composited data for each unit to determine appropriate top cuts to apply to the data. The method of top cutting was to look at grade distribution histograms and the change in coefficient of variation (CV) and Sichel mean with the progressive removal of high grades. Top-cuts result in the cutting back of a small number of outlier high grade samples that can have a disproportionately large effect on the estimated grade. Figures 17.2.3_1 and 17.2.3_2 show type examples of the graphs that were examined from each mineralised zone to assess the top cutting of outlier assays. Coffey Mining has determined that no copper top-cut is required for either domain at Safari. A 50g/t and a 20g/t top cut was applied to the silver composites for the HW and FW domains respectively.
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Figure 17.2.3_1 Safari - Histogram Plot for Cu% - Zone 200 (Hangingwall) Histogram Plot (Weighted) (ZONE 200)
7
6
5
4
3
2 Frequency(%)
1
0 0 1 2 3 4 5 6 7 8 9 101112 CU_PCT (%)
Figure 17.2.3_1 shows the histogram plot for the Hangingwall zone (200) copper composites illustrating the moderate positive tail to the distribution.
Figure 17.2.3_2 shows the histogram plot for the Footwall zone (210) copper composites illustrating the weaker positive tail to the distribution.
Figure 17.2.3_2 Safari - Histogram Plot for Cu% - Zone 210 (Footwall) Histogram Plot (Weighted) (ZONE 210)
25
20
15
10 Frequency(%) 5
0 0 1 2 3 4 5 6 7 8 CU_PCT (%)
17.2.4 Bulk Density Data
A total of 543 density readings were taken from water immersion of whole core from 33 diamond holes drilled at Safari. These readings were grouped by lithology type and weathering profile and an average bulk density reading was calculated for each sub-domain excluding the mineralised domains. Bulk densities were calculated for mineralisation domains and weathering profile using regression statistics vs. copper grade. Sub-domains with insufficient readings to generate averaged or regression statistics were assigned a weighted value derived from the fresh measurements and the weathering profile of the undifferentiated granitoid. Table 17.2.4_1 summarises the density readings.
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Table 17.2.4_1 Safari Deposit, Kapulo Copper Project Summary Bulk Density Measurements By Lithology and Weathering Profile.
Lithology Average Bulk Density (t/m³) Average Bulk Density (t/m³) “GROUND” Field Undiff. Fault K-feldspar Weathering Profile Sandstone Sediment Ultramafic (in Final Model) Granitoid Gauge Granitoid Completely Oxidised 3 2.31 2.38 2.12 2.28 2.36 -- Partially Oxidised 4 2.53 258 2.32 2.50 2.59 -- Transitional 5 2.58 2.58 2.37 2.55 2.64 2.77 Fresh 6 2.59 2.59 2.39 2.56 2.65 2.78
Mineralisation domains were assigned a bulk density values based upon the regression statistics for the SG and copper values, weighted by weathering profile. Table 17.2.4_2 summarises the density assignments.
Table 17.2.4_2 Safari Deposit, Kapulo Copper Project Safari Mineralisation Bulk Density Statistics (Regression Equations)
Hangingwall (HW) Footwall (FW) Weathering Profile GROUND Zone 200/300 210/310 Completely Oxidised 3 0.0178*Cu% + 2.147 0.0362*Cu% + 2.223 Partially Oxidised 4 0.1036*Cu% + 1.680 0.1036*Cu% + 1.680 Transitional 5 -0.0276*Cu% + 2.770 -0.4134*Cu% + 2.791 Fresh 6 -0.0163*Cu% + 2.623 -0.0238*Cu% + 2.578
Corresponding density values were calculated for every top-cut copper composite assay. Block model densities were estimated from the calculated drillhole density values and applied if greater than the background host rock bulk density.
17.2.5 Variography
The Isatis geostatistical software package was used to analyse the Safari Deposit variography. Traditional semi-variograms were initially used to analyse the spatial variability of the copper, silver and density 2m composites. In order to obtain a better structured variography, Gaussian transformed variograms were calculated for the silver assays, with traditional variograms calculated for the density composites copper composites and combined zone density composites.
In this document, the term ‘variogram’ is used as a generic word to designate the function characterising the variability of variables versus the distance between two samples.
Variographic analysis for Safari South (Zones 300/310) produced very poor variograms due to the limited number of available composite values; therefore the variograms from Safari North were applied (see Figures 17.2.5_1 to 17.2.5_3). The rotations are tabulated as input into Isatis (geological convention), with X representing rotation around Z axis, Y representing rotation around Y` axis and Z representing rotation around X``. Dip and dip direction of major, semi-major and minor axes of continuity are also referred to in the text.
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Figure 17.2.5_1 Safari Deposit (Combined Zones) Copper – Correlogram
upper left (Major) 0º 000º upper right (Semi-Major) 60º 090º, lower left (Minor) -30º 270º
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Figure 17.2.5_2 Safari Deposit (Combined Zones) Silver – Correlogram
upper left (Major) 0º 000º upper right (Semi-Major) 60º 090º, lower left (Minor) -30º 270º
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Figure 17.2.5_3 Safari Deposit (Combined Zones) Density – Traditional Variography
upper left (Major), upper right (Semi-Major), lower left (Minor)
Search elli psoids for the OK estimate were oriented with the major axis along the plunge direction, approximately 25 ° south-west, and the ellipsoids were flattened in the plane of the mineralisation.
Table 17.2.5_1 summarises the resulting variogram model parameters for copper, silver and density, from the combined zones with sills adjusted to the domain variances.
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Table 17.2.5_1 Safari Deposit, Kapulo Copper Project Variogram Parameters for Mineralised Zones
Orientation Range 1 (m) Range 2 (m) Major Semi-Major Minor Zone C0 C1 Semi- C2 Semi- Major Minor Major Minor Dip Azi Dip Azi Dip Azi Major Major (o) (o) (o) (o) (o) (o) 200/205 -24 208 45 145 35 280 2.75 3.95 20 15 15 1.75 30 52.5 10 Cu 210/215 -24 208 45 145 35 280 1.0 0.425 30 47 10 0.75 85 55 10 200/205 -24 208 45 145 35 280 14.19 26.03 40 25 40 19.4 60 90 7.5 Ag 210/215 -24 208 45 145 35 280 11.25 6.159 55 25 60 3.413 70 50 10 200/205 -24 208 45 145 35 280 0.009 0.006 175 25 5 0.022 400 35 7.5 Density 210/215 -24 208 45 145 35 280 0.007 0.005 175 25 5 0.018 400 35 7,5 300 -24 208 45 145 35 280 2.75 3.95 20 15 15 1.75 30 52.5 10 Cu 310 -24 208 45 145 35 280 1.0 0.425 30 47 10 0.75 85 55 10 300 -24 208 45 145 35 280 14.19 26.03 40 25 40 19.4 60 90 7.5 Ag 310 -24 208 45 145 35 280 11.25 6.159 55 25 60 3.413 70 50 10 300 -24 208 45 145 35 280 0.009 0.006 175 25 5 0.022 400 35 7.5 Density 310 -24 208 45 145 35 280 0.007 0.005 175 25 5 0.018 400 35 7,5
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17.2.6 Block Model
Introduction
A three-dimensional block model was constructed for the Safari deposit, covering all the interpreted mineralisation zones and including suitable additional waste material to allow later pit optimisation studies.
Block Construction Parameters
A sub-block model was created in the local grid using Datamine mining software. Block coding was completed on the basis of the block centroid, wherein a centroid falling within any wireframe was coded with the wireframe solid attribute. The model extents, block sizes and attributes are summarized in Table 17.2.6_1 below.
Table 17.2.6_1 Safari Deposit, Kapulo Copper Project Block Model Summary
Type Y X Z Minimum Coordinates 16,700 9,300 950 Maximum Coordinates 17,800 10,050 1,285 # Blocks 55 75 75 User Block Size (m) 20 10 5 Minimum Block Size (m) 2.0 1.0 0.50 Rotation 0° 0° 0°
Total Blocks – Mineralisation 39,466 Total Blocks – Final Model 792,679
Attribute Name Type Decimals Background Description Classification RESCAT Integer - 4 (1=measured; 2=indicated; 3=inferred, 4=unclassified) Classification RCLASS Character - UNC (MEAS; IND; INF, UNC) Estimated Density from top-cut composites or DENSITY Real 2 2.65 assigned by rock/weathering profile Weathering Code GROUND Integer - 6 (3=oxidised, 4=partially oxidised, 5=Transitional, 6=fresh, 99=dumps, 199=open pit, 0=air) Weathering OXIDN Character - PMRY (OXID, TRN1, TRN2, PMRY, DUMP, AIR) 4 mineralisation domains; Safari North: HW ZONE Integer - 0 Zone=200/205, FW Zone=210/215 Safari South: HW Zone=300, FW Zone=310 Rock types GEOL Character - unknown (GG, SST, FG, GA, SEDT, MM) Is Mined Flag MINED Integer - 0 (0=not mined, 1=mined open pit) CUc Real - 0.005 Estimated Copper from top-cut composites AGc Real - 0.005 Estimated Silver from top-cut composites
The parent block size was selected on the basis of the average drill spacing (40m section spacing) and the variogram models. A parent block size of 10mE x 20mN x 5mRL was selected as appropriate. Sub-blocking to a 1.0mE x 2.0mN x 0.5mRL size (1/10 th parent block) was completed to ensure adequate volume representation.
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The attributes coded into the block models included the weathering and mineralisation models. A visual review of the wireframe solids and the block model indicates robust flagging of the block model.
Bulk density has been coded to the block model based on the weathering profile and lithology type. The average bulk density for each subdivision, as presented in Table 17.2.4_2, was coded via a block model script. Bulk densities for mineralised domains were estimated directly into the block model. A description of the bulk density measurement methodology can be found in Section 11.4.
wf_org_topo_tr/pt.dm Original pre-mined surface topography DTM
wf_SN_pit_tr/pt.dm Surveyed extent of Safari North artisanal workings
wf_SS_pit_tr/pt.dm Surveyed extent of Safari South artisanal workings
wf_BOX_tr/pt.dm Base of complete oxidation
wf_TOT_tr/pt.dm Top of transitional material
wf_TOF_tr/pt.dm Top of fresh material
wf_GG_tr/pt.dm Undifferentiated Granitoid DTM
wf_UFG_tr/pt.dm Upper DTM surface for Fault Gouge
wf_LFG_tr/pt.dm Lower DTM surface for Fault Gouge
wf_SEDT_tr/pt.dm Sediment 3D model
wf_MM_tr/pt.dm Ultramafics 3D model
wf_Z200_tr/pt.dm Safari North – Main lode high grade mineralisation – Zone 200
wf_Z205_tr/pt.dm Safari North – Footwall high grade mineralisation – Zone 205
wf_Z210_tr/pt.dm Safari North – Main lode low grade mineralisation – Zone 210
wf_Z215_tr/pt.dm Safari North – Footwall low grade mineralisation – Zone 215
wf_Z300_tr/pt.dm Safari South – Hangingwall high grade mineralisation – Zone 300
wf_Z310_tr/pt.dm Safari South – Footwall low grade mineralisation – Zone 310
17.2.7 Grade Estimation
Introduction
Resource estimation for the Safari copper and silver mineralisation was completed using OK within all Domains. Inverse Distance Squared estimates were also completed within these domains to allow comparison with the OK estimate.
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OK Estimate
Copper, silver grades and density values were estimated into the block model using OK after top cuts had been applied to the original 2m composite data.
No neighbourhood testing was carried out for this estimation (Table 17.2.7_1). Second and third passes with 2x and 3x multipliers for the search radii were applied, although the majority of the blocks were estimated in the first pass.
Table 17.2.7_1 Safari Deposit, Kapulo Copper Project Sample Search Parameters for OK Estimate
Search Orientation Search Radii Number of Samples Pass Semi- Domain Unit Major Minor Max / No. Major Semi-Major Minor Major Min Max Axis (m) Axis (m) Hole Axis (m) 200/300 Copper 1 120 60 25 8 24 8 -24° => 208° 45° => 145° 35° => 280° 210/310 Copper 1 120 60 25 8 24 8 200/300 Silver 1 100 50 25 8 24 8 -24° => 208° 45° => 145° 35° => 280° 210/310 Silver 1 100 50 25 8 24 8 200/300 Density 1 100 50 25 8 24 8 -24° => 208° 45° => 145° 35° => 280° 210/310 Density 1 100 50 25 8 24 8
Validation
A variety of validation checks were undertaken on the data prior to estimation to ensure that composite values and locations matched the original data in the database. After estimation was complete validation checks on the block model included:
Checks that the majority of blocks had filled with grade.
Correct assignment of density, classification, unit, domain and lithology information.
Volume comparison between the mineralisation wireframe and the mineralized units in the block model.
Visual inspection of estimated blocks against the informing composites and original drillhole data. (Figure 17.2.7_1)
Comparison plots of average informing composite grade and average block model grade by mineralized unit (Figure 17.2.7_2 to Figure 17.2.7_3).
Statistical comparison of 2m composites with declustered composites and block model grades.
Overall, there was a good comparison between the informing composite data input into the model and the resulting block grades. Blocks filled by direct assignment e.g. density, lithology and classification had all filled correctly.
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Figure 17.2.7_1 NS Long Section Safari North Block Model Grades & Informing Composites
Figure 17.2.7_2 Example of Comparison Plots of Cu Block Grades and Informing Composite Grades for Individual Units by 20m Northing Windows
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Figure 17.2.7_3 Example of Comparison Plots of Ag Block Grades and Informing Composite Grades for Individual Units by 20m Northing Windows
17.2.8 Resource Reporting and Classification
Introduction
The resource estimate for Safari update has been classified in accordance with the criteria laid out in the JORC Code (2004) and Canadian National Instrument 43-101. Indicated and Inferred Resources were defined based on data quality, data density and geological and/or grade continuity after detailed consideration of the JORC and CNI43-101 guidelines.
Criteria for Resource Categorisation
The classification of the Safari Resource was developed from the confidence levels of key criteria including drilling methods, geological understanding and interpretation, sampling, data density, data location, data quality, grade estimation and quality of the estimates. The current Safari Resource is reported as a combination of Indicated and Inferred material, depleted for surface artisanal workings, the key criteria are listed in Table 17.2.8_1 below. Figure 17.2.8_1 illustrates the classification applied to the resource model.
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Table 17.2.8_1 Safari Deposit, Kapulo Copper Project Confidence Levels of Key Categorisation Criteria
Item Discussion Confidence Drilling Techniques RC/Diamond – standard industry approach. High Logging Standard nomenclature applied and apparent high quality. High Drill Sample Recovery Recorded as good. High Sub-sampling Techniques Industry standard for both RC and diamond. High and Sample Preparation Quality of Assay Data Good internal and external QAQC check data for majority of the chemical High assays. Verification of Sampling QAQC analysis is within industry acceptable standards High and Assaying Location of Sampling Drillhole collars surveyed by DGPS and all drillholes have been surveyed High Points downhole for deviation Data Density and Nominal 40 x 40m overall spacing over whole deposit, Moderate Distribution Audits or Reviews Site drilling and sampling procedures reviewed by Coffey Mining. High Database Integrity No material errors identified High Geological Interpretation Further infill and extensional drilling may change the mineralisation shapes Moderate and the geological interpretation. Estimation and Estimates based on statistical and geostatistical analysis. Moderate Modelling Techniques Cutoff Grades Range of cutoff grades reported. High Mining Factors or No ore loss or dilution factored in. The effect of emulating SMU (change N/A Assumptions of support) has not been investigated.
Figure 17.2.8_1 Safari North Classified Block Model
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Grade Tonnage Reporting
The resources for the Safari deposits above a 0.3% copper cutoff are reported by mineralisation domain in Tables 17.2.8_2 and 17.2.8_3 below.
Table 17.2.8_2 Safari North Deposit, DRC Indicated/Inferred Resource Estimate Reported Above 0.3% Cu Cutoff Ordinary Kriged Estimate Using 2m Cut Cu% Composites Parent Cell Dimensions of 10mEW by 20mNS by 5mRL Depleted for Artisanal Surface Mining Categorised according to JORC Code (2004)
Grade Cu Metal Grade Ag Metal Classification Zone Mt (Cu%) (Tonnes) (Ag ppm) (kOz) HW 0.509 4.8 24,360 10 157 Indicated FW 0.451 1.8 7,950 3 48 Subtotal 0.959 3.4 32,310 7 205 HW 0.313 4.1 12,875 8 77 Inferred FW 0.693 1.3 9,075 3 65 Subtotal 1.006 2.2 21,950 4 142 Note: Figures have been rounded
Table 17.2.8_3 Safari South Deposit, DRC Indicated/Inferred Resource Estimate Reported Above 0.3% Cu Cutoff Ordinary Kriged Estimate Using 2m Cut Cu% Composites Parent Cell Dimensions of 10mEW by 20mNS by 5mRL Depleted for Artisanal Surface Mining Categorised according to JORC Code (2004)
Grade Cu Metal Grade Ag Metal Classification Zone Mt (Cu%) (Tonnes) (Ag ppm) (kOz) HW Indicated FW Subtotal HW 0.098 4.2 4,150 1 1.6 Inferred FW 0.294 1.1 3,095 1 8.6 Subtotal 0.392 1.8 7,245 1 10.2 Note: Figures have been rounded
The resources for the Safari North & South deposits above a 0.3% Cu cutoff are reported below, in Tables 17.2.8_4 by oxidation profile.
Adjacent to the artisanal workings are approximately 86,000t of artisanal spoil dumps with an average grade of ~4% Cu and ~3ppm Ag, calculated from global average grades from similar trenches at Shaba. These dumps are not classified and do not form part of the existing Resources. The potential quantity and grade of the dumps is conceptual in nature, and there has been insufficient exploration to define a Mineral Resource on the dumps and it is uncertain if further exploration will result in discovery of a Mineral Resource .
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Table 17.2.8_4 Safari North & South Deposits, DRC Indicated/Inferred Resource Estimate by Oxidation Reported Above 0.3% Cu Cutoff Ordinary Kriged Estimate Using 2m Cut Cu% Composites Parent Cell Dimensions of 10mEW by 20mNS by 5mRL Depleted for Artisanal Surface Mining Categorised according to JORC Code (2004)
Grade Cu Metal Grade Ag Metal Classification Zone Mt (Cu%) (Tonnes) (Ag ppm) (kOz) Oxide 0.212 3.8 8,000 5 34 Upr Trans 0.098 2.9 2,860 5 15 Indicated Lwr Trans 0.372 3.7 13,850 8 101 Fresh 0.272 2.7 7,600 6 55 Subtotal 0.959 3.4 32,315 7 205 Oxide 0.085 2.4 2,065 1 3 Upr Trans 0.061 2.4 1,455 1 3 Inferred Lwr Trans 0.308 2.5 7,825 3 32 Fresh 0.944 1.9 17,845 4 113 Subtotal 1.398 2.1 29,195 3 152 Note: Figures have been rounded
17.2.9 Comments and Recommendations
Drilling at the Safari prospects has defined an Indicated and Inferred Resource as defined in the above tables.
Further infill and extensional drilling is required to raise the level of confidence and extend the Inferred Resources.
17.3 Mineral Reserve Estimates
The Mineral Reserve estimates are based on the input parameters described in Section 18 of this report. Table 17.3_1 provides a summary of the Mineral Reserves that were determined for the Project, based on a 0.3% Cu cutoff. The Mineral Reserves pertain to the Shaba deposit only and no Mineral Reserves were determined for the Safari deposits at this stage. All stated Mineral Reserves are completely included within the quoted Mineral Resources.
This reserve estimate has been determined and reported in accordance with Canadian National Instrument 43-101, ‘Standards of Disclosure for Mineral Projects’ of December 2005 (the Instrument) and the classifications adopted by CIM Council in November 2010.
The reported Mineral Reserves have been compiled by Mr Harry Warries. Harry Warries is a Member of the Australasian Institute of Mining and Metallurgy and an employee of Coffey Mining. He has sufficient experience, relevant to the style of mineralisation and type of deposit under consideration and to the activity they are undertaking, to qualify as a Competent Person as defined in the JORC code.
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Table 17.3_1 Kapulo Mineral Reserves Shaba Deposit Based on a 0.3% Cu Cutoff 5% dilution applied at zero grade
Mineral Reserves Proven Probable Total Deposit Grade Insitu Metal Grade Insitu Metal Grade Insitu Metal Tonnes Tonnes Tonnes Cu Ag Cu Ag Cu Ag Cu Ag Cu Ag Cu Ag [Mt] [%] [g/t] [kt] [koz] [Mt] [%] [g/t] [kt] [koz] [%] [%] [g/t] [kt] [koz] Shaba - - - - - 3.6 3. 6 8.3 128.0 954.8 3.6 3.6 8.3 128.0 954.8
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18 OTHER RELEVANT DATA AND INFORMATION,
18.1 Mining
18.1.1 Mining Approach
The Project will involve a conventional open pit, selective mining exploitation method, employing a mining contractor.
Drilling and blasting will be performed on 5m high benches, with blasted material excavated in two discrete flitches, each nominally of 2.5m height.
The use of RC drilling, assays every 2.0m sample and interpretation of the results by mine geologists is the primary method of grade control envisaged for the Project. Further investigation is warranted with regards to optimising sample length and sample pattern to ensure appropriate statistical methodologies can be utilised to minimise dilution and maximise ore extraction.
A notional grade control drilling pattern of 8m x 5m was adopted for the study, with 120% of the expected ore zones assumed to be grade control drilled, to ensure sufficient overlap into adjacent low grade or waste such that ore is not missed. All RC holes are assumed to be drilled at a 50° angle.
The main mine production equipment that envisaged for the Project includes a 65t to 100t back hoe excavators and articulated haul trucks with a payload capacity of between 40t to 50t.
18.1.2 Geotechnical Input
All pit geotechnical work for the DFS was completed by George, Orr and Associates (Australia) Pty Ltd (GOA) and their work has been summarised in the report ‘Kapulo Copper Project, Democratic Republic of Congo: Shaba and Safari North Deposits: Geotechnical Evaluation for Open Pit Mining Feasibility Purposes’ dated May 2010 and a subsequent amendment in November 2010 as detailed in the memorandum ‘Kapulo Copper Project, Democratic Republic of Congo: Proposed Shaba Pit Final East Wall – Base Case Wall Design Amendment’ dated November 2010.
The geotechnical investigation was based on nine HQ3 size geotechnical boreholes.
Copper mineralisation at the Project is associated with the Kapulo Fault Zone (KFZ). This approximately north trending regional geological structure has an interpreted width of up to several hundreds of metres and dips to the west at angles ranging from around 60° to 70°.
The KFZ separates Banguelen Granite (in the east) from sedimentary rocks (sandstones and greywackes, with minor siltstone and shale) of the Kundelungu Group (in the west).
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Copper mineralisation appears to be associated with the north-north westerly striking, westerly dipping thrust faults. Locally, mineralisation dips to the west at around 55° to 65°, and plunges to the south at around 70°.
Alluvial materials (sand and clay) occur within the banks of the Kapulo River valley. The thickness of sediment is not known but is not expected to be greater than about 2m.
Figure 18.1.2_1 is an extremely simplified (schematic) geological cross-section, depicting the principal geological features (lithology and major thrust faults) interpreted to occur at the Shaba deposit.
Figure 18.1.2_1 Schematic Simplified Geological Cross-section of the Shaba Deposit
Thrust faulting has resulted in granitic bodies of varying dimensions (several metres to many tens of metres wide) being emplaced within the sediments forming the hangingwalls to mineralisation. This results in "mixed" and highly variable geological conditions. The mixed granitic and sedimentary rocks are referred to as "polymict breccias".
Interpretations made from the geotechnical drilling information indicate that:
Depths to the Base of Complete Oxidation (BOCO) at the Shaba deposit ranges from around 20m to 40m in the footwall and from around 30m to 50m in the hangingwall.
Depths to the Top of Fresh (TOFR) at the Shaba deposit range from around 50m to 80m (in the footwall) and from around 80m to 85m in the hangingwall.
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Inspection of the Rock Quality Designation (RQD) and Fracture Frequency (FF) indicates that generally "very poor" and "poor" quality ground conditions occur within the variably weathered and fractured rocks making up the hangingwalls, and occurring near surface in the footwalls at both deposits.
Similar predominantly "poor" quality rock conditions are interpreted to occur within the zone of sheared "brecciated" granite which extends eastwards from the sandstone – granite contact within the footwalls. This sheared rock envelope is interpreted to occur over a width of up to 20m to the east of the aforementioned contact. It is currently interpreted to represent the "sole" thrust fault.
Within the fresh unsheared granitic rocks occurring within the footwall, rock quality generally classifies as "good" to "very good".
Test results obtained for the Unconfined Compressive Strength (UCS) testwork indicated the following:
Kundelungu Group Sediments: Highly weathered rock 2MPa; Moderately weathered rock 50MPa; Fresh rock 150MPa.
Banguelen Group Granite and Mafic Intrusions: Highly weathered rock 5MPa; Moderately weathered rock 50MPa; Fresh rock >200MPa.
Pit wall stability is expected to be governed almost exclusively by the presence, attitude and shear strengths of geological structures such as faults, shears, and joints.
For the East Wall, GOA recommended that the final wall profile be mined as close as possible to the TOFR boundary, with berm crests located either on or immediately to the east of the TOFR boundary.
The pit wall slope parameters determined for Shaba are summarised in Table 18.1.2_1.
Table 18.1.2_1 Summary Pit Wall Slope Parameters
Sector Bearing Level [mRL] Parameter 60° & 10m high batters, 6m berms 1,260 - 1,140 0° – 45° 14m berms at 1,220 / 1,180 / 1,040 65° & 10m high batters, 8m berms and 1,140 - 1,080 East Wall 14m berms at 1,110 / 1,080 120° – 360° 75° & 10m high batters, 8m berms 1,080 - 980 14m berm at 1,020 45° – 120° 1,260 – 980 85° & 5m batters, 5m berms West Wall 65° & 10m high batters, 6m berms
Depressurised wall conditions were assumed, using the installation of sub-horizontal slope depressurisation boreholes.
Pre-split blasting is not expected to be required, providing good quality trim blasting (firing to a free face) is carried out.
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Groundwater is unlikely to have a significant influence on the overall pit slope stability and it is expected that the de-watering necessary to maintain reasonably dry operating conditions in the pit bottom will be adequate for slope stability purposes.
18.1.3 Hydrogeology and Hydrology Input
Hydrogeology
The hydrogeology aspects at the Project were assessed by Groundwater Resource Management Pty Ltd (GRM) in January 2010. Their findings are summarised in the report “Kapulo Project, Democratic Republic of Congo, Stage 2 Hydrogeological Investigation”.
The hydrogeological testwork was based on seven groundwater exploration holes with small scale permeability testing conducted. No high yielding groundwater intersects were observed during the groundwater exploration drilling, therefore no production bore drilling testing was undertaken.
Bulk permeabilities in the country rock around the Shaba deposit, away from the zones of mineralisation, is thought to be generally modest. Groundwater level measurements are highly variable ranging from 1,050mRL to 1,225mRL. This variation supports the assumption that permeabilities are modest, causing compartmentalised groundwater systems and high hydraulic gradients.
Permeabilities in the weathered zones at Shaba is believed to be low as the rock remains fairly massive with joints and breccia zones intersected in drill-core appearing to be re-healed with hematite.
Zones of high permeability are likely to be associated with the main north trending fault zone that runs through both the deposits and therefore associated with the mineralised zones. Thrust zones associated with mafic emplacement and en-echelon metasediment lenses in the ore zones may also have higher permeability.
The highest airlift yield obtained was 2L/s.
Notwithstanding the low predicted permeabilities, GRM believe that the hanging wall sediments have sufficient permeability to depressurise nominally 10m to 20m into the pit wall. However, it is recommended that this estimation is confirmed through monitoring of pit wall pressures during mining.
The footwall rocks are generally considered to be low permeability and therefore the depressurisation extent is likely to be restricted, i.e. less than 20m.
Given the predicted low pit in-flows for Shaba, dewatering should be able to be adequately managed by pumping from in-pit sumps.
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Hydrology
The hydrology and surface water aspects at the Project were assessed by Coffey Mining in December 2010. Their findings are summarised in the report “Kapulo Project, Democratic Republic Of Congo, Surface Water Management report”.
The Project is in the western foothills of the Bangulian Granite complex. The Kapulo Copper deposits outcrop in the steep north-south valley which forms the contact with the lower Makana and Kapona hills and the Marungu Plateau. The valley has an average elevation of ~1,200m above sea level. Few rivers drain into Lake Moero from the DRC. A minor ephemeral stream runs nearby the Kapulo site. Lake Moero is fed by the Luapula River, which along with Lake Moero, forms the international boundary between Zambia and the DRC.
The climate of the area is tropical wet and dry. The wet season begins towards the end of September and finishes at the end of April. The average rainfall, as indicated by mission records is 1,260mm, with a range of 800mm to 2,200mm.
The drainage management for the Project has been designed such that clean runoff water from areas upslope of the project infrastructure is diverted, where possible, around the site via diversion drains. Where drainage diversion is not possible (i.e. in the proposed plant area upstream of the Shaba Pit), runoff will be collected in a pump sumps and either used in dust suppression, as process water, or disposed of downstream if water quality guidelines are met. A general arrangement plan is provided in Figure 18.1.3_1 and shows the location of various project infrastructure, and the proposed drainage lines and plant site sump.
18.1.4 Mine Waste Geochemical Testwork Results
Coffey Mining undertook an assessment of the acid producing potential of the waste rock, the results of which are summarised in the memorandum “Mawson West, Kapulo Mine Waste Geochemistry Testwork Results”, dated 16 December 2010.
The waste was categorised by the three lithologies as listed below:
SST: Sandstone
BXGA: Granite Breccia
BXPG: Polymet breccia / Granite dom.
16 Samples were collected from which three composites were made based on lithology.
Based on the testwork results obtained, it was concluded that the waste material contains portions which are classified as having moderate to high acid forming potential. The relative proportion of this material to the total waste volume should be assessed to allow this material to be accounted for in the overall Acid Rock Drainage (ARD) management plan.
It is recommended that further column kinetic testing be performed, particularly on representative samples of Composite 2 (BXGA) waste material. The kinetic testing should be carried out over a longer time period and therefore it should be able to more accurately measure the extent of oxidation, rate of acid generation and water quality information.
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Figure 18.1.3_1 Surface Water Management - General Arrangement
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
18.1.5 Contract Mining
Mawson West obtained a quote from MCK Trucks sprl (MCK), a locally based mining contractor, in March 2011 to undertake the contract mining work at the Project.
Mawson West envisage that the blasting will be undertaken by an explosives supplier and a quote was obtained from AEL Mining Services.
Based on the above quotations the contract mining operating costs as shown in Table 18.1.5_1 were estimated.
Table 18.1.5_1 Kapulo Copper Project Summary Contract Mining Costs
Item Value [$/t mined] Monthly management fee 0.32 Drill & blast 0.57 Load & haul 2.14 Crusher Feed and Dayworks 0.08 Total 3.11
18.1.6 Pit Optimisation
Introduction
Pit optimisation studies have been undertaken using the resource models as developed by Coffey Mining and described in Section 17 as the basis for pit optimisation.
The Whittle Four-X pit optimisation software package was used for this work. The Whittle Four-X model development was carried out in Vulcan. Whittle Four-X deals with the amount of metal in a block, not grade. Knowing the tonnage and the metal in a block, Whittle Four-X can calculate the grade of a block. To that end, the metal content of a block was calculated using the grade estimate derived from the OK resource estimate.
Input Parameters
A mining dilution of 5% at 0.0g/t grade was added to the pit optimisation, whilst the mining recovery was set at 95%.
The input parameters adopted for the pit optimisation cover a wide range of disciplines and as a result a number of specialists have been involved in determining these parameters. The principal input parameters used in the pit optimisation and the specialists responsible for determining these parameters are listed in Table 18.1.6_1 below. All references to monetary values are denominated in United States of America dollars, unless specifically stated otherwise.
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Table 18.1.6_1 Kapulo Copper Project Source of Main Input Parameters
Input Parameter Source Commodity price Mawson West Treatment charges and Refining costs Mawson West Transport and marketing Mawson West Contract Mining Costs MCK and AEL contract quotation submission Owner’s mining associated costs Mawson West Metallurgical and Processing Sedgman Metals Engineering Services General and Administration cost Sedgman Metals Engineering Services, Mawson West Geotechnical George, Orr and Associates Hydrogeology Groundwater Resource Management Governmental Mawson West
A flat copper price of US$2.50/lb and a silver price of US$12.50/oz were adopted for the pit optimisation work that was carried out for the Project.
A Net Smelter Return (NSR) of 97% and 94% was used for Cu and Ag respectively.
Marketing and transport costs were estimated at US$1,250/t Cu.
The average mining costs were based on an initial mining quote from contractors and were estimated at US$2.58/t mined. It is noted that updated contractor quotes were obtained subsequent to the pit optimisation and the updated mining costs are shown in Table 18.1.4_1.
Miscellaneous mining costs such as mine supervision and de-watering were estimated by Mawson West and were included in the project general and administration cost.
Grade control costs were estimated by Mawson West at US$0.30/t based on a RC drilling rate of $30/m per meter and assaying cost of $10 per assay.
The processing costs that were determined for the Project were supplied by Sedgman Metals Engineering Services (Sedgman) and were estimated at $24.30/t.
The recoveries were supplied by Sedgman and are summarised in Table 18.1.6_2.
Table 18.1.6_2 Kapulo Copper Project Summary Processing Recoveries
Material Type Recovery (%) Cu 75 Weathered Ag 70 Cu 93 Transitional and Fresh Ag 92
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
The Project general and administration costs were determined by Sedgman and Mawson West and were estimated at $8.0 million per annum. In addition, a US$1 million per annum allowance was made for sustaining capital.
Mawson West advised that no government royalty was applicable.
The pit wall slope parameters determined for the Project were based on the geotechnical parameters as described in Section 18.1.2. In summary, the overall slope angles, which incorporate a ramp system, that were adopted for the pit optimisation ranged from 37º to 49º.
A summary of the principal economic input parameters used in the pit optimisations are summarised in Table 18.1.6_3 below.
Table 18.1.6_3 Kapulo Copper Project Summary Whittle Four-X Input Parameters
Item Unit Value Mill throughput Mtpa 0.5 Cu US$/lb 2.50 Commodity price Ag US$/oz 12.50 Cu % 97 NSR Ag % 94 Marketing and transport US$/t Cu 1,250 State government royalty % Nil Processing cost $/t 24.30 General and Administration $/t mill feed 8.0 Sustaining capital US$/t mill feed 1.0 Average mining cost $/t mined 2.58 Grade control $/t mill feed 0.30 Cu % 75 Weathered Ag % 70 Processing recovery Cu % 93 Transition/Fresh Ag % 92 Mining dilution (at zero grade) % 5 Mining recovery % 95 Overall pit wall slope angle (inclusive of a ramp system) degrees 37 - 49
Pit Optimisation Setup
The pit optimisations were carried out for a wide range of Cu and Ag prices, from as low as US$1.00/lb to a maximum of $5.00/lb for Cu and US$5.00/oz to a maximum of US$25/oz for Ag. Figure 18.1.6_1 displays the pit size in relationship to the copper and silver price, for the pit optimisation based on the Indicated Resources only.
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Figure 18.1.6_1 Commodity Price vs Mill Feed - Based on Indicated Resources only Ag Price (US$/oz) - 5.00 10.00 15.00 20.00 25.00 30.00 5,000,000 4,500,000 4,000,000 3,500,000 3,000,000 2,500,000 2,000,000
Mill FeedMill (t) 1,500,000 1,000,000 500,000 0 - 1.00 2.00 3.00 4.00 5.00 6.00
Cu Price (US$/lb)
The Whittle Four-X financial analysis was carried out using the following base assumptions and parameters:
Mill throughput: 0.5Mtpa.
Mill limiting: i.e. sufficient waste is removed each period to enable the required milling rate to be maintained.
Discount rate: 10%.
Base case commodity price : Cu: US$2.50/lb : Ag: US$12.50/oz
Three cashflows were produced for each analysis:
Undiscounted Operating Cashflow.
Best Case Discounted Operating Cashflow – Each incremental pit is removed prior to advancing to the next adjacent incremental pit. The cashflow schedule is the equivalent of multiple pushbacks.
Worst Case Discounted Operating Cashflow – Each bench is mined out prior to moving to the next bench, using the optimisation block height as the default bench height. The cashflow schedule is the equivalent of top down ‘flat’ mining.
An actual mining schedule will most likely lie between the two extremes of Worst Case and Best Case as described above.
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The cashflows, as described above, are exclusive of any capital expenditure or Project start- up costs and should be used for pit optimisation comparison purposes only. No Net Present Value (NPV) can be derived from these cashflows.
Pit Optimisation Results
Based on Indicated Resources only and at a copper price of US$2.50/lb and a silver price of US$12.50/oz, the optimum pit shell, based on the maximum un-discounted cash flow, is pit shell 12. Pit shell 12 contains some 2.9 million tonnes of ore at a grade of 4.5% Cu and 11.3g/t Ag, for approximately 118,000t of recovered copper and 904,000 ounces of recovered silver. Some 37.2 million tonnes of waste are contained within the pit shell with a stripping ratio of 12.9:1. The undiscounted operating cashflow, exclusive of capital and start up costs, is $259 million. The Worst Case discounted cashflow is $181 million, whilst the Best Case discounted cashflow is $201 million. To ascertain the likely discounted cashflow derived from a realistic mine production schedule, the average discounted cashflow was calculated and is $191 million. The cash operating cost derived from the pit optimisation is $1.51/lb, inclusive of silver credits.
Table 18.1.6_4 provides a summary of the pit optimisation results, with Figure 18.1.6_2 displaying the pit optimisation results graphically.
Figure 18.1.6_2 Summary Pit Optimisation Results - Based on Indicated Resources only 70.0 300.0
60.0 250.0
50.0 200.0 40.0 150.0 30.0 Tonnes [Mt] Tonnes
100.0 [M$] FlowCash 20.0
10.0 50.0
0.0 0.0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 Pit Waste Mill Feed Undiscounted operating cash flow Discounted cash flow - Worst
Discounted cash flow - Best Discounted cash flow - Average
Kapulo Copper Project, DRC – MINEWPER00482AC National Instrument 43-101 Technical Report – 30 June 2011 Coffey Mining Pty Ltd
Table 18.1.6_4 Kapulo Copper Project Summary Whittle Four-X Pit Optimisation Results Based on Indicated Resources Only
Mill Feed Cu Metal Ag Metal Operating Cashflow Cashcost Total Strip Pit Base Waste Grade Discounted @ 10% Ag Credits Pit Material Ratio Tonnes Insitu Rec Insitu Rec Undisc. Cu Ag Worst Best Avg. Excl Incl (mRL) (Mt) (Mt) (w:o) (Mt) (%) (g/t) (kt) (koz) US$M [US$/lb] 1 1120 1.5 1.1 2.5 0.43 6.34 13.2 27.3 22.9 174 150 72.0 66.4 66.4 66.4 1.11 1.07 2 1060 6.0 5.0 4.8 1.05 5.41 12.2 56.4 49.2 389 346 143.4 122.3 124.5 123.4 1.22 1.18 3 1050 8.0 6.6 4.9 1.35 5.13 11.5 69.0 60.6 473 422 172.6 142.4 146.2 144.3 1.25 1.21 4 1030 10.7 9.1 5.8 1.59 5.02 11.6 79.5 70.2 561 503 195.2 156.8 161.9 159.4 1.28 1.24 5 1010 15.2 13.3 7.2 1.86 4.98 11.8 92.2 82.0 670 603 220.9 172.1 179.3 175.7 1.32 1.28 6 1010 15.9 14.0 7.3 1.92 4.93 11.7 94.4 84.0 685 617 224.6 173.7 181.5 177.6 1.33 1.29 7 1000 16.8 14.9 7.6 1.96 4.93 11.7 96.2 85.7 703 633 227.7 175.1 183.4 179.3 1.34 1.30 8 970 25.0 22.6 9.4 2.40 4.69 11.3 112.2 100.6 829 749 248.3 183.1 195.7 189.4 1.42 1.38 9 970 25.7 23.3 9.5 2.44 4.66 11.3 113.6 101.8 840 759 249.7 183.2 196.4 189.8 1.43 1.39 10 970 26.3 23.8 9.7 2.47 4.64 11.2 114.5 102.7 848 767 250.4 183.1 196.7 189.9 1.44 1.39 11 960 31.4 28.7 10.9 2.63 4.60 11.2 120.7 108.5 903 817 253.8 182.7 198.5 190.6 1.48 1.44 12 940 40.1 37.2 12.9 2.89 4.54 11.3 131.0 118.0 997 904 258.6 180.7 200.9 190.8 1.55 1.51 13 930 47.3 44.2 14.3 3.11 4.48 11.4 138.8 125.3 1,085 984 260.2 177.1 201.4 189.2 1.60 1.56 14 920 51.0 47.8 15.1 3.19 4.47 11.6 142.2 128.4 1,131 1,027 259.6 174.7 200.9 187.8 1.63 1.58 15 910 56.6 53.3 16.1 3.33 4.43 11.8 147.0 132.9 1,200 1,090 257.0 170.0 199.3 184.6 1.67 1.62 16 910 57.1 53.8 16.1 3.35 4.41 11.8 147.5 133.4 1,204 1,094 256.4 169.0 199.0 184.0 1.67 1.63 17 910 57.8 54.4 16.2 3.36 4.41 11.8 147.9 133.8 1,211 1,101 255.9 168.3 198.7 183.5 1.68 1.63 18 910 59.7 56.3 16.7 3.39 4.41 11.9 149.1 134.9 1,232 1,120 254.1 166.2 197.7 181.9 1.69 1.65 19 910 59.8 56.4 16.7 3.39 4.41 11.9 149.2 135.0 1,234 1,122 253.9 165.9 197.6 181.8 1.69 1.65 20 910 60.5 57.1 16.8 3.40 4.41 11.9 149.6 135.3 1,238 1,125 253.3 165.3 197.3 181.3 1.70 1.65 21 910 60.7 57.3 16.9 3.41 4.41 11.9 149.6 135.4 1,238 1,126 252.9 164.8 197.0 180.9 1.70 1.65 22 910 60.8 57.4 16.9 3.41 4.40 11.9 149.7 135.5 1,239 1,126 252.5 164.4 196.8 180.6 1.70 1.65 23 910 61.1 57.7 16.9 3.42 4.40 11.9 149.9 135.6 1,240 1,128 252.1 164.0 196.6 180.3 1.70 1.66 24 900 65.6 62.2 18.1 3.45 4.41 12.1 151.9 137.5 1,280 1,164 245.9 158.4 193.4 175.9 1.74 1.69 25 900 65.6 62.2 18.1 3.45 4.41 12.1 151.9 137.5 1,280 1,164 245.9 158.4 193.4 175.9 1.74 1.69 26 900 66.1 62.7 18.2 3.46 4.41 12.1 152.0 137.6 1,281 1,165 244.8 157.5 192.8 175.2 1.74 1.69 27 900 66.2 62.7 18.2 3.46 4.41 12.1 152.0 137.6 1,281 1,165 244.6 157.3 192.7 175.0 1.74 1.69 28 900 66.2 62.7 18.2 3.46 4.41 12.1 152.0 137.6 1,281 1,165 244.6 157.3 192.7 175.0 1.74 1.69 29 900 66.2 62.8 18.2 3.46 4.41 12.1 152.0 137.6 1,281 1,165 244.5 157.2 192.7 175.0 1.74 1.69 30 900 66.5 63.0 18.3 3.46 4.41 12.1 152.0 137.6 1,281 1,165 243.9 156.8 192.4 174.6 1.74 1.70
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18.1.7 Mine Design
Pit Design
Figure 18.1.6_2 shows that the pit shell versus cash flow curve is very flat near the maximum cash flow inflection point, which indicates that selecting any of the pit shells that lie on either side of the inflection point will generate similar cash flows. Pit shell 12 was selected for the detailed life of mine pit design work.
Many starter shells comprised a number of pods and shared a common FW. As such, a single starter shell was selected with pit shell 10 being used as a guide for the design of the starter pit.
The pit designs were based on the pit wall slope configuration as described in Section 18.1.2.
The ramp system was based on a 16m wide ramp at a 10% gradient.
The starter pit reaches a depth of 1,020mRL, whilst the life of mine pit design reaches a depth of 980mRL.
A minimum mining width of 20m was assumed. This width suits the backhoe excavator loading method and 50t articulated trucks turning circle of 10m.
Table 18.1.7_1 provides a summary of the material breakdown as contained within the starter pit and final pit designs.
Table 18.1.7_1 Kapulo Copper Project Summary Material Breakdown by Pit Design
Reserves
Total Strip Proved Probable Total Waste Pit Material Ratio Grade Grade Grade Tonnes Tonnes Tonnes Cu Ag Cu Ag Cu Ag [Mt] [Mt] [w:o] [Mt] [%] [g/t] [Mt] [%] [g/t] [Mt] [%] [g/t] Starter pit 26.2 23.5 8.8 - - - 2.7 3.84 7.44 2.7 3.84 7.44 Final pit 12.5. 11.6 12.7 - - - 0.9 2.79 7.42 0.9 2.79 7.42 Total 38.7 35.1 9.8 - - - 3.6 3.58 8.30 3.6 3.58 8.30
Figure 18.1.7_1 shows the starter pit and final pit design as developed for the Project.
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Figure 18.1.7_1 Plan View of Starter Pit and Final Pit Design
Waste Dumps
The waste dumps have been designed to Western Australian standards and the parameters used are:
Face slope 22º
Bench height 20m
Berm width 6m
Overall slope 20º
The waste dump capacities have been based on a swell factor of 30%. The top of the waste dump is at 1,235mRL.
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.
Figure 18.1.7_1 shows the waste dumps, pit designs and associated roads and other site infrastructure.
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Figure 18.1.7_1 Mine Infrastructure Layout
The waste mined during the pre-production period will be utilised to build the ROM pad, the tailings storage facility (TSF) starter dam, raw water dam (RWD) and other infrastructure items, such as haul roads, as required.
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18.1.8 Mine Production Schedule
The mine production schedule was based on the pit designs as described in Section 18.1.7.
The scheduling periods adopted for the mine production schedule comprises of annual increments, although pre-production has been estimated at approximately six months.
The mine production schedule was developed using Microsoft Excel.
Scheduling was carried out on a bench by bench basis for the starter pit and final cutback.
The following constraints were set as a target for the mine production schedule:
Ore processing rate Oxide 0.385Mtpa Sulphide 0.5Mtpa
Maximum total material movement 8.0Mt
Vertical mining advance rate 60m
Pre-production material requirements Mill Feed 0.03Mt Waste 3.0Mt
The pre-production waste requirements comprise the following infrastructure items:
TSF (starter dam) 0.28Mm³
Raw water dam 0.46Mm³
ROM pad 2.0 Mm³
In order to reduce pre-production capital cost, it was assumed that only part of the ROM pad will be built, ready for plant commissioning and that the ROM pad will be extended with suitable waste during ongoing mining after plant commissioning.
The Project produces an average of 15.8kt of Cu per year, ranging from a minimum of 14.2kt in Year 6 to a maximum of 17.3kt in Year 2. In addition, the Project produces an average of 78.0kozof Ag per year, ranging from a minimum of 52.7koz in Year 1 to a maximum of 85.7koz in Year 7.
The average total material movement is approximately 6.7Mtpa, with a maximum mining rate of 7.8Mtpa in Year 1, Year 2 and Year 3 and a minimum annualised mining rate of 4.8Mtpa in Year 5.
In order to maximise the mine production fleet efficiency, the mining contractor will complete mining in six years with excess ROM being stockpiled. The ROM stockpile will reach a maximum capacity of 0.7Mt.
Table 18.1.8_1 summarises the mine production schedule that has been developed for the Project.
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Table 18.1.8_1 Kapulo Copper Project Summary Mine Production Schedule
Year Item Unit Pre- Total 1 2 3 4 5 6 7 8 Prod Total Material [Mt] 3.1 7.8 7.8 7.8 5.2 4.8 2.1 38.7 Waste [Mt] 3.1 7.0 7.3 7.1 4.6 4.4 1.6 35.1 Strip Ratio [w:o] NA 8.7 13.2 10.1 7.9 11.3 3.0 9.8 [Mt] 0.03 0.81 0.55 0.70 0.58 0.39 0.52 3.58 Mill Feed [% Cu] 2.4% 4.1% 3.7% 3.5% 3.6% 2.3% 3.8% 3.58% Mined [g/t Ag] 2.3 6.9 8.3 8.9 10.3 5.5 9.7 8.30 [Mt] 0.42 0.50 0.50 0.50 0.50 0.50 0.50 0.16 3.58 Mill Feed [% Cu] 4.4% 3.8% 3.6% 3.6% 3.3% 3.1% 3.4% 3.3% 3.58% Processed [g/t Ag] 6.44 7.60 8.21 9.47 8.81 8.15 8.89 9.01 8.30 [Mt] 0.03 0.41 0.46 0.66 0.74 0.63 0.66 0.28 Stockpile [% Cu] 2.4% 3.7% 3.6% 3.5% 3.5% 2.9% 3.4% 3.3% Balance [g/t Ag] 2.30 7.10 8.00 8.83 9.55 7.66 8.92 8.95 Insitu [kt] 18.5 19.2 17.8 18.2 16.5 15.7 17.2 5.1 128.0 Copper Rec [kt] 16.7 17.3 16.0 16.4 14.9 14.2 15.5 4.6 115.5 Insitu [koz] 87.8 122.2 132.0 152.3 141.6 131.1 142.9 44.9 954.8 Silver Rec [koz] 52.7 73.3 79.2 91.4 84.9 78.6 85.7 27.0 572.9
The mine production schedule is presented graphically in Figure 18.1.8_1.
Figure 18.1.8_1 Summary Mine Production Schedule 9.0 8.0 7.0 6.0 5.0 4.0
Tonnes [Mt] Tonnes 3.0 2.0 1.0 0.0 Pre -Prod. Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Period [Years] Waste Mill feed mined
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The processing schedule is shown in Figure 18.1.8_2.
Figure 18.1.8_2 Summary Processing Schedule 0.60 10.00
9.00 0.50 8.00
7.00 0.40 6.00
0.30 5.00 Grade
Tonnes [Mt] Tonnes 4.00 0.20 3.00
2.00 0.10 1.00
0.00 - Pre -Prod. Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Period [Years]
Mill Feed Cu grade [%] Ag grade [g/t]
18.2 Proposed Processing Operations
The process plant has been designed for campaign processing of either 0.385Mtpa of oxide material or 0.5Mtpa of sulphide material.
The flowsheet was based on conventional comminution, flotation and filtration processes. The plant design is based upon crusher feed grades of 5.40% Cu for the oxide material and 4.37% Cu for the sulphide material, these values being derived from the head assays of the completed metallurgical testing. The design criteria used are summarised in Table 18.2_1.
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Table 18.2_1 Summary Process Design Criteria
Value Description Unit Oxide Sulphide Nominal Annual Dry Ore Throughput t/a 385,000 500,000 Ore Grade % Cu 5.40 4.37 Crushing Circuit Type 2 Stage 2 Stage Crushing Crushing Throughput t/h 117 152 Crushing Availability % 75 75 Grinding Circuit Type Single stage ball mill Single stage ball mill Grinding Grinding Throughput t/h 48.1 62.5 Grinding Circuit Availability % 93.1 93.1 Roughing stage cells No. 5 5 Rougher cell volume m³ each 16 16 Rougher residence time Min 42.6 30.2 Cleaner 1 stage cells No. 6 6 Flotation Cleaner 1 cell volume m³ each 3 3 Cleaner 1 residence time Min 42.8 20.4 Cleaner 2 stage cells No. 4 4 Cleaner 2 cell volume m³ each 1.5 1.5 Cleaner 2 residence time min 38.3 14.7 Concentrate Thickening Specific settling rate m²/t/h 0.15 0.25 Concentrate Filtration Cake Flux kg/m².h 100 300
The general plant design philosophy has been to provide a cost effective processing solution, whilst maintaining high levels of reliability, operability and maintainability. In order to achieve this, the following strategies were adopted;
The crushing plant will operate on a 365 days/year, 12h/day operating cycle, with a design availability within this period of 75% for a nominal throughput of 117t/h on oxide ore and 152t/h on sulphide ore.
Downstream of the crushing circuit, the plant will operate on a 365day/year, 24h/day operating cycle with a design availability of 91.3% for a nominal combined ore throughput of 48.1 dry t/h on oxide crusher feed and 62.5 dry t/h on sulphide crusher feed.
The design basis assumes a moderate level of instrumentation and automation to minimise the operator requirement without introducing undue complexity and expense.
Adherence to well-proven and conservative design practice appropriate to the copper flotation industry.
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The processing facility design is based upon the supply and installation of new processing equipment and comprises the following unit operations:
Sulphide crusher feed secondary crushing and reclaim.
Single stage closed circuit ball milling.
Slurry conditioning.
Flotation.
Concentrate thickening and filtration.
Tailings thickening and disposal.
Reagent storage and distribution.
Water supply, recovery and distribution services.
Air supply and distribution services.
18.2.1 Crushing Circuit
The crushing circuit will consist of a primary jaw crusher, a secondary cone crusher, a vibratory grizzly, an apron feeder, a double deck screen, associated covered conveyors, dust extraction system and two self cleaning tramp metal magnets. The crushing circuit will operate 12 hours per day for seven days per week.
Mill feed will be delivered to the Run of Mine (ROM) pad via articulated dump trucks with an approximate capacity of 40t. The mill feed will either be stockpiled on the ROM or direct tipped into the ROM bin as required. Stockpiled mill feed will be reclaimed by FEL and fed to the ROM bin.
The ROM bin will be fitted with a heavy duty fixed grizzly, with an aperture of 700mm to remove oversize material from the crusher feed. A static rock breaker will be installed to help remove oversize rocks which may jam in the grizzly.
The ROM bin will discharge via a variable speed apron feeder (2,200mm x 9,000mm), which will convey the crusher feed to the jaw crusher. Minor spillage from the apron feeder will be collected in a dribble chute which will direct the spillage to the combined crusher discharge conveyor. Crusher discharge will gravitate, along with dribble conveyor discharge, to the combined crusher discharge conveyor.
The jaw crusher, nominally 1,100mm x 850mm, 160kW, will have a closed side setting (CSS) of approximately 70mm. The jaw crusher discharge conveyor will transfer the crushed mill feed to the double deck screen from which both the oversize and intermediate fractions will be fed into the secondary cone crusher feed bin, which will have approximately 30t capacity. The crusher feed will be reclaimed from the secondary feed bin via a vibratory feeder ahead of the secondary cone crusher which will have an operating CSS of 16mm. The secondary cone crusher discharge will be combined with the jaw crusher discharge as feed to the double deck screen.
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Both the cone crusher and vibrating screen will be protected against ferrous metal damage by two self cleaning tramp metal magnets.
A dust extraction system will be installed to achieve dust control on two fronts:
All conveyor transfer points will be enclosed and ducted to a common baghouse. Collected dust will be discharged to the stockpile feed conveyor.
The installation of dust suppression sprays above all feed and surge bins.
To assist with dust extraction as well as minimising water ingress during the wet season roofing will be provided for the crushers and screening areas. In addition, all conveyors and fine ore stockpile will be covered.
18.2.2 Crushed Ore Stockpiles
The mill feed will be crushed to -15mm before being conveyed to a covered fine crusher feed stockpile from which it will be reclaimed for feeding to the comminution circuit. The screen undersize, at -15mm, will be conveyed to the fine crusher feed stockpile from which it will be reclaimed and fed to the comminution circuit.
18.2.3 Milling
Crushed mill feed will be reclaimed from the fine crusher feed stockpile by one of two variable speed vibratory feeders which will transfer the crusher feed to the ball mill via the ball mill feed conveyor. A weightometer fitted to the ball mill feed conveyor will allow measurement of the feed rate for metallurgical accounting purposes. Ground crusher feed will be discharged from the ball mill via a trommel, the undersize of which will gravitate to the cyclone feed sump. Trommel oversize will be conveyed to scats bin for discard.
The ball mill will operate in closed circuit with a set of hydrocyclones, mill discharge will be combined with make-up water, and be pumped to the cyclones for classification. Target
cyclone overflow sizing will be P 80 106µm for the oxide crusher feed and P 80 75µm for the sulphide crusher feed. The cyclone overflow will gravitate to the rougher flotation feed for mill feed conditioning. Cyclone underflow will be gravitate to an underflow box from where it will be primarily returned to the ball mill feed. Provision will be made for a bleed stream to be extracted from this box to supply a future unit flotation cell to be located in the grinding circuit. This will only be operated during sulphide crusher feed processing and will produce a concentrate of product grade. This will supply the concentrate thickener directly with the underflow being returned to the ball mill feed.
A hoist and a ball kibble will be provided for loading of grinding media into the ball mill.
The ball mill will be approximately 4.0m diameter by 6.0m effective grinding length (EGL) and fitted with a 1.5MW motor.
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Key parameters for the Project’s grinding circuit are presented in Table 18.2.3_1.
Table 18.2.3_1 Kapulo Grinding Circuit Key Parameters
Value Criteria Units Oxide Ore Sulphide Ore Throughput Mt/a 0.385 0.500 t/h 48.1 62.5 Availability % 91.3 91.3
Feed Size F 80 mm 12.0 12.0
Target Product Size P 80 µm 106 75 RWi kWh/t 6.8 13.1 Ore Characteristics BWi kWh/t 16.9 17.4 Ai g 0.287 0.572
18.2.4 Flotation
In the design of the flotation circuit, it has been assumed that the oxide mill feed will be processed prior to the sulphide mill feed and will use the same circuit. The circuit will include some equipment that will become redundant once all of the oxide mill feed has been processed.
The oxide mill feed primary cyclone overflow will gravitate to the rougher conditioners where it will be conditioned with NaHS for 15 minutes prior to being conditioned for five minutes with PAX and Aeropromoter 407 prior to being pumped to the rougher flotation cells.
The sulphide mill feed primary cyclone overflow will gravitate to the rougher conditioners where it will be conditioned with NaHS for five minutes with Aerophine 3418A prior to being pumped to the rougher flotation cells.
A bank of five 16m³ tank cells will be installed as the roughers giving a total residence time of approximately 42.5 minutes for the oxide mill feed and approximately 30.0 minutes for the sulphide mill feed. The rougher tails will be pumped to the tailings thickener, while the rougher concentrate will be pumped to the Cleaner 1 conditioner tank where it will be conditioned with NaHS for 5 minutes.
The Cleaner 1 concentrate will be fed to the Cleaner 2 conditioner tank for five minutes, conditioning with NaHS prior to being sent to the Cleaner 2 flotation cells from which the final copper concentrate will be produced and sent to the concentrate thickener. The Cleaner 2 tail will be pumped to the Cleaner 1 conditioner tank, while the Cleaner 1 tail will be pumped to the rougher conditioner tank.
The Cleaner 1 circuit will consist of six 3m³ conventional cells, with a total residence time of approximately 43 minutes for the oxide mill feed and 20 minutes for the sulphide mill feed. The Cleaner 2 circuit will consist of four 1.5m³ conventional cells, with an approximate residence time of 38 minutes for the oxide mill feed and 15 minutes for the sulphide mill feed.
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Collector, i.e. Aerophine 3418A for the sulphide mill feed and AP407 plus PAX for the oxide mill feed, NaHS for the oxide mill feed only and frother will be added stage-wise throughout the rougher and cleaner circuits.
An allowance will be made for the provision of a fully automated Controlled Potential Sulphidisation (CPS) system, complete with automatic addition of NaHS to optimise the recovery of the oxide copper minerals.
An in-stream analysis system will be installed to help maintain the efficient operation of the processing plant.
18.2.5 Concentrate Handling
The final concentrates will be thickened to approximately 50% to 55% solids for the oxide mill feed concentrate and to approximately 55% to 60% solids for the sulphide mill feed concentrate in a thickener, prior to being pumped to the filter stock tank. The filter stock tank will be agitated and offer a residence time of 24 hours when processing oxide mill feed and 15 hours when processing sulphide mill feed.
Concentrates will be withdrawn from the stock tank and will be pumped to a plate and frame filter operating in automated batch mode for final dewatering. The filter cake will be discharged to a concentrate bagging facility for loading into bulka-bags each of 2t capacity in preparation for transportation.
The filtrate will be pumped back to the concentrate thickener feed box and the concentrate thickener overflow will gravitate to the process water dam for re-use in the processing plant.
A hi-rate thickener, of 7.0m diameter, will be used for the concentrate dewatering. Flocculant will be added to the thickener feed box to assist with settling rate and overflow clarity.
An expandable plate and frame filter with 33 plates each 1.5m by 1.5m, will be installed for concentrate filtration. It is anticipated that the filtration circuit will operate for approximately 20 hours per day. Filtered concentrate will be discharged to a reversible conveyor where it will be normally directed to a concentrate bin of the concentrate bagging facility. During periods where the bagging facility is unavailable, the filtered concentrate will be directed to a bulk stockpile for later re-feeding to the concentrate bin by a Front End Loader (FEL).
The concentrate bagging system will incorporate a concentrate storage bin, a bag feeder, a bag filling conveyor, a weigh scale and a roller conveyor. Empty bulk bags will be loaded in to a support frame over the weigh scale and be filled to the predetermined amount (nominally 2,000kg). Filled bags will be transported down the roller conveyor for transfer to the storage area. Transfer will be achieved by an overhead gantry hoist contained within the filter building which will allow loading of the bags on to trucks for shipment.
18.2.6 Reagents
A number of reagents and consumables will be used at Kapulo. Table 18.2.6_1 summarises the major process inputs.
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Table 18.2.6_1 Kapulo Copper Project Reagents and Consumables
Usage Rate (kg/t) Annual Usage (t) Reagent Area Used Packing Distribution Oxide Sulphide Oxide Sulphide Grinding Media Ball Mills 1t Ball Drum Batch by Kibble 1.47 2.25 565.2 1,124.0 PAX Flotation Bulk Bag 20.0% w/w Solution 0.75 0.00 288.8 0.0 Aeropromoter 407 Flotation 205L Drum Neat 0.45 0.00 173.3 0.0 Aerophine 3418A Flotation 205L Drum Neat 0.00 0.06 0.0 30.0 Frother Flotation 205L Drum Neat 0.04 0.04 15.4 20.0 NaHS Flotation Bulk Bag 30.0% w/w Solution 5.42 0.00 2,086.7 0.0 Flocculant Concentrate Handling & Tailings 25kg Bag 0.25% w/w Solution 0.03 0.03 11.6 15.0
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Reagents will be stored in a covered area and accessed by fork lift. The reagent mixing will be in covered tanks, outside the storage area. The reagents will be transferred to holding tanks for dosing throughout the processing plant.
18.2.7 Tailings
A tailings thickener, of 11.0m diameter, will be installed to recover water from the flotation tailings prior to being pumped to the tailings storage facility.
Slurry will be pumped from the flotation tails sump to the thickener feed well. Mixed flocculant solution will be added to aid in the thickening process prior to the slurry entry to the thickener feed well. Target thickener underflow density is 55% w/w solids for the oxide mill feed and 60% w/w solids for the sulphide mill feed. Thickener underflow will be pumped to the tailings storage facility located to the south of the processing plant. Thickener overflow will gravitate to the process water pond for reuse within the circuit.
It is anticipated that water will decant from the tailings dam to a decant catchment pond located adjacent to the dam. This water will be pumped back to the process water pond by one of two decant pumps located at the decant pond.
18.2.8 Air Supply
General purpose plant and instrument air will be provided by two air compressors operating at 750kPa with associated filters, refrigerative air dryers and associated receivers. These compressors will also provide air to the concentrate filter via dedicated receivers.
Low pressure air for flotation will be provided by two air blowers (one duty and one standby).
18.2.9 Water Services
A raw water dam will be constructed in the valley immediately east of the process plant. The dam wall will be waste rock fill and will have a storage capacity of 347,000m³ of water. A natural spring feeds into the dam area and additional supplies will be pumped from the Lukinda River to the dam during start-up and extended dry periods.
Raw water will be pumped from the raw water dam and distributed to the plant for water treatment, plant feed, low pressure gland water, fire water, reagent makeup and process water make up. Gland water for high pressure applications will be supplied from the raw water system and boosted by dedicated pumps fitted to the gland water lines to supply the majority of the slurry pumps.
The process water dam (2,000m³) will collect water from the concentrate thickener overflow, tailings thickener overflow and tailings storage facility decant pond as well as make up water from the raw water dam. The process water will be used as dilution water, hose up water and spray water throughout the plant.
A water treatment plant, fed from the raw water system will be used to generate potable water for distribution to the plant and associated infrastructure including the camp, offices, and ablutions as well as for safety showers.
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18.3 Tailings Storage and Raw Water Dam
A Tailings Storage Facility (TSF) and Raw Water Dam (RWD) embankment have been designed for the Project by Coffey Mining as detailed in the report MINEWPER00482AD-AI TSF RWD Design Report RevC.docx.
The tailings are to be discharged at between 55% solids (oxide ore) and 60% solids (fresh rock ore).
Tailings from the process operation will be discharged from the plant to a valley type TSF, located to the south west of the process plant. The RWD site is proposed in the side valley to the east of the processing plant. The RWD and TSF sites and design concepts were selected to maximise the use of existing topography and available mine waste material.
The designs were developed based on information supplied by Mawson West, tailings physical characterisation testwork, and a geotechnical site investigation that included borehole drilling, test pit excavation, sampling, and laboratory testwork.
The tailings have been assessed as Potentially Acid Forming (PAF) and further analysis of the geochemistry of the tailings is required in the detailed design stage.
The TSF embankment will be constructed in stages using mine waste to suit tailings storage requirements and as mine waste becomes available for construction. At full size, the TSF will cover an area of approximately 39 hectares with an embankment crest level of RL1278m. The design of the TSF includes a pontoon mounted pump allowing water to be pumped to the process plant for reuse. The tailings will be discharged via spigots from the embankment crest, ensuring the beach slopes towards the western end of the TSF, maintaining a supernatant pond away from the embankment. The conceptual closure design includes placement of a mine waste cover (nominal 1m thick) and construction of surface water management features.
The RWD embankment will also be constructed using mine waste and includes a moisture conditioned and roller compacted clay core, rip rap rock armour of the upstream slope, a spillway designed for a 1:10,000yr, 24 hour storm event, and a downstream blanket rock drain. Water is to be returned to the process plant using a pontoon mounted pump (design by others).
The TSF has been assigned a hazard rating of Significant and the RWD has been assigned a hazard rating of ‘Significant’/’High C’ in accordance with ANCOLD guidelines. The designs have been developed in accordance with the guidelines requirements for these hazard ratings.
Design work included seepage and stability assessment. Under normal operating conditions, seepage from the TSF is expected to be minimal as a relatively small decant pond is expected on the TSF. The stability analyses indicate that the proposed RWD and TSF embankments have adequate factors of safety when compared with the recommended minimum factors of safety in ANCOLD guidelines.
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The results of the assessments for the development of the design of the TSF and RWD indicate that these facilities can be safely operated on the basis that:
Liberated water is continually removed from the surface of the TSF.
The pontoon mounted pump is continuously operated to remove water from the top of the tailings in the TSF.
Tailings deposition to the TSF is regularly cycled along the main embankment and adjacent hills forming the abutments to the main embankment to maximise tailings density and therefore the storage volume of the facility. The sloped tailings beaches that are developed will concentrate water in the western end of the facility.
The storages are constructed in accordance with the design and operated in accordance with the Operating Manuals.
The safe operation of the tailings storage relies upon the implementation of the tailings operation, management inspection and maintenance procedures.
Recommendations for follow-up study include:
Development of detailed TSF and RWD construction specifications and scopes of work.
Further geochemical characterisation of the tailings and mine waste.
Assess raw water dam water balance and filling rate schedule
18.4 Infrastructure
The infrastructure required for the development of the Kapulo mine is as follows:
The process plant area which includes the mine administration building.
The mine services area, including workshops, fuel farm and powerhouse.
Explosives storage area.
The staff village.
The process water dam.
The tailings storage facility.
Access road from Pweto is being upgraded by the company at present.
The airstrip at Pweto will be used for the foreseeable future as the mine is only 45min by road from the airstrip.
Site access roads will require upgrading as part of mine construction.
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18.5 Personnel
The Kapulo mine will have 450 employees of which 30 will be expatriates.
A breakdown of the anticipated personnel by department is shown below:
Health/Safety/Environmental 15
Training/Social Development 10
Administration 20
Logistics 30
Camp 40
Geology 35
Mining 110
Process Plant Operations 75
Maintenance 110
Security 5
In addition to this contractors will be used for mining, security and laboratory analysis.
18.6 Markets
Copper is a major industrial metal (ranking third after iron and aluminium by consumption) because it is highly conductive (electrically and thermally), highly ductile and malleable, and resistant to corrosion. Electrical applications of copper include power transmission and generation; building wiring; motors; transformers; telecommunications; electronics and electronics products; and renewable energy production systems.
Research by CMCC has indicated there is an attractive market for high grade copper concentrates. Demand for Copper is expected to remain high.
The copper price forecast used in the economic evaluation of the project is shown in the Table 18.6_1. The forecast is based on published LME monthly futures prices using the June 2011 contract as the basis for each prospective year through to 2019.
Table 18.6_1 Kapulo Copper Project Copper Price Projection
Year Cu Price 2013 2014 2015 2016 2017 2018 2019 2020 US$/t 8,778 8,551 8,305 8,042 7,734 7,438 7,190 6,974 US$/lb 3.98 3.88 3.77 3.65 3.51 3.37 3.26 3.16
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Copper Concentrate will be sold either directly to smelters in country or worldwide dependent on the available capacity via an agent or directly to metal trading companies.
18.7 Contracts
CMCC is not a party to any contract for the sale of copper concentrate from its Project.
Initial Tenders have been requested by various construction companies for the construction of the plant.
18.8 Environmental
A biophysical and social environment baseline study of the Kapulo permit was conducted by African Mining Consultants between November 2009 and September 2010. The following are the principal findings:
There are no human settlements or villages on the mine site; the nearest settlement is Kapulo village (1.4km east),
There are two streams (Mposha and Kasongo) and one river (Lunkinda) within the permit,
The predominant flora is Miombo forest vegetation,
The sulphur dioxide ambient air quality is (AMC) <0.2 µg/m³ µg/m³ (WHO guideline limit is 12.5µg/m³),
The "equivalent" average sound level measured at Kapulo was 49dB (DRC limit is 70dB),
The soils in the Kapulo permit are strongly to slightly acidic and the predominant soil groups are:
cambisols,
regosols, and
alisols.
There is no fauna that will be displaced by the project implementation.
The Project will have the following significant positive impacts:
It will contribute to the central and provincial governments by remitting statutory taxes,
The project will generate more than 350 direct and indirect employment opportunities,
It will increase local and national economic development by creating business opportunities for contractors and service providers, and
It will provide skills training opportunities for unskilled staff.
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To enhance these positive impacts, Impact Enhancement Plans have been developed in terms of:
Maximising the sustainability of the Project,
Offering competitive salaries and wages to employees,
As far as practical, offering business opportunities to local contractors and service providers,
Introducing a skills training program for unskilled staff.
The Project’s significant negative impacts identified are:
Possible contamination of air,
Possible contamination of surface water,
Possible increase in noise levels,
Possible contamination of soils,
Stress on the social and public health services due to population influx, and
Danger to worker’s health and welfare due to the spread of HIV/AIDS.
To mitigate these negative impacts, Impact Mitigation Plans have been developed in terms of:
Suppressing dust generation by spraying water on haulage roads,
All plant effluents will be settled in settling ponds before discharge,
Encasing noise sources and providing noise protection equipment,
Constructing a sustainable tailings dam and disposing of tailings and other solid waste at the waste disposal site,
Supporting the local government by helping to provide social and health services, and
Educating workers on the dangers of HIV/AIDS and providing Voluntary Counselling and HIV Testing.
Mawson West will ensure that their environmental and social management plans and environmental best practice is followed by all contractors and sub-contractors. Furthermore, Mawson West will maintain a continuous public consultation process with local and regional administrators, Kapulo community and other interested and affected stakeholders.
18.8.1 Legislative Background
The Environmental Impact Assessment (EIA) and the Environmental Management Plan (EMP) for the Project has been conducted in compliance with the DRC regulations and laws. The relevant laws and regulations are discussed below.
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DRC Mining Code – Law No 007/2002
The Mining Code of 2002 provides legislative policy for development and operations of any mining project in DRC. The Mining Regulations (Decree No 038/2003), discussed in Articles of the Mining Code that are relevant to the Project are discussed below.
Article 15 – Department for the Protection of the Mining Environment
The Department in the Ministry of Mines is responsible implementing and enforcing legal provisions of the Mining Code and the Mining Regulations. To that effect, statutory environmental reports are submitted to the Department for the Protection of the Mining Environment (DPEM) for evaluation and approval.
Article 204 – Obligations Relating to Mining Titles During Exploitation
Any applicant for an Exploitation Licence, an Exploitation Licence for Tailings, a Small-scale Mining Exploitation Licence, or an Authorization for Quarry Exploitation must submit an environmental impact study (EIS) together with an EMP for the Project, and obtain the approval of his EIS and EMP, as well as implement the EMP.
The EIS will include a description of the ecosystem before commencing mining operations, including the flora and fauna, soil and topography, air quality, underground and surface water. It specifies the aspects which may be affected qualitatively and quantitatively by the mining or quarry exploitation activity.
It will also include the measures planned for the protection of the environment, the elimination or the reduction of pollution, the rehabilitation of the sites, as well as the verification of the effectiveness of said measures. The holder of mining or quarry rights must provide security in order to guarantee the compliance with the environmental obligations relating thereto during exploration and/or exploitation.
DRC Mining Regulations – Decree No 038/2003
The Mining Regulations provides the legislative regulations for implementing the policies of the Mining Code. Articles of the Mining Regulations that are relevant to the Project are discussed below.
Article 407 – Operations to Submit an Environmental Impact Study (EIS)
Except for temporary quarrying, any operation shall be the object of a Project Environmental Impact Study and a Project Environmental Management Plan established and approved in advance, pursuant to the provisions of Chapter V of the Mining Regulations.
The Environmental Impact Study and the Environmental Management Plan shall be submitted at the same time as the application for the mining/exploitation right. Their approval by the appropriate authority is a condition for award of the right.
With respect to the Environmental Impact Study, the appropriate authority is the Directorate responsible for Mining Environment Protection pursuant to the provisions of Article 42 of the Mining Code.
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Article 450 – Environmental Impact Study and Management Plan
Pursuant to Article 204 of the Mining Code, all mining or quarrying operations except for temporary quarrying shall be the object of a Project Environmental Impact Study and a Project Environmental Management Plan.
The Project Environmental Management Plan is an implementation plan for mitigation and rehabilitation measures developed under Title V of the Environmental Impact Study in accordance with the directive in Schedule IX of the Mining Regulations.
All mining operations stemming from integrated activity, including concentration, processing and transport operations, are part of the same the Project Environmental Impact Study.
Article 451 – Objective of the Public Consultation
Public consultation during the development of the Project Environmental Impact Study shall allow the local populations affected by the mining or quarrying project to participate actively in developing the Project Environmental Impact Study.
The public consultation program during development of the Project Environmental Impact Study shall involve presenting and explaining the mining or quarrying work program, the project’s negative and positive impacts and the mitigation and rehabilitation measures to the local populations affected and gathering their reactions, questions and concerns. The representative of the mining company responsible for public relations with the local populations shall, as soon as possible, send the Territory Administrator and the representatives of every community concerned a written summary of the Project Environmental Impact Study or the Project Environmental Impact Study in the local language summarizing the mining or quarrying work program, the project’s negative and positive impacts and the proposed mitigation and rehabilitation measures.
The applicant, as the Holder of a mining or quarrying exploration right, shall have established sound relations with every community directly affected by the project and undertaken the following measures in particular:
Be familiar with the populations concerned, their main activities, their social and cultural values;
Inform the local populations regarding the exploration work program and exploration project’s negative and positive impacts;
Consult the populations affected during development of the mitigation and rehabilitation measures program;
Compensate persons affected by the exploration project.
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Measures establishing the basis for relations and targeting a good understanding between the mining company and the local populations affected by the Project that were already part of the applicant’s Mitigation and Rehabilitation Plan shall be implemented during preparation of the Project Environmental Impact Study. If, for some reason, these measures were not carried out during the exploration work or there are points of dissension between the mining and the local populations, the applicant shall remedy those shortcomings before establishing its consultation public program during the development of the Project Environmental Impact Study or the Project Environmental Impact Study.
Article 452 – Objectives Environmental Management Plan
The objectives for development of the Project Environmental Management Plan shall be as follows:
ensure the safety of the site during and after the mining or quarrying operation;
reduce the adverse effects of the mining or quarrying operation on the atmosphere, on water sources and watercourses to an acceptable level;
harmonize the mine or quarry and the infrastructures with the landscape through appropriate development to protection wildlife and vegetation;
reduce erosion, leakage of water or chemicals and irregularities in the landscape resulting from the mining or quarrying operation, as well as its adverse effects on the habitat of wildlife species and local flora;
improve the well-being of local populations by implementing economic and social development programs and by providing for compensation of the populations in the event of the displacement of their home;
reduce adverse effects of the mining or quarrying operation such as shock, noise, dust, etc., on the activities of the human and animal populations that live in the surrounding area;
prevent the introduction of parasites and undesirable plants as well as the development or propagation of disease in areas where none previously existed; and
encourage rapid re-growth and renewal of plant species that are indigenous or compatible with the ecosystem of the site.
Schedule IX – Directive In Respect of the Environmental Impact Study
This whole schedule provides the framework for conducting the Environmental Impact Study and developing the Environmental Management Plan.
Article 43 of Schedule IX
This Article obligates the project developer to develop and submit a mitigation and rehabilitation program to reduce or eliminate any negative impacts identified in the environmental impact assessment.
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Article 46 and 47 of Schedule IX
These articles provide the noise pollution thresholds for three categories of land.
Article 50 of Schedule IX
This article provides the air pollution thresholds within and outside the permit area.
Article 66 of Schedule IX
This article provides the maximum concentration of contaminants in surface water and the effluent limits at the discharge points.
Article 80 of Schedule IX
This article provides the guidelines for storage of mine wastes and tailings.
Schedule XI – Classification of Mine Waste / Tailings
This schedule provides the guidelines for classification of mine solid wastes.
Flora Code
The flora of DRC covers 1.3 million square kilometres of the country or 54.6% and is governed by a New Forest Code enacted in 2002. This Code manages the flora especially forested lands under three themes; exploitation, community use and conservation. Threats to conservation of forest resources include expansion of urban areas, logging industry, conflicts, forest fires, mining and harvesting for fuel-wood (Sebastien and N’yanga-Nzo (2001)). The logging industry is of particular interest as it opens up once closed off forest areas and access into logged areas may create conservation problems for flora and fauna. Mining on the other hand is governed by guidelines outlined in the Mining Code. The objective of such guidelines is to ensure that mining does not result into undue environmental damage that may be expensive to repair. The proposed mining project at Kapulo will be governed by similar guidelines and this paper has been written with particular reference to the flora, how it will be affected by mining and what conservation measures are needed to be put in place to mitigate negative effects of mining on forest resources.
18.8.2 Baseline Study
The extent of the physical environmental baseline study undertaken for the Kapulo licence area is shown in Figure 18.8.2_1.
Topography
Field observations together with satellite imagery have been used to define the topography of the Kapulo area (Figure 18.8.2_2). The topography is defined by the valley for the Mposha and Kasongo streams. Mposha is a relatively small and intermittent stream that flows from east to west into the Lunkinda River. The Kasongo stream flows from west to east into the Lwawu River. Both Kasongo and Lwamu rivers flow into Lake Mweru.
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Figure 18.8.2_1 Proposed Kapulo Exploitation Permit Inside Current PR
Figure 18.8.2_2 Topography of Kapulo
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There is a hill range on the western and eastern side of the proposed open pits. The landscape is dominated by Miombo forests, minor artisanal workings and linear features like roads and watercourses with riparian vegetation.
Soils
The soil quality assessment that was conducted in November 2009 by African Mining Consultants Limited revealed that the soil groups in the Kapulo permit area are (in descending order of predominance):
Cambisols with different suborders,
Regosols due to accidental geomorphology that hinders soil weathering and keeps rogoliths in permanent rejuvenation, and
Alisols, podzols, and fluvisols wherever deep soil weathering is possible.
Land evaluation for suitability classes revealed that:
About 60% of soils under natural condition are unsuitable for exacting crops (maize),
Unsuitable to slightly suitable for moderately exacting crops, and less exacting crops (cassava) due to topography, soil depth and chemical fertility;
Almost 30% of soil type are only slightly suitable for exacting crops, slightly too moderately suitable for moderately exacting crops and moderately suitable for less exacting crops.
Prior to mining activities, normal soils in the upper horizon, except where informal mining was performed (Safari, Katanga, and Shaba), have copper level ranging from 10ug/g up to 18ug/g and cobalt level from 38ug/g up to 320ug/g.
Risk of Natural Disasters
The risk of earthquakes in the Project area is considered as low. The seismic hazard map of Africa (G. Grünthal, C. Bosse, Geoforschungs Zentrum, Potsdam, Germany) shows that there is a 10% probability of a peak ground acceleration of between 0.4m/s² and 0.8m/s² being exceeded every 50 years in the Kapulo project area (Figure 18.8.2_3). Therefore, the Project area has a low to medium susceptibility to an earthquake.
Climate
There are three main weather stations with reliable weather data for the Kapulo project area. These are:
The weather station at Kapulo mine camp
The weather station at Dikulushi mine (125km south west), and
The Kawambwa meteorological station (165km south).
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Figure 18.8.2_3 Seismic Risk Map of Southern Africa
Meteorological data from these three stations (Kapulo (June 2009 – April 2010), Dikulushi (2006-2008) and Kawambwa (1970-2000)) indicate that the regional weather is defined by two distinct seasons in a year – the dry and cool season (May to September) and the rainy season (October to April).
Rainfall
Meteorological data (1970-2000) from the Kawambwa station predicts that the mean annual rainfall for Kapulo is 1, 320mm. Climatic data (1970-2000) from Kawambwa meteorological station shows that wettest month was March. On the other hand, the weather data (2006- 2010) from the Kapulo and Dikulushi weather stations show that the wettest month has been January.
Figure 18.8.2_4 through to Figure 18.8.2_6 below show monthly rainfall and temperature for the three stations described above.
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Figure 18.8.2_4 Kapulo Monthly Rainfall and Temperature
Figure 18.8.2_5 Dikulushi Monthly Rainfall and Temperature
Rainfall and Temperature (Dikulushi (2006-2008)) 180
160 32 140 C) 120 27 o 100 80 22 60 Rainfall(mm)
40 17 Temperature( 20 0 12
Month
Rainfall (mm) Temp Min Temp Max
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Figure 18.8.2_6 Seismic Risk Map of Southern Africa
Rainfall and Temperature (Kawambwa (1970-2000)) 280 30 240 26 200 C) o
160 22 120 18 Rainfall Rainfall (mm) 80 Temperature( 14 40
0 10 Jan Feb Mar Apr May Jun Jul Aug Sep Oct Nov Dec Month
Month Min Temp Max Temp
Rainfall mainly occurs as heavy thunderstorms, which can produce between 10mm and 40mm of rain during a typical precipitation event.
Extreme weather events such as floods, droughts and high winds do occur from time to time. The 30 year and 100 year maximum 24 hour storm event is calculated as 135mm and 162mm respectively. Annual rainfall of between 1,400mm and 1,500mm per annum is likely to be reached or exceeded in one year out of five.
Temperature
Average annual temperatures in the region vary between 11°C and 31°C throughout the year (as shown in Figure 18.8.2_4 above). Minimum temperatures occur during the cold season months of June and July. Maximum temperatures occur in the month preceding the onset of the rains (September/October).
Wind
Wind frequency data for Kawambwa based on five years of daytime observations indicates that the prevailing dry season (July) wind direction is from the east to south east with mean and maximum wind speeds of 0.9m/s-1 and 9.1m/s-1 respectively. The mean number of calm days in July is 2.8. Observations taken at the beginning (October) and the end (April) of the rainy season indicate that the prevailing wind direction is from the east, south-east with maximum wind speeds of 9.3m/s-1 (October) and 9.1m/s-1 (April).
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The prevailing wet season (January) wind direction is from the west to northwest with minimum and maximum wind speeds of 0.8m/s-1 and 4.8m/s-1 respectively. The mean number of calm days in January is 11.2.
Air Quality
The determination of baseline air quality of the proposed project is essential and to that
extent, African Mining Consultants conducted ambient sulphur dioxide (SO 2) passive sampling at Kapulo in October 2010.
One SO 2 passive samplers from IVL laboratories in Sweden were exposed for 30 days at the Main Driller’s camp as shown in Table 18.8.2_1 below. The Main Driller’s camp was selected because it provided the needed security for the passive sampler against vandals. Such security was not guaranteed in Kapulo village. Nevertheless, the ambient air quality at the two sites is representative of the ambient air quality in Kapulo village.
Table 18.8.2_1 Kapulo Copper Project
Ambient SO 2 Quality Results
SO DRC Limit WHO Guideline Exposure Period Exposure Site Location 2 (µg/m³) (µg/m³) (µg/m³) UTM 9084014 27/10/2010 - 26/11/2010 Main Driller’s Camp <0.2 5, 000 125 35L 0745255
After exposure, the passive sampler was dispatched to IVL for analysis under standard temperature and pressure. The analytical results obtained from the passive sampler, the DRC
SO 2 limit within the mine permit (Article 50, Schedule IX of the DRC Mine Regulations and the
World Health Organisation (WHO) SO 2 guideline are shown in Table 18.8.2_1
It is evident from the table above that the ambient SO 2 quality for Kapulo (<0.2 µg/m³) is less than the DRC limit of 5,000 µg/m³ and the World Health Organisation guideline of 125 µg/m³.
Noise Quality
Prior to the noise baseline study conducted by African Mining Consultants in November 2009, there was no historical environmental noise quality data for the Project area. Current noise is associated with exploration drilling operations and power generators at the mine camp.
African Mining Consultants conducted a noise baseline study of the Project area at three selected points that are shown in Table 18.8.2_2 and Figure 18.8.2_7. The noise quality assessment was conducted between 07:00 hours and 19:00 hours using a hand-held Rion NL 14 Noise level Meter.
The noise quality (in dB) results as shown in Table 18.8.2_2 are expressed in LAeq – the "equivalent" average sound level measured using the A-weighting which is most sensitive to speech intelligibility frequencies of the human ear.
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Table 18.8.2_2 Kapulo Copper Project Noise Quality Results
DRC Limit Site Description Coordinates LAeq Day Time Night Time 0746485 Safari North Open Pit 49.8 70 70 9079146 0746034 Shaba Open Pit 50.8 70 70 9081514 Kapulo Village (1.3km and 2.4km from 0747324 41.1 45 40 Shaba and Safari North open pits) 9081110
Figure 18.8.2_7 Noise Contour Map of Kapulo Permit
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According to Article 47, Schedule IX of the DRC Mining Regulations, the ‘land category’ of Kapulo permit is Category C – ‘land where mainly industrial activities are being conducted’. With reference to Article 46 of the Mining Regulations – Schedule IX, the noise level limit for ‘land category’ C is 70 dB (A).
According to the same Article 47, the ‘land category’ for Kapulo village is A – ‘land comprising several dwellings of a residential nature constituting a village, a school or a hospital or any other establishment procuring educational, health or convalescence services’. The noise level limit for Kapulo Village is 40dB and 45dB during night time and day time respectively.
Figure 18.8.2_7 shows that the LAeq range in the Kapulo permit ranges from 41.1dB to 50.8dB with the highest being at Shaba open pit where exploration drilling was taking place. All the three baseline analytical results for ambient noise are less than the DRC limit (45dB or 70dB during the day time).
Surface Water
Surface Water Drainage
The main surface water drainage in the Project area is the Lunkinda River whose source is in Miombo woodlands and if flows from north to south towards Lake Mweru. As shown in Figure 18.8.2_8, the other rivers in the project area are:
Mposha Stream A relatively small and intermittent stream that flows from east to west towards the Lunkinda River,
Kasongo Stream Also a relatively small river that flows from west to east to Lwawu River, a tributary of Lake Mweru.
Surface Water Sampling
Two main representative surface water-monitoring sites were selected as sampling points for surface water. The location of the monitoring sites is shown in Table 18.8.2_3 below and in Figure 18.8.2_8 referred to above.
Analytical Results
The surface water analytical results for Lunkinda River upstream of Mposha stream confluence are shown in Table 18.8.2_4 below and these are compared to the 2004 WHO guideline for drinking water (all units are in g/ml, unless shown otherwise). Water samples were collected in December 2009 and in August 2010.
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Figure 18.8.2_8 Kapulo Surface Water Drainage
Table 18.8.2_3 Kapulo Copper Project Surface Water Monitoring Sites
Sampling Site Location Description 0745291 SW-01 Lunkinda River upstream of Mposha stream confluence. 9084164 Lunkinda River Bridge – downstream of Mposha stream confluence 0745155 Lunkinda River upstream of Mposha stream confluence Lunkinda River SW-02 upstream of Mposha stream confluence Lunkinda River upstream of Mposha 90839438803636 stream confluence Lualaba river upstream Lualaba river upstream Lualaba river upstream Lualaba river upstream Lualaba river upstream.
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Table 18.8.2_4 Kapulo Copper Project Lunkinda Upstream – Surface Water Quality
WHO Drinking Lunkinda Upstream Dec-09 Aug-10 Water Guideline Aluminium <0.3 0.2 Arsenic <0.01 <0.01 0 Cadmium <0.01 <0.01 0 Calcium 1.65 <0.02 - Chloride <1.2 <1.2 - Chromium <0.03 <0.03 0.1 Cobalt <0.03 <0.03 - Coliform-Faecal (/100ml) >60 >60 0 Coliform-Total (/100ml) >60 >60 0 Conductivity (µS/cm) 58 37 - Copper <0.02 0.04 2 Cyanide <0.05 <0.05 0.7 Floride 0.09 0.12 1.5 Hardness 16 - Iron 2.04 0.46 - Lead <0.04 <0.02 0.01 Magnesium 0.24 0.13 50 Manganese 0.03 <0.03 - Mercury <0.001 <0.001 0.01 Nickel <0.03 <0.03 0 Nitrate <0.07 - Nitrite <0.004 - pH (Unit) 7.7 7.11 - Potassium 2.3 - Selenium <0.01 <0.01 0.01 Sodium 4.31 4.66 - Sulphate <2.8 5 - Total Dissolved Solids 24 20 600 Total Suspended Solids 48 4 - Turbidity (NTU) 7.26 - Zinc <0.02 <0.02 3
The analytical results for Lunkinda River upstream of Mposha stream confluence shown in Table 18.8.2_4 above demonstrate that the surface water analytical parameters are in compliance with the 2004 WHO guideline for drinking water except for:
Total Coliform (WHO guideline – 0/100ml),
Faecal Coliform (WHO guideline – 0/100ml).
The surface water analytical results for Lunkinda River Bridge – Downstream of Mposha stream confluence are shown in Table 18.8.2_5 and these are compared to the 2004 WHO guideline for drinking water (all units are in g/ml, unless shown otherwise). Water samples were collected in December 2009 and in August 2010.
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Table 18.8.2_5 Kapulo Copper Project Lunkinda Downstream – Surface Water Quality
WHO Drinking Lunkinda Downstream Dec-09 Aug-10 Water Guideline Aluminium * <0.3 0.2 Arsenic <0.01 <0.01 0 Cadmium ,0.01 <0.01 0 Calcium 1.3 <0.02 - Chloride <1.2 <1.2 - Chromium <0.03 <0.03 0.1 Cobalt <0.03 <0.03 - Coliform-Faecal (/100ml) >60 >60 0 Coliform-Total (/100ml) >60 >60 0 Conductivity (µS/cm) 59 37 - Copper <0.02 0.04 2 Cyanide <0.05 <0.05 0.7 Floride 0.09 0.13 1.5 Hardness 4 - Iron 0.94 0.5 - Lead <0.04 <0.02 0.01 Magnesium 0.28 0.09 50 Manganese <0.03 <0.03 - Mercury <0.001 <0.001 0.01 Nickel <0.03 <0.03 0 Nitrate <0.07 - Nitrite <0.004 - pH (Unit) 7.72 7.17 - Potassium 2.11 - Selenium <0.01 <0.01 0.01 Sodium 5.05 4.58 - Sulphate <2.8 3 - Total Dissolved Solids 24 22 600 Total Suspended Solids 48 4 - Turbidity (NTU) 7.05 - Zinc <0.02 <0.02 3
The analytical results for Lunkinda River Bridge - Downstream of Mposha stream confluence show that the surface water analytical parameters are in compliance with the 2004 WHO guideline for drinking water except for:
Total Coliform (WHO guideline – 0/100ml),
Faecal Coliform (WHO guideline – 0/100ml).
Groundwater
The Kapulo copper deposits are situated in the Kundelungu lithologies, which are predominantly non-water bearing. A program of RC drilling to define groundwater in the current pit designs was completed by Groundwater Resource Management (GRM) consultants and found the maximum water flow in a drillhole to be less than 2litres/sec. As such, the groundwater in the pits is expected to be managed through a series of in-pit sumps and small pumps.
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Nevertheless, allowances have been made to carry out a detailed ground water assessment during the first six months of mining to confirm observations made during the exploration program.
18.9 Taxes, Duties and Royalties
The Kapulo mine will operate under the Dikulushi Mining Convention which provides for concessionary rates of taxation for each new mine. The first five years of production will be tax free, the effective tax rate from the sixth through to the tenth year of production will be 16%, from the eleventh through to the fifteenth year of production it will be 18% and 40% thereafter.
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
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 3% for Fuel, lubricants and mining consumables and 5% of all other supplies.
No government royalties are applicable at the Project.
18.10 Capital Cost Estimates
The development of the capital cost estimate for the process plant and general infrastructure was provided by Sedgman, whilst the capital cost estimate for the tailings storage facility and the raw water dam were developed by Coffey Mining.
18.10.1 Initial Fixed Capital Costs
Initial fixed capital costs, including first fill, total US$69.5 million and are summarized in Table 18.10.1_1. No escalation has been assumed in the capital costs, nor have allowances been made for capitalized interest or country risk insurance. Interest charges and country risk insurance are treated as financing costs and considered as part of an overall financing package.
18.10.2 First Fill and Working Capital
First fill includes the initial stocks of consumables and strategic spare parts. Ongoing purchases of supplies and parts are treated as operating costs.
During the initial weeks of mine operation and mill commissioning while copper inventory is building up in the process circuit, and before cash is received from the sale of copper, working capital is required to meet operating expenses. Working capital is estimated to be 90 days of Year 1 operating expenses.
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Table 18.10.1_1 Kapulo Copper Project Summary Initial Capital Cost Estimate
Value Item (US$M) Process Plant Direct Costs 40.6 Support Infrastructure 5.8 Tailings/Raw Water Dam 3.8 Insurance/Consumables 1.1 EPCM 9.6 Contingency 6.6 Sub Total 67.5 First fill 1.5 Mobilisation Mining contractor 0.5 Grand Total 69.5
18.10.3 Deferred Capital Costs
Deferred capital costs are associated with the staged construction of the TSF as shown in Table 18.10.3_1.
Table 18.10.3_1 Kapulo Copper Project Summary Deferred Capital Cost Estimate
Value Item (US$M) TSF - Stage 2 – Year 1 0.6 TSF - Stage 3 – Year 2 1.0 TSF – Stage 4 – Year 4 1.0 TSF – Stage 5 – Year 5 2.6 Total 5.2
Over the life of the Project there has been allowance of US$13.9 million for sustaining capital. This was based on a notional 2.5% of the initial capital expenditure.
18.11 Operating Costs
The principal cost items comprising the overall project operating cost can be categorised as follows:
Mining cost.
Processing cost.
General and administration.
Product transport and refining.
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A summary of the derivation of the cost items listed above is provided in Table 18.11_1.
Table 18.11_1 Kapulo Copper Project Summary Operating Cost Derivation
Item Responsible Entity Source Mining Cost Mawson West Contract submission obtained from MCK Trucks First principals, supplier quotations, Sedgman cost Processing Cost Sedgman database, Mawson West in-country experience Sedgman cost database and Mawson West in- General and Administration Sedgman / Mawson West country experience Product Transport and Refining Mawson West Mawson West in-country experience
A summary of the mining cost estimate is provided In Table 18.11_2.
Table 18.11_2 Kapulo Copper Project Summary Mining Costs
Value Item (US$/t milled) Contract Mining (Refer Section 18.1.5) 3.11 Grade Control 0.04 Total 3.15
A summary of the processing cost broken down by cost centre is provided in Table 18.11_3.
Table 18.11_3 Kapulo Copper Project Summary Processing Cost Estimate
Value Item (US$/t milled) Labour 6.15 Operating Consumables 7.33 Maintenance Consumables 1.43 Power 18.08 Transport 2.53 Total 35.53
The labour costs were determined using remuneration levels supplied by Mawson West based upon current labour rates at the Dikulushi operation.
The power cost of was estimated at US$0.3784/kWh, based on a diesel consumption of 0.26L/kwh, a diesel cost of US$1.34/L and a maintenance cost of US$0.03/kWh.
The life of mine operating costs for the Project are presented in Table 18.11_4 and include all mining, processing, general and administrative costs and product transport and refining costs. The operating costs exclude depreciation and amortization.
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Table 18.11_4 Kapulo Copper Project Life of Mine Operating Costs
Value Unit Cost Item (M$) ($/t lb) Contract Mining Costs 121.4 0.47 Processing Costs 128.5 0.50 On Site General and Administration Costs 43.8 0.17 Sub Total 293.7 1.14 Marketing Duties and Taxes 46.5 0.18 Metal Sales and Transport Costs 166.1 0.65 Sub Total 212.6 0.83 Total 506.3 1.97
18.12 Economic Analysis
18.12.1 Introduction
This section presents a valuation of 100% interest in the Project. The analysis used a Discounted Cash Flow (“DCF”) methodology, which determines the value of an asset by calculating the net present value of the future cash flows over the useful life of that asset.
The model is presented as an equity model assuming 100% equity financing. No allowance has been made in the model for the effects and levels of debt financing available or required.
Mawson West overall interest in the Project is 90%.
18.12.2 Project Economics
The results of the DCF model, based on the economic input parameters as described in the previous sections, including the copper prices as shown in Table 18.6_1, are shown in Table 18.12.2_1.
Table 18.12.2_1 Kapulo Copper Project Life of Mine Project Cash Flows
Item Unit Value Net Revenue US$M 735 Capital Expenditure US$M (90) Operating Expenses US$M (294) Marketing Duties and Indirect Taxes US$M (46) Free Cash Flow US$M 305 Net Present Value (@ 10%) US$M 157 Mawson West portion (90%) US$M 141 Internal Rate of Return % 61
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The principal results of the financial evaluation are as follows:
Initial Capital Investment US$69.5 million
Net Present Value (NPV) at 10% US$157 million
Internal Rate of Return 61%
The pay back for the capital invested in the project is less than 2 years.
18.12.3 Sensitivity Analysis
Sensitivity analyses have been undertaken on the financial model for the following parameters:
Copper Price ± 20%, in 10% increments.
Operating Costs ± 20% in 10% increments.
Capital Costs ± 20% in 10% increments.
Discount Rate ± 2% in absolute terms.
Table 18.12.3_1, Table 18.12.3_2 and Table 18.12.3_3 show the impact on the NPV of a change in Cu price, operating cost and capital cost, at different discount rates.
Table 18.12.3_1 Kapulo Copper Project Project Sensitivity to a Change in Cu Price
NPV (US$ million) Discount Rate Change in Cu Price -20% -10% 0% 10% 20% 8% 60 118 176 235 293 10% 50 103 157 211 265 12% 41 90 140 190 239
Table 18.12.3_2 Kapulo Copper Project Project Sensitivity to a Change in Operating Costs
NPV (US$ million) Discount Rate Change in Operating Costs -20% -10% 0% 10% 20% 8% 217 197 176 156 135 10% 195 176 157 138 119 12% 175 158 140 122 105
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Table 18.12.3_3 Kapulo Copper Project Project Sensitivity to a Change in Capital Costs
NPV (US$ million) Discount Rate Change in Capital Costs -20% -10% 0% 10% 20% 8% 193 185 176 168 160 10% 173 165 157 149 141 12% 155 148 140 132 125
Figure 18.12.3_1 displays a spider diagram showing the sensitivities that were run at a 10% discount rate.
Figure 18.12.3_1 Kapulo Copper Project Summary Project Sensitivity Analysis 300
250
200
150
NPV (US$ (US$ NPV million) 100
50
0 -20% -15% -10% -5% 0% 5% 10% 15% 20%
Percent Change fro Base Case
Cu Price Operating Costs Capital Costs
The sensitivity analysis showed that the Project is most sensitive to a change in Cu price. This is a common phenomenon where the parameters that affect the revenue will have the greatest impact on a project, with a change in the Cu grade or processing recovery exhibiting a similar result as shown for the copper price. A 10% change in the Cu price resulted in an approximate 34% change in the NPV.
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18.13 Project Implementation
In terms of the DRC Mining Code and the DRC Mining Regulations (Decree No 038/2003 of March 2003) the applicant of an exploitation licence is required to submit an EIS and EMP to DPEM.
The EIS and EMP have been prepared by AMC Consulting and have been lodged with DPEM in 22 May 2011. Whilst Mawson West doesn’t foresee any issues with the granting of an exploitation licence, due process is to be followed. At this stage Mawson West anticipates that the exploitation licence will be granted in the third quarter of 2011.
Until such time the exploitation licence is granted, Mawson West is not in a position to enter into an Engineering, Procurement and Construction Management (EPCM) contract for the design and construction of the treatment plant, associated infrastructure and services that are the subject of the DFS.
Nevertheless, for the purposes of the DFS, it has been assumed that after a decision has been taken to proceed with the development of the Project, Mawson West will call for tenders from several experienced and reputable engineering companies for the implementation of the Project on an EPCM basis. This will include the undertaking of preliminary front end engineering design (FEED) prior to project award to allow for the timely ordering of the long lead items such as the ball mill.
It is envisaged that the Contractor will execute most of the engineering, design and procurement from its home office and will maintain a site based construction management team during the construction phase. The site based team will include engineers and discipline supervisors to suit the activities being undertaken at any given time.
The Contractor will manage all equipment supply contracts and the site works, which will be undertaken by experienced project contractors/suppliers who have been successful in open tenders for the respective work packages. All Contracts will be between Mawson West and the Project Contractor for equipment and material supply and site works contracts. During the latter phases of construction, a testing and pre-commissioning crew will, under the direction of the Contractor’s Commissioning Manager, carry out all relevant checks and dry run tests and ensure that the plant circuits are water commissioned.
After Contract Practical Completion has been achieved, commissioning with ore will commence under the direction of the Commissioning Manager. The commissioning team will include the Contractor’s metallurgists, discipline engineers and tradesmen, as well as Mawson West operations and maintenance labour forces.
The Mawson West labour force will have been recruited, inducted in pre-established operating procedures and be trained to an appropriate level prior to the commencement of ore commissioning. It may be necessary for some equipment supplier’s representatives to return to the site to assist during the ore commissioning phase.
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The preliminary Project schedule is expected to take approximately 74 weeks from the project award to completion of ore commissioning. This schedule assumes the commencement of the FEED 8 weeks prior to the Project award to allow sufficient preliminary engineering to be undertaken to allow timely placement of the order for the ball mill. Practical completion of the processing plant is estimated to be 68 weeks after Project award. The expected key project milestone dates are provided in Table 18.13_1.
Table 18.13_1 Kapulo Copper Project Summary Key Project Milestones
Milestone Time (Weeks from Project Award) Commencement of FEED -8 Project Award 0 Placement of Mills Order 1 Delivery of Mills 54 Commence Commissioning 62 Practical Completion 72 Completion of Ore Commissioning 78
These key milestone dates are largely dependent on award of the long lead equipment, in particular the ball mill and the flotation cells.
18.14 Risk Assessment
The principal risks to the Project that were assessed as either extreme or high require further work on mitigation during detailed engineering and are discussed below.
The highest risk identified concerns the political risk to the mining operation. This is high because political influences in the country could stop the mining operation.
The risk to the copper and silver produced is minimal, as the mining rights are well defined. However, if there were to be political upheavals in the country, the copper concentrate and silver produced on the plant, were the plant to continue to operate, may be under threat. This would directly threaten the Project due to the potential loss in revenues.
The time taken to ramp up to full production as a result of late mobilisation of crews with the necessary skill levels to meet production targets is a risk for the Project.
Pay scales have been developed to allow the company to attract the necessary quality staff. A comprehensive training strategy will be implemented using international trainers, OEM training specialists, experienced staff from within the DRC and other West African mining regions, and an in-house training function.
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Whilst proven and relatively simple technology will be utilised, due to the plant’s remote location, the plant must carry a comprehensive store of spares. This is critical as the availability of suitable spares in the country is unreliable at best, and specialised spare parts must be imported. Delays to operations will occur if no suitable spares are available on site.
The region where the Project is located has a well defined rainy season when torrential rains fall over short periods. This rain could potentially damage the roads on site, cutting off supplies and feed from the mine. Good road construction and prompt maintenance will minimise the risk to the roads.
All projects of this nature are subject to an element of risk due to the potential for loss, damage or delay occurring in international transport operations. The overall transport periods may thus also influence the end dates for the installation of the equipment items concerned.
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19 INTERPRETATION AND CONCLUSIONS
The pertinent observations and interpretations which have been developed in producing this Technical Report of the Kapulo Copper Project are detailed in the sections above.
The key recommendations are contained in Section 20 below.
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20 RECOMMENDATIONS
The current Project financial model indicates that the Project is economically viable and it is recommended that Mawson West proceeds with:
The construction of the process plant and other infrastructure.
Additional metallurgical testwork on the Safari North oxide and sulfide material to assess processing recoveries.
Further geochemical testwork on tailings and waste rock samples to further assess the potential of acid rock drainage.
Development of detailed TSF and RWD construction specifications and scopes of work.
Assess RWD water balance and filling rate schedule.
Consideration and evaluation of Mawson West supplying some major mining consumables (e.g. fuel and explosives) to share some risk and potentially reduce contract mining costs.
Detailed ground water assessment during the first six months of mining to confirm observations made during the exploration program.
Further infill and extensional drilling to raise the level of confidence and extend the Inferred Resources at both Shaba and Safari North.
Further resource definition drilling as the geophysics completed during 2010 has provided a number of targets for drilling in 2011 and, as such, there is reasonable potential to increase resources within trucking distance of the Project.
Mawson West has provided a comprehensive 12 months exploration and development program and budget for the Project, which incorporates the recommendations listed above. Expenditure for the Project (US$) is detailed below in Table 20_1.
Table 20_1 Kapulo Copper Project Future Work Summary
Activity Cost Further Geotechnical Drilling and Study $75,000 Resource Definition Drilling RC and DDH (Met) $250,000 Metallurgical Testwork Safari North $100,000 Regional Drilling $450,000 Further Geochemical Testwork $20,000 Geophysics $150,000 Resource Estimation $75,000 Detailed Engineering Studies $400,000 Total $1,520,000
Coffey considers that the proposed exploration and evaluation strategy is consistent with the potential of the project. The proposed expenditure is also generally considered to be adequate to cover the cost of the proposed programs.
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21 REFERENCES
Cailteux, J.L.H., 1994, Lithostratigraphy of the Neoproterozoic Shaba-type (Zaïre) Roan supergroup and metallogenesis of associated stratiform mineralisation: Journal of African Earth Sciences, v. 19, p. 279-301.
Dewaele, S., Muchez, P., Heijlen, W., Boutwood, A., Lemmon, T., and Tyler, R., 2006, Reconstruction of the hydrothermal history of the Cu-Ag vein-type mineralisation at Dikulushi, Kundelungu foreland, Katanga, D.R. Congo: Journal of Geochemical Exploration, v. 89, p. 376-379.
El Desouky, H.A., Muchez, P., Dewaele, S., Boutwood, A., and Tyler, R., 2007, The stratiform copper mineralization of the Lufukwe Anticline, Lufilian Foreland, Democratic Republic Congo: Geologica Belgica, v. 10, p. 148-151.
François, A., 1974, Stratigraphie, tectonique et minéralisations dans l'arc cuprifère du Shaba (République du Zaïre), in Bartholomé, P., ed., Gisements Stratiformes et Provinces Cuprifères: Liège, La Société Géologique de Belgique, p. 79-101.
Haest, M., 2009, Metallogenesis of the Dikulushi Cu-Ag Deposit in Katanga (DRC) [PhD thesis]: Leuven, Belgium, Catholic University of Leuven.
Haest, M., and Muchez, P., 2011, Stratiform and vein-type deposits in the Pan-African Orogen in central and southern Africa: evid ence for multiphase mineralization: Geologica Belgica, v. 14, p. 23-44.
Haest, M., Muchez, P., Dewaele, S., Boyce, A.J., Quadt, A., and Schneider, J., 2009, Petrographic, fluid inclusion and isotopic study of the Dikulushi Cu-Ag deposit, Katanga (D.R.C.): implications for exploration: Mineralium deposita, v. 44, p. 505-522.
Haest, M., Muchez, P., Dewaele, S., Franey, N., and Tyler, R., 2007, Structural control on the Dikulushi Cu-Ag deposit, Katanga, Democratic Republic of Congo: Economic Geology and the Bulletin of the Society of Economic Geologists, v. 102, p. 1321-1333.
Haest, M., Schneider, J., Cloquet, C., Latruwe, K., Vanhaecke, F., and Muchez, P., 2010, Pb isotopic constraints on the formation of the Dikulushi Cu-Pb-Zn-Ag mineralization, Kundelungu Plateau (Democratic Republic of Congo): Mineralium deposita, v. 45, p. 393-410.
Hitzman, M.W., Kirkham, R., Broughton, D., Thorson, J., and Selley, D., 2005, The Sediment- Hosted Stratiform Copper Ore System: Economic Geology 100th Anniversary Edition, p. 609-642.
Jackson, M.P.A., Warin, O.N., Woad, G.M., and Hudec, M.R., 2003, Neoproterozoic allochtonous salt tectonics during the Lufilian orogeny in the Katangan Copperbelt, Central Africa: Geological Society of America - Bulletin, v. 115, p. 314-330.
Kampunzu, A.B., and Cailteux, J., 1999, Tectonic evolution of the Lufilian Arc (Central Africa Copper Belt) during Neoproterozoic Pan African orogenesis: Gondwana Research, v. 2, p. 401-421.
Porada, H., and Berhorst, V., 2000, Towards a new understanding of the Neoproterozoic- Early Palaezoic Lufilian and northern Zambezi Belts in Zambia and the Democratic Republic of Congo: Journal of African Earth Sciences, v. 30, p. 727-771.
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Richards, J.P., Cumming, G.L., Krstic, D., Wagner, P.A., and Spooner, E.T.C., 1988, Pb isotope constraints on the age of sulfide ore deposition and U-Pb age of late uraninite veining at the Musoshi stratiform copper deposit, Central Africa copper belt, Zaire: Economic Geology and the Bulletin of the Society of Economic Geologists, v. 83, p. 724-741.
Selley, D., Broughton, D., Scott, R., Hitzman, M., Bull, S.W., Large, R.R., McGoldrick, P.J., Croaker, M., Pollington, N., and Barra, F., 2005, A new look at the geology of the Zambian Copperbelt: Economic Geology 100th Anniversary Edition, p. 965-1000.
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22 CERTIFICATES OF QUALIFIED PERSONS
Coffey Mining Pty. Ltd.
Certificate of Qualified Person
1. As an author of the report entitled “Kapulo Copper Project, National Instrument 43-101 Technical Report” dated 30 June 2011, on the Kapulo Copper Project, DRC of Mawson West Ltd (the “Study”), I hereby state:
2. My name is Steven Le Brun and I am an Associate Resource Geologist with the firm of Coffey Mining Pty. Ltd. of 1162 Hay Street, West Perth, WA, 6005, Australia.
3. I am a practising geologist and a member of the AusIMM (202832) and of MICA.
4. I am a graduate of Leeds University in the United Kingdom with a BSc (hons) in Geological Sciences in 1984. In 1987 I graduated from the University of Leicester, United Kingdom with an MSc in Mineral Exploration and Mining Geology.
5. I have practiced my profession continuously since 1987.
6. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
7. I visited the Kapulo Copper Project property and surrounding areas for 2 days in April 2010. I have performed consulting services during and reviewed files and data supplied by Mawson West between April 2010 and the present.
8. I am responsible for all sections of the study except Section 16 and Section 18. I have contributed to Section 17 of the Study and the associated text in the summary, conclusions and recommendations.
9. 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.
10. I am independent of Mawson West Ltd pursuant to section 1.4 of the Instrument.
11. 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.
12. I do not have nor do I expect to receive a direct or indirect interest in the Kapulo 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 30 June 2011.
[Signed] Steve Le Brun BSc (Hons) (Geology Sciences) MSc (Mineral Exploration and Mining Geology) Associate Resource Consultant (Coffey Mining Pty Ltd)
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Coffey Mining Pty. Ltd.
Certificate of Qualified Person
1. As an author of the report entitled “Kapulo Copper Project, National Instrument 43-101 Technical Report” dated 30 June 2011, on the Kapulo Copper Project, DRC of Mawson West Ltd (the “Study”), I hereby state:
2. My name is Harry Warries and I am a Principal Mining Engineer with the firm of Coffey Mining Pty. Ltd. of 1162 Hay Street, West Perth, WA, 6005, Australia.
3. I am a practising mining engineer and a member of the AusIMM (111318).
4. I am a graduate of Delft University of Technology, Holland, and hold and hold a Masters degree majoring in Mining (1989).
5. I have practiced my profession continuously since 1990.
6. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
7. I have not visited the Kapulo Copper Project property. I have performed consulting services during and reviewed files and data supplied by Mawson West between April 2010 and the present.
8. I am responsible for Section 17.3, Section 18.1 and Section 18.4 through to Section 18.14. I have contributed to the associated text in the summary, conclusions and recommendations.
9. 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.
10. I am independent of Mawson West Ltd pursuant to section 1.4 of the Instrument.
11. 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.
12. I do not have nor do I expect to receive a direct or indirect interest in the Kapulo 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 30 June 2011.
[Signed] Harry Warries MSc (Mine Engineering) Principal Mining Engineer (Coffey Mining Pty Ltd)
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Coffey Mining Pty. Ltd.
Certificate of Qualified Person
1. As an author of the report entitled “Kapulo Copper Project, National Instrument 43-101 Technical Report” dated 30 June 2011, on the Kapulo Copper Project, DRC of Mawson West Ltd (the “Study”), I hereby state:
2. My name is Chris Johns and I am an Associate Geotechnical Engineer with the firm of Coffey Mining Pty. Ltd. of 1162 Hay Street, West Perth, WA, 6005, Australia.
3. I am a practising geotechnical engineer and a registered Professional Engineer with the Association of Professional Engineers, Geologists, and Geophysicists of Alberta.
4. I am a graduate of Queen’s University, Canada and hold a B.Sc.E. (Hons.) in Geological Engineering (1994), as well as a M.Sc. in Environmental Engineering, University of Alberta, Canada (1999).
5. I have practiced my profession continuously since 1995.
6. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
7. I have not visited the Kapulo Copper Project property. I have performed consulting services during and reviewed files and data supplied by Mawson West between April 2010 and the present.
8. I have contributed to Section 18.1.3 and I am responsible for Section 18.3. I have contributed to the associated text in the summary, conclusions and recommendations.
9. 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.
10. I am independent of Mawson West Ltd pursuant to section 1.4 of the Instrument.
11. 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.
12. I do not have nor do I expect to receive a direct or indirect interest in the Kapulo 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 30 June 2011.
[Signed] Chris Johns M.Sc. MIEAust, P.Eng. (AB) Associate Geotechnical Engineer (Coffey Mining Pty Ltd)
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Chris Orr and Associates
Certificate of Qualified Person
1. As an author of the report entitled “Kapulo Copper Project, National Instrument 43-101 Technical Report” dated 30 June 2011, on the Kapulo Copper Project, DRC of Mawson West Ltd (the “Study”), I hereby state:
2. My name is Christopher Martin Orr and I am a Director and Principal of the company George, Orr and Associates (Australia) Pty Ltd, of 11 Southport Street, Leederville, Perth, 6007
3. I am a practising geotechnical engineer and a member of the Australasian Institute of Mining and Metallurgy and the Australian Institute of Geoscientists.
4. I am a graduate of the University of Natal, Republic of South Africa and hold BSc, BSc (Hons) and MSc degrees.
5. I have practiced my profession continuously since 1973.
6. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
7. I have visited the Kapulo Copper Project property in November 2008. I have performed consulting services and reviewed files and data supplied by Mawson West Ltd between 2009 and the present.
8. I am responsible for Section 18.1.2.
9. 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.
10. I am not independent of Mawson West Ltd pursuant to section 1.4 of the Instrument.
11. 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.
12. I own Ordinary Fully Paid Shares in Mawson West Ltd, issued in March 2009 as part payment of consulting fees. Apart from these shares, I do not have nor do I expect to receive a direct or indirect interest in the Kapulo property of Mawson West Ltd and do not own, directly or indirectly, any other securities of Mawson West Ltd or any associate or affiliate of such companies.
Dated at Perth, Western Australia, on 30 June 2011.
[Signed] Chris Orr MSc, MAusIMM, MAIG Principal Engineering Geologist (Chris Orr and Associates)
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Sedgman Ltd.
Certificate of Qualified Person
As an author of the report entitled “Kapulo Copper Project, National Instrument 43-101 Technical Report” dated 30 June 2011, on the Kapulo property of Mawson West Ltd (the “Study”), I hereby state:
1. My name is Peter George Hayward and I am a Senior Process Engineer with the firm of Sedgman Ltd. of Suite 3, 3 Craig Street, Burswood, Western Australia, 6100.
2. I am a practising metallurgist and a member of the Australian Institute of Mining and Metallurgy.
3. I have practiced my profession continuously since 1974.
4. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
5. I have not visited the Kapulo Copper Project property. I have performed consulting services and reviewed files and data supplied by Mawson West Ltd in 2010.
6. I prepared Section 16 of the Study, as well as the associated text in the summary, conclusions and recommendations.
7. I am not aware of any limitations imposed upon my access to persons, information, data or documents that I consider relevant to the subject matter of the Study.
8. I am not aware, as of the date of this Certificate, of any material fact or material change with respect to the subject matter of the Study, which is not reflected in the Study, the omission of which would make the Study misleading.
9. I am independent of Mawson West Ltd and AMC SPRL. pursuant to section 1.4 of the Instrument.
10. I have not had any prior involvement with the properties that are subject of the Study, except as described in paragraph 6 above.
11. 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.
12. I do not have nor do I expect to receive a direct or indirect interest in the Kapulo property, 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 30 June 2011.
[Signed] Peter Hayward Senior Process Engineer
Sedgman Metals Engineering Services
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Sedgman Ltd.
Certificate of Qualified Person
As an author of the report entitled “Kapulo Copper Project, National Instrument 43-101 Technical Report” dated 30 June 2011, on the Kapulo property of Mawson West Ltd (the “Study”), I hereby state:
1. My name is Aaron Wade Massey and I am a Senior Process Engineer with the firm of Sedgman Ltd. of Suite 3, 3 Craig Street, Burswood, Western Australia, 6100.
2. I am a practising metallurgist and a member of the Australian Institute of Mining and Metallurgy.
3. I have practiced my profession continuously since 1995.
4. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
5. I have not visited the Kapulo Copper Project property. I have performed consulting services and reviewed files and data supplied by Mawson West Ltd in 2010.
6. I prepared Section 18.2 and contributed to Section 18.4 through to Section 18.14, as well as the associated text in the summary, conclusions and recommendations.
7. I am not aware of any limitations imposed upon my access to persons, information, data or documents that I consider relevant to the subject matter of the Study.
8. I am not aware, as of the date of this Certificate, of any material fact or material change with respect to the subject matter of the Study, which is not reflected in the Study, the omission of which would make the Study misleading.
9. I am independent of Mawson West Ltd and AMC SPRL,. pursuant to section 1.4 of the Instrument.
10. I have not had any prior involvement with the properties that are subject of the Study, except as described in paragraph 6 above.
11. 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.
12. I do not have nor do I expect to receive a direct or indirect interest in the Kapulo property, 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 30 June 2011.
[Signed] Aaron Massey Senior Process Engineer
Sedgman Metals Engineering Services
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