Husab Project, Namibia National Instrument 43-101 Technical Report
Husab Uranium Project - May 2011 Project Update
Prepared by Coffey Mining Pty Ltd on behalf of: Extract Resources Limited
Effective Date: 20th May 2011 Qualified Person : Neil Inwood - BSc (Geol.), PGradDip (Geol), MSc, MAusIMM Steve Le Brun MAusIMM Steve Craig – AusIMM Ross Cheyne - M AusIMM Mike Valenta - M SAIMM Hugh Browner - F SAIMM Steve Amos - F SAIMM
MINEWPER00713AE
Coffey Mining Pty Ltd DOCUMENT INFORMATION
Author(s): Coffey Mining Pty Ltd Neil Inwood Principle Consultant MSc, MAusIMM Steve Le Brun Principle Consultant M AusIMM
ORElogy Steve Craig Managing Director M AusIMM Ross Cheyne Director M AusIMM
Metallicon Mike Valenta Managing Director M SAIMM
AMEC Minproc Hugh Browner Engineering Manager F SAIMM Steve Amos Technical Manager F SAIMM
Location: Husab Uranium Project, Namibia
Date: May 2011
Project Number: MINEWPER00713AE
Version / Status: Final
Path & File Name: F:\MINE\Projects\Extract Resources\MINEWPER00713AE(continuation of 713AD)_Rossing South Res Modelling & Ongoing Tech Serv\Report\CMWPr_713AE_Extract_43-101_20May2011_SEDAR.docx
Print Date: Friday, 20 May 2011
Copies: Extract Resources Limited (2) Coffey Mining – Perth (1)
Document Change Control
Version Description (section(s) amended) Author(s) Date
Document Review and Sign Off
[signed] Principal Consultant Resources Neil Inwood
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
Table of Contents
1 Summary ...... 1 1.1 Property ...... 1 1.2 Location ...... 1 1.3 Ownership ...... 1 1.4 Geology, Mineralisation and Resources ...... 1 1.5 Exploration Concept ...... 2 1.6 Reserve Estimate and Mining ...... 4 1.7 Metallurgical ...... 5 1.8 Definitive Feasibility Study (DFS) ...... 8 1.9 Exploration Status ...... 9 1.10 Conclusions and Recommendations ...... 9 2 Introduction and Terms of Reference ...... 11 2.1 Scope of the Report ...... 11 2.2 Principal Sources of Information and Scope of Inspection ...... 11 2.3 Participants, Qualifications and Experience ...... 12 2.4 Independence ...... 13 2.5 Abbreviations ...... 13 3 Reliance on Other Experts ...... 15
4 Property Description and Location ...... 18 4.1 Background Information on Namibia ...... 18 4.1.1 Demographics and Geographic Setting ...... 18 4.1.2 History and Political Status ...... 19 4.1.3 Infrastructure ...... 19 4.1.4 Industry ...... 19 4.1.5 Mining ...... 20 4.2 Mineral Tenure ...... 20 4.2.1 Non-Exclusive Prospecting Licenses (NEPL) ...... 20 4.2.2 Reconnaissance Licenses (RL) ...... 21 4.2.3 Exclusive Prospecting License (EPL) ...... 21 4.2.4 Mineral Deposit Retention Licenses (MDRL) ...... 21 4.2.5 Mining Licenses ...... 21 4.3 Project Locations and Land Area ...... 21 4.4 Agreements and Encumbrances ...... 25 4.4.1 Environmental Liabilities and Permits ...... 25 5 Accessibility, Climate, Physiography, Local Resources and Infrastructure ...... 27 5.1 Access ...... 27 5.2 Climate ...... 27 5.2.1 Topography, Elevation and Vegetation ...... 27 5.2.2 Local Resources and Infrastructure ...... 28 5.3 Future Mining Operations ...... 29 6 History ...... 30 6.1 Ownership History ...... 30
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
6.2 Exploration History ...... 30 6.3 Resource History ...... 31 6.3.1 Production History ...... 32 7 Geological Setting ...... 33 7.1 Regional and Local Setting ...... 33 7.2 Project Geology...... 36 7.2.1 Husab Project ...... 36 8 Deposit Types ...... 38
9 Mineralisation ...... 39 9.1 Primary Uranium Mineralisation...... 39 9.2 Secondary Uranium Mineralisation ...... 40 10 Exploration ...... 41 10.1 Husab Uranium Project, Zones 1 to 4 ...... 41 11 Drilling ...... 46 11.1 Introduction ...... 46 11.2 Diamond Core Drilling ...... 46 11.3 RC Drilling ...... 47 11.4 Drilling Orientation ...... 48 11.5 Drilling Results ...... 48 11.6 Drilling Quality – Drillhole Database Verification ...... 49 12 Sampling Method and Approach ...... 50 12.1 Diamond Core Sampling ...... 50 12.2 RC Sampling and Logging ...... 51 12.2.1 Resource Drilling ...... 51 12.2.2 Exploration Drilling ...... 52 12.2.3 QAQC Sampling ...... 52 12.3 Sampling Quality ...... 52 13 Sample Preparation, Analyses and Security ...... 53 13.1 Sample Security ...... 53 13.2 Analytical Laboratories ...... 53 13.3 Sample Preparation and Analytical Procedure ...... 53 13.4 Bulk Density Determinations ...... 55 13.5 Radiometric Downhole Assaying ...... 56 13.6 Adequacy of Procedures ...... 56 14 Data Verification ...... 57 14.1 Standards and Blanks ...... 57 14.1.1 Extract Submitted Standards and Blanks ...... 57 14.1.2 Laboratory Blanks and Standards ...... 59 14.2 Duplicates ...... 60 14.3 Data Quality Summary ...... 61 15 Adjacent Properties ...... 62 15.1 Rössing Uranium Mine ...... 63 15.2 Etango Deposit...... 63 15.3 Langer Heinrich Deposit ...... 63
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
16 Mineral Processing and Metallurgical Testing ...... 64 16.1 Introduction ...... 64 16.2 Metallurgical Batch Testwork ...... 65 16.2.1 Sample Selection and Composite Formation ...... 66 16.2.2 Head Assays ...... 66 16.2.3 Mineralogy ...... 66 16.2.4 Comminution ...... 67 16.2.5 Dilute Leach Testwork ...... 67 16.2.6 Agitated Leach Testwork ...... 68 16.2.7 Flotation ...... 70 16.2.8 Heap Leach Amenability ...... 70 16.2.9 Ion Exchange ...... 70 16.2.10 Solvent Extraction...... 71 16.2.11 Precipitation ...... 72 16.2.12 Radiometric Ore Sorting ...... 72 16.2.13 Other Testwork ...... 73 16.3 Conclusions from Batch Laboratory Testwork...... 74 16.4 Metallurgical Pilot Plant Testwork ...... 75 16.4.1 Introduction ...... 75 16.4.2 Ore Preparation ...... 75 16.4.3 Batch Dry Milling...... 76 16.4.4 Re-pulp ...... 76 16.4.5 Leaching ...... 76 16.4.6 Dilution and Screening ...... 77 16.4.7 Fines Thickening ...... 77 16.4.8 Underflow Re-leach ...... 77 16.4.9 Filtration of Leach-End Residue ...... 77 16.4.10 Continuous Ion Exchange ...... 78 16.4.11 Solvent Extraction...... 79 16.4.12 Recovery and Refinery ...... 81 16.4.13 Vendors ...... 82 16.4.14 F.L. Smidth – Filtration ...... 82 16.4.15 F.L. Smidth – Thickening Testwork ...... 83 16.4.16 Larox ...... 86 16.4.17 Delkor ...... 87 16.4.18 RPA Filtres Philippe ...... 89 16.5 Comminution Pilot Plant Testwork ...... 90 16.5.1 Introduction ...... 90 16.5.2 Sample Preparation ...... 91 16.5.3 Summary of Results ...... 91 16.5.4 Conclusion ...... 92 16.6 Development of Flowsheet Based on Testwork ...... 92 16.7 Key Technical Features ...... 94 16.8 Process Plant ...... 94 16.8.1 Crushing ...... 94 16.8.2 Milling ...... 94 16.8.3 Leach ...... 94 16.8.4 Solid-l Liquid Separation ...... 96
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
16.8.5 Ion Exchange ...... 96 16.8.6 Solvent Extraction...... 96 16.8.7 Precipitation ...... 96 16.8.8 Tailings ...... 96 17 Mineral Resource Estimates ...... 97 17.1 Zone 1 Resource Estimate ...... 97 17.1.1 Resource Database and Validation ...... 97 17.1.2 Geological Interpretation and Modelling ...... 98 17.1.3 Radiometric Data ...... 101 17.1.4 Statistical Analysis of Composites and Top Cuts...... 101 17.1.5 Bulk Density Data ...... 103 17.1.6 Variography ...... 104 17.1.7 Block Model ...... 107 17.1.8 Grade Estimation ...... 107 17.1.9 Resource Reporting and Classification ...... 111 17.2 Zones 2 to 4 Resource Estimates ...... 112 17.2.1 Resource Database and Validation ...... 112 17.2.2 Geological Interpretation and Modelling ...... 113 17.2.3 Statistical Analysis of Composites and Top Cuts...... 119 17.2.4 Block Model ...... 125 17.2.5 Grade Estimation ...... 126 17.2.6 Comments and Recommendations ...... 134 18 Mineral Reserve Estimates and Mining Methods ...... 135 18.1 Mining operations ...... 135 18.1.1 Mining Methodology ...... 135 18.2 Open Pit Optimisation ...... 136 18.2.1 Resource Model ...... 137 18.2.2 Density ...... 137 18.2.3 Ore Loss and Dilution ...... 138 18.2.4 Bench Height Selection ...... 139 18.2.5 Geotechnical Parameters ...... 140 18.2.6 Mining Costs ...... 142 18.2.7 Processing Costs, Mill Throughput and Timing ...... 144 18.2.8 Processing Recovery and Recovery Ramp-Up ...... 145 18.2.9 Price, Selling Cost and Royalties...... 145 18.2.10 Capital Expenditure and Discount Rate ...... 146 18.2.11 Whittle Optimisation Process ...... 147 18.2.12 Whittle Optimisation Results and Shell Selection ...... 148 18.2.13 Sensitivity ...... 150 18.2.14 Ore Tonnage Sensitivity ...... 152 18.2.15 NPV Sensitivity on Best Case Shell...... 153 18.2.16 Shell Geometry and the Effect of Inferred Mineralisation ...... 154 18.3 Pit Designs ...... 159 18.3.1 Mineable Ore Reserve Estimates ...... 162 18.3.2 Mining Fleet ...... 162 18.3.3 Compliance with Whittle ...... 163 18.4 Reconciliation to Final DFS Parameters...... 164
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
19 Energy, Water and Process Materials ...... 167 19.1 Water Supply ...... 167 19.2 Power Supply ...... 168 19.3 Project Supply Chain And Logistics ...... 169 20 Infrastructure ...... 171 20.1 Access Roads ...... 171 20.2 Port Facilities ...... 171 20.3 Housing ...... 172 20.4 Personnel Transport ...... 172 21 Environmental Studies, Permitting and Social or Community Impact ...... 174 21.1 Location ...... 174 21.2 Legislation Governing Uranium Mining In Namibia ...... 174 21.3 The Impact Assessment Process ...... 175 21.4 An Overview of The Project Area ...... 176 21.5 Major Issues of Concern That Were Identified and Recommended Mitigation Measures ...... 177 21.6 Conclusion ...... 178 22 Market Studies and Contracts ...... 179 22.1 Demand for Uranium...... 179 22.2 Marketing of Uranium ...... 179 22.3 Taxation...... 180 23 Capital and Operating Costs ...... 181 23.1 Cost Estimates ...... 181 23.1.1 Capital Cost Estimate ...... 181 23.1.2 Operating Cost Estimate ...... 182 24 Economic Analysis ...... 183 24.1 Taxation...... 183 24.2 Economic Analysis – Results ...... 183 24.3 Economic Analysis - Sensitivity ...... 184 24.4 Mine Life Extension ...... 186 25 Interpretation and Conclusions ...... 187
26 Recommendations ...... 189
27 References ...... 191
28 Date and Signature Page ...... 195
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
List of Tables
Table 1.4_1 – Husab Uranium Project Zones 1 to 4, August 6 2010 Resource Estimate 2 March, 2011 DFS Reserves 3 Table 1.6_1 – Husab Uranium Project Zone 1 and Zone 2 March, 2011 DFS Reserves 4 Table 1.6_2 – Husab Uranium Project Zones 1 and 2 Ultimate Design Open Pit Dimensions 4 Table 2.5_1 – List of Abbreviations 14 Table 4.2.5_1 – Namibian Mineral Exploration and Mining Rights 22 Table 4.3_1 –Tenement Schedule 24 Table 6.3_1 – Husab Uranium Project, Namibia - August 2008 Resource Estimates 31 Table 6.3_2 – Superseded August 2010 Zone 1 and Zone 2 Resource Estimates 32 Table 10.1_1 – Husab Uranium Project Zone 1 and Zone 2- Significant Drilling Intersections 44 Table 10.1_2 – Husab Uranium Project - Zone 2- Significant Drilling Intersections 45 Table 13.3_1 – Summary of Assaying by Laboratory for Au, Ag, Cu, Ni, U, Zn and SG 55 Table 14.2_1 – Summary of Uranium Precision Data 60 Table 16.2.11_1 – Uranium Product Analysis 73 Table 16.4.14_1 – Husab Vacuum Filtration Results 83 Table 16.4.16_1 – Thickening Testwork Summary 84 Table 16.4.16_2 – Thickening Testwork Summary 85 Table 16.4.17_1 – Soluble Recoveries - Factor 1 Displacement Wash 88 Table 16.4.17_2 – Soluble Recoveries - 2 Stage Counter-Current Wash 89 Table 17.1.4_1 – Statistics and Top Cuts Applied to the Various Mineralised Units Zones 103 Table 17.1.5_1 – Density Readings taken from Drill Core at Husab Uranium Project 104 Table 17.1.6_1 – Grouped Zone Variography – Husab Uranium Project Zone 1 104 Table 17.1.7_1 – Block Model Parameters – Husab Uranium Project Zone 1 107 Table 17.1.8_1 – Sample Search Parameters for OK Estimate – Husab Uranium Project Zone 1 108
Table 17.1.8_2 – Comparison of Model Grades and Informing Composite U 3O8 Grades for Mineralised Units 109 Table 17.1.8_3 – Density Values Applied to the Various Rocktypes within the Resource Model 111 Table 17.1.9_1 – Confidence Levels of Key Categorisation Criteria 112 Table 17.1.9_2 – Husab Uranium Project Zone 1 – August 6 2010 Resource Estimate 112 Table 17.2.1_1 – Husab Uranium Project (Zone 2 - 4), Number, Metres Drilled and Type of Hole 113 Table 17.2.3_1 – Statistics and Top Cuts Applied to the Various Mineralised Zones – Zones 2 to 4 120 Table 17.2.3_2 – Bulk Density Values Assigned by Supplied Lithological Wireframes 121 Table 17.2.3_3 – Variogram Parameters for the Mineralised Zone 2 122 Table 17.2.3_4 –Variogram Parameters for the Mineralised Zones 3 and 4 123 Table 17.2.4_1 – Block Model Parameters, Zones 2 and 4 125 Table 17.2.4_2 – Block Model Variables, Zones 2 and 4 125 Table 17.2.5_1 – Sample Search Parameters, Ordinary Kriging - Husab Uranium Project (Zone 2) 130 Table 17.2.5_2 – Sample Search Parameters, Ordinary Kriging – Husab Uranium Project (Zones 3 and 4) 131 Table 17.2.5_3 – Confidence Levels of Key Categorisation Criteria – Zones 2 to 4 132 Table 17.2.5_4 – Husab Uranium Project - Zones 2 to 4 Resource Estimate 133 Table 17.2.5_5 – Husab Uranium Project - August 6 2010 Resource Estimate, All Zones 133 Table 18.2.2_1 – Material Density 138
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
Table 18.2.3_1 – Global Ore Loss and Dilution by Zone 139 Table 18.2.5_1 – Geotechnical Slope Design Parameters 141 Table 18.2.2_2 – Ramp Width Calculation 141 Table 18.2.5_3 – Overall Slope Calculations 141 Table 18.2.6_1 – Global Average Unit Mining Costs (US$/t) 143 Table 18.2.7_1 – Process Throughput and Costs 144 Table 18.2.8_1 – Process Losses 145 Table 18.2.8_2 – Process Recovery Ramp-Up 145 Table 18.2.10_1 – Capital Costs (US$ M) 147 Table 18.2.12_1 – Summary of Selected Shell 36 150 Table 18.2.13_1 – Sensitivity Ranking for a 15% Variation in Key Parameter 152 Table 18.2.13_2 – Sensitivity Summary 152 Table 18.2.13_3 – Economic Cutoff Grade 153 Table 18.3.1_1 – Husab Probable Ore Reserve by Stage Designs 162 Table 18.3.3_1 – Correlation between Pit Designs and Original Optimal Shell 36 163 Table 18.3.3_2 – Correlation between Pit Designs and Revised Optimal Shell 36 164 Table 18.4_1 – Variation in Key Parameters Optimisation to Final DFS 166 Table 19.3_1 – Reagent Supply Quantities 170 Table 23.1.1_1 – Estimated Capital Cost 181 Table 23.1.2_1 – Estimated Operating Costs 182 Table 24.2_1 – Forecast Cashflows (Reserves only) 185 Table 24.2_2 – Project NPV (Reserves only) 186 Table 24.3_1 – Valuation Sensitivity 186 Table 27_1 – Estimated Budget for Recommendations for 2011/2012 190
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
List of Figures
Figure 1.6_1 – Husab Uranium Project Mine Site Layout 5 Figure 4.1.1_1 – Geography of Namibia 18 Figure 4.3_1 – Project Location Map 23 Figure 5.2.1_1 –Desert Plains in the Rossing South Project Area 28 Figure 7.1_1 – Damara Orogenic Belt : Regional Geological Setting 33 Figure 7.1_2 – Stratigraphic Column of the Damara Orogen 35 Figure 7.2.1_1 – Husab Project: Geological Setting 36 Figure 7.2.1_2 – Husab Uranium Project Zone 1 Geological Cross Section (7506000N) 37 Figure 7.2.1_3 – Husab Project: Tenement Outline over First Vertical Derivative of Magnetics 37 Figure 9.1_1 – Beta-Uranophane Mineralisation in Hole RDD002 39 Figure 10.1_1 – Husab Uranium Project: Significant Drilling Intercepts 43 Figure 10.1_2 – Husab Uranium Project Mineralisation from Hole RDD002 45 Figure 11.2_1 – Husab Uranium Project Core – Hole RD002 46 Figure 11.3_1 – RC Drilling at Husab Uranium Project 48 Figure 12.3.1_1 – Husab Uranium Project RC Drilling 51 Figure 13.3_1 – Flow Diagram - Sample Preparation at the Genalysis Preparation Facility in Johannesburg, South Africa 54 Figure 14.1.1_1 – Extract Submitted Standards and Blanks 58 Figure 14.1.2_1 – Laboratory Submitted Standards and Blanks 59 Figure 15_1 – Location of Uranium Deposits on Properties Adjacent to the Extract Tenement Holdings 62 Figure 16.8_1 – DFS Flowsheet 95 Figure 17.1.2_1 – Husab Uranium Project Zone 1 - Mineralised Zones and Drill Type 99 Figure 17.1.2_2 – Husab Uranium Project Zone 1, Sectional Interpretation (7,506,800mN) 100 Figure 17.1.4_2 – Histogram Plot from Mineralised Zones 2 to 4 102 Figure 17.1.6_1– Grouped Domains – Husab Uranium Project Zone 1 105 Figure 17.1.6_2– Modelled Variography – Husab Uranium Project, Zone 1 106 Figure 17.1.8_1 – Example Northing Validation Plots – Husab Uranium Project Zone 1 110 Figure 17.1.9_1 – Oblique View of the Classified Zone 1 Resource Model and Drillholes 111 Figure 17.2.2_1 – Husab Uranium Project, Zones 1 – 4 - Drillhole Location Plan 114 Figure 17.2.2_2 – Husab Uranium Project Zone 2 Mineralised Zones and Drill Type 115 Figure 17.2.2_3 – Husab Uranium Project Zones 3 and 4 - Mineralised Zones and Drill Type 116 Figure 17.2.2_4 – Husab Uranium Project Zone 2 - Sectional Interpretation (7,503,600mN) 117 Figure 17.2.5_5 – Husab Uranium Project Zone 3 - Sectional Interpretation (7,501,600mN) 118 Figure 17.2.3_1 – Cross-Section of Final Block Model at 7503500mN 121 Figure 17.2.3_2 – Correlogram for Husab Uranium Project Zone 2 Combined Lodes 124 Figure 17.2.5_1 – Comparative Plots of Informing Composites and Block Model Grade –Zone 2 127 Figure 17.2.5_2 – Comparative Plots of Informing Composites and Block Model Grade – Zone 3 128 Figure 17.2.5_3 – Comparative Plots of Informing Composites and Block Model Grade – Zone 4 129 Figure 18.1.1_1 – Trolley-Assist Layout 135 Figure 18.2.3_1 – Ore Loss Dilution Geometry (Orebody Dip ≥ 10°) 139 Figure 18.2.5_1 – Whittle Slope Areas 142
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
Figure 18.2.7_1 – Process Throughput Ramp-Up 144 Figure 18.2.12_1 – Base Optimisation Results for 15Mtpa Throughput Rates 149 Figure 18.2.13_1 – % Variation from Base Case - Ore Tonnes 151 Figure 18.2.13_2 – % Variation from Base Case - Best Case Discounted Cashflow 151 Figure 18.2.13_3 – Grade Tonnage Curve for Resource 154 Figure 18.2.16_1 – Zone 1 Cross-Section 7507000 N 155 Figure 18.2.16_2 – Zone 1 Cross-Section 7506500 N 155 Figure 18.2.16_3 – Zone 1 Cross-Section 7506000 N 156 Figure 18.2.16_4 – Zone 1 Cross-Section 7505300 N 156 Figure 18.2.16_5 – Zone 1 Cross-Section 7504000 N 157 Figure 18.2.16_6 – Zone 1 Cross-Section 7503500 N 157 Figure 18.2.16_7 – Zone 1 Cross-Section 7503000 N 158 Figure 18.3_1 – Zone 1 Stage Layouts 160 Figure 18.3_2 – Zone 1 Stage Layouts 160 Figure 18.3_3 – Mine Site Layout 161 Figure 19.1_1 – Water Supply to Husab Mine 167 Figure 19.2_1 – Modified 220kV Supply 169 Figure 20.1_1 – Alternative Routes Considered 172
List of Appe ndices
Appendix 1 – QAQC Report Appendix 2 – Certificates of Qualified Persons
Husab Uranium Project, Namibia – MINEWPER00713AE 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
1 SUMMARY
In April 2011, Coffey Mining Pty Ltd (Coffey Mining) was commissioned by Extract Resources Ltd. (ASX:EXT, TSX:EXT) (Extract) to prepare an updated Independent Technical Report (ITR) on the uranium Resources and Reserves and the associated mining studies of the Husab Uranium Project (previously referred to as Rössing South - Zone 1 and 2 Prospects) in Namibia, Southern Africa.
Coffey Mining had previously prepared the ITR on Extract’s Husab Uranium Project in 2009
that estimated an Indicated mineral resource of 21Mt at 527ppm U3O8 and an Inferred
Resource of 126Mt at 436ppm U3O8, at Zone 1; and at Zone 2 an Inferred Resource of 102Mt
at 543ppm U3O8. All resources are reported above a 100ppm U3O8 lower cutoff.
1.1 Property
Extract Resources Ltd controls a portfolio of uranium properties in western Namibia through its wholly owned Namibian subsidiaries Extract Resources Namibia (Pty) Ltd (Extract Namibia) and Swakop Uranium (Pty) Ltd. The Husab Uranium Project (formerly known as Rössing South) is located within the broader Husab Project area, which covers an area of approximately 637km².
1.2 Location
The Husab Project area comprises two granted Exclusive Prospecting Licenses (EPL 3138 and 3439) and is located in the Namib Desert, approximately 50km east of Swakopmund in central western Namibia.
The Zone 1 and 2 Prospects are the northern-most region of the Husab Uranium Project which is located in the northern portion of EPL 3138. The Husab Uranium Project is approximately 8km south of the Rössing Uranium Mine. The Husab Uranium Project forms part of the broader Husab Project area.
1.3 Ownership
The Husab Project EPL 3138 and EPL 3439 are owned by Swakop Uranium (Pty) Ltd. The tenement schedule is included as Table 4.4_1.
1.4 Geology, Mineralisation and Resources
The Husab Project (EPLs 3138 and 3439) is situated within the central Damara Orogenic Belt (DOB). The area is dominated by a series of north-northeast to northeast trending regional- scale antiforms and synforms, which make up the main structural architecture of the entire Central Zone of the Damara. These meta-sedimentary folds or dome-like structures of the DOB are cored by gneissic and metasedimentary rocks of the Abbabis Formation. The basement rocks are covered to the northeast and south by stranded cover sequences of flat- lying calcrete and alluvial deposits, which are associated with a broad northeast trending valley marginal to the Khan River. Regional magnetic data indicate that the regional structural history is complex.
Husab Uranium Project, Namibia – MINEWPER00713AE Page: 1 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
The Husab Project area contains primary uranium mineralisation hosted in uraniferous leucocratic granites (alaskites) within the highly prospective Central Zone of the Damara Orogeny. The mineralised alaskites occur mainly within the Rössing Formation but locally also intrude the Khan and Chuos Formations.
In addition to primary alaskite-hosted uranium mineralisation, the tenement holding also hosts occurrences of secondary uranium vanadate (carnotite).
Regional magnetic data show that the mineralised rock units can be traced beneath cover. Drilling completed at the Husab Uranium Project has confirmed the existence of high grade uranium mineralisation beneath these cover sequences for bedrock occurrences and for secondary calcrete uranium mineralisation within the cover rocks.
In August 2008, Coffey Mining completed resource estimates for alaskite hosted uranium mineralisation at the Garnet Valley, New Camp and Ida Central prospects. The resources
included a combined Indicated Mineral Resource of 0.6Mt at 246ppm U3O8 and Inferred
Mineral Resource of 52.7Mt at 213ppm U3O8 above a 100ppm U3O8 lower cutoff. These deposits are part of the Ida Dome project area.
In July 2009, Extract completed an updated resource estimate (now superseded) for Zone 1
of the Husab Uranium Project with an Indicated Resource of 21Mt at 527ppm U 3O8 and an
Inferred Resource of 126Mt at 436ppm U 3O8, above a 100ppm U3O8 lower cutoff. Also at this time, Coffey Mining completed a maiden resource estimate for Zone 2 with an Inferred Mineral
Resource of 102Mt at 543ppm U 3O8, above a 100ppm U3O8 lower cutoff.
The August 2010 resource update represents a significant increase in Indicated Mineral Resource material for Zones 1 and 2, and includes maiden resource estimates for Zones 3 and 4. The updated Resource estimates are tabulated below in Table 1.4_1. The preferred
cutoff grade for reporting is 100ppm U3O8.
The Definitive Feasibility Study (DFS) has been conducted on the basis of the August 2010 resource model, but has focussed only on the more advanced Zone 1 and Zone 2 deposits where Indicated Resources have been defined. Exploration activities remain ongoing and a further resource update for Zones 1, 2, 3 and 4 and a maiden resource for Zone 5 is scheduled for release in Q2, 2011.
1.5 Exploration Concept
Extract is targeting primary, alaskite-hosted uranium mineralisation at the Husab Project. Mineral Resource definition drilling is currently underway at the Husab Uranium Project Zone 1 and Zone 2. Extensive exploration potential remains throughout the Husab Project with the main area of immediate interest the 9km trend between the northern end of Ida Dome and the southern end of Husab Uranium Project Zone 2.
Husab Uranium Project, Namibia – MINEWPER00713AE Page: 2 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
Table 1.4_1 Husab Uranium Project Zones 1 to 4 August 6 2010 Resource Estimate
Reported at various cutoffs, Preferred cutoff : 100ppm U 3O8 Ordinary Kriged Estimate based upon 3m cut U 3O8 Composites Parent Cell Dimensions of 25m NS by 25mEW by 15mRL
Lower Cutoff Tonnage Grade Contained U 3O8 Contained U 3O8 (ppm U 3O8) (Mt) (ppm U 3O8) (MKg) (MLb) Zone 1 100 122.2 450 55.0 120.1 200 104.5 490 51.2 113.9 Indicated 300 78.0 580 45.3 99.3 400 56.2 670 37.6 82.6 100 41.3 420 17.4 37.8 200 34.2 470 16.1 35.2 Inferred 300 24.7 550 13.6 29.9 400 17.0 640 10.9 24.1 Zone 2 100 118.8 520 61.8 136.9 200 110.0 550 60.5 133.7 Indicated 300 87.7 630 55.3 121.2 400 66.3 720 47.8 104.8 100 26.8 520 13.9 30.5 200 24.5 550 13.5 29.7 Inferred 300 19.3 630 12.1 26.8 400 14.1 740 10.4 22.9 Zone 3 100 43.0 250 10.7 24.0 200 24.3 330 8.0 17.5 Inferred 300 11.8 410 4.9 10.7 400 4.3 530 2.3 5.0 Zone 4 100 14.4 570 8.2 17.9 200 13.0 610 7.9 17.5 Inferred 300 11.4 660 7.5 16.6 400 9.0 740 6.7 14.7 All Zones 100 241.0 480 117 257.0 200 214.5 520 112 247.6 Indicated 300 165.7 610 101 220.6 400 122.5 700 85 187.4 100 125.5 400 50 110.3 200 96.1 470 45 99.9 Inferred 300 67.2 570 38 84.0 400 44.4 680 30 66.7
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1.6 Reserve Estimate and Mining
The mineable reserves, based on pre-defined pit optimisation parameters, are summarised in Table 1.6_1.
Table 1.6_1 Husab Uranium Project Husab Uranium Project Zone 1 and Zone 2 March, 2011 DFS Reserves
Cutoff Grade Ore Grade* U O Probable Reserves 3 8 (ppm) (Mt) (ppm) (M lbs) Zone 1 148 97.1 477 102.2 Zone 2 138 107.8 515 122.5 Total 205.0 497 224.8 *Mineral Reserve tonnage and grade estimates include allowance for mining dilution and ore loss.
The DFS envisages conventional truck and shovel mining from two open pits focussing on mineralisation in Zones 1 and 2. Following initial pre-stripping, ore will be hauled to a primary crusher, positioned on surface to the west and between the two open pits. Waste rock removed from the pits will be hauled to a Mine Residue Facility (MRF) located to the east of the open pits which serves as a co-disposal facility for the process plant tailings. As appropriate during operation and at mine closure, the MRF will be capped with a layer of waste rock. The site layout is shown in Figure 1.6_1.
Based on the maiden Mineral Reserve estimate the mine life (including pre-strip) is 16 years. Extract expects to define additional Mineral Reserves to extend the mine life substantially beyond the current mine plan.
Both pits have been designed based on the outputs from the open pit optimisation simulation software Whittle FOUR-X. The current design pit dimensions are shown in Table 1.6_2.
Table 1.6_2 Husab Uranium Project Husab Uranium Project Zones 1 and 2 Ultimate Design Open Pit Dimensions
Length N-S Width E-W Depth Zone (km) (km) (m) 1 2.4 0.7 412 2 1.9 1.2 330
The primary loading fleet comprises three large electric-powered rope shovels to mine the bulk waste on 15m benches whilst a fleet of smaller diesel hydraulic shovels address all ore loading requirements on 7.5m benches in order to minimise ore loss and dilution. The smaller shovels are able to manage both a 7.5m and 15m bench height. In-pit blending will minimise the extent of re-handling of ore from stockpile to crusher to cater for short-term grade variations over life of mine.
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Figure 1.6_1 Husab Uranium Project Mine Site Layout
Large electric-powered drill rigs service the large electric loading units whilst the smaller and more manoeuvrable diesel rigs would similarly service the smaller diesel-powered loading units.
Ore and waste are transported by a fleet of 39 diesel electric drive haul trucks in the +300 tonne class. Trolley-assisted hauling has been included in the base case, and will be implemented on most up-ramp sections of the open pits and ramps accessing the MRF.
The remainder of the mining production fleet consists of support equipment that includes graders, track and wheel dozers, front-end loaders, rock breakers and utility excavators.
Specific mining activities are planned to be outsourced. These include the repair and maintenance of the mobile mining fleet, blasting operations, tyre and haul road management as well as drilling operations relating to grade control and resource definition.
Following initial pre-stripping of 85Mt overburden, the mine plan envisages production at a rate of 15Mt ore per year, at an average strip ratio of 7.0:1.
1.7 Metallurgical
The metallurgical testwork commenced with laboratory scale batch testwork conducted at a scoping level in July 2008 and continued through to a hydrometallurgical and comminution pilot plant testwork phase that commenced in April 2010 and concluded in November 2010.
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The following laboratory scale batch testwork was conducted on selected drill cores that were deemed to be representative of the ore body:
Head assays
Mineralogy
Comminution
Dilute leach testwork
Agitated leach testwork
Flotation
Heap leach amenability
Radiometric ore sorting
Ion exchange
Solvent extraction using ammonium hydroxide strip
Ammonium diuranate precipitation.
Comments on the outcomes of the laboratory-scale testwork programme include the following:
SAG mill amenability tests indicate the samples are in the softest quartile of all samples tested.
High uranium recoveries with >60% of the drillhole composite leach recoveries >90%.
The leach acid consumptions are generally below 25kg/t with excursions above this due to high Ca and Mg ore or high Fe releasing ore, which contributes to oxidation of excessive ferrous to ferric and a resultant increase in pyrolusite and acid consumptions.
Reducing the leach temperature from 35ºC to 40ºC to 25 ºC to 30ºC impacts on the leach kinetics but has minimal impact on the final leach recovery.
IX testwork shows good selectivity of uranium from other impurities with low levels of contaminants feeding downstream processing.
High quality precipitate produced from either H 2O2 precipitation or Ammonium Diuranate (ADU).
Heap leach amenability testing produced moderate recoveries of 64% to 68% at crush sizes of -6.3mm and -12.5mm.
Radiometric ore sorting and dense media separation (DMS) or heavy media separation (HMS) were determined to be ineffective in upgrading the ore.
Pre-concentration by flotation showed some promise, but the testwork data was not reproducible.
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The following aspects of the selected process were tested at pilot plant scale:
Leaching
Dilution and screening
Fines thickening
Thickener underflow re-leach
Pressure filtration of leach residue
Continuous ion exchange
Solvent extraction using sodium carbonate strip
Sodium diuranate precipitation
Uranyl peroxide precipitation
During the hydrometallurgical pilot plant phase, selected equipment vendors were invited to conduct equipment specific tests to develop design data.
Comminution pilot plant testwork was also conducted to confirm the selection of single stage semi-autogenous (SAG) milling as the preferred milling option.
The flowsheet used in the DFS is primarily derived from the flowsheet used in the hydrometallurgical pilot plant, the comminution pilot plant and the findings from a PLS option. There are three main differences to that flowsheet that are incorporated into the DFS flowsheet:
The inclusion of a comminution circuit
The replacement of the leach discharge screen and fines thickener with a coarse filtration and fines CCD wash circuit; and
The replacement of the SX carbonate strip process with the strong acid strip process.
Key technical features of the DFS flowsheet are:
Mineral Sizer (product top size 250mm) for primary crushing of the run-of-mine (ROM) ore followed by semi-autogenous (SAG) milling to produce a P80 of 780µm;
Atmospheric leach process with 14 hour residence time. Importing of sulphuric acid and pyrolusite as lixiviant and oxidant respectively;
Solid / liquid separation using belt filtration with option of counter current decantation for excessive fines. Leach residue deposited as filter cake for minimising water requirement;
Continuous ion-exchange using the NIMCIX technology and conventional solvent extraction (SX) uranium upgrading and refining process; and
Production of uranyl peroxide by precipitation with hydrogen peroxide following a strong acid strip process.
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1.8 Definitive Feasibility Study (DFS)
Extract has defined a base case mine plan and process plant design, including plans for delivery of the infrastructure necessary to support the project. The DFS has demonstrated the technical and economic viability of developing Husab, the world’s fifth-largest uranium-only deposit.
The DFS is based on:
Indicated Minerals Resources defined at Zones 1 and 2, based on the August 2010 resource model;
Open pit mining by truck and shovel from two separate pits to maintain a sustained rate of 15Mt pa over the life of mine with an average strip ratio of 7:1 (waste:ore);
A waste and plant tailings storage facility (the mine residue facility);
Ore crushing and overland conveying to a new processing facility employing milling, leaching, ion exchange, solvent extraction and precipitation plant and equipment to produce approximately 15 million lbs pa of U3O8 equivalent; and
Provision of temporary and permanent power and water supplies, access roads, temporary and permanent buildings and structures necessary to support the Project.
Capital costs for the Project are estimated at US$1,480 million, including initial mine fleet, process plant and supporting infrastructure. Inclusive of pre-strip and other pre-production operating costs of US$179 million, the Project Cost is estimated at US$1,659 million. This estimate excludes allowance for finished goods inventory in transit and held at conversion facilities, debtor payment terms, creditor payment terms, escalation, and financing costs (including fees and interest during construction).
Production costs are estimated at US$28.5/lb, excluding royalties, marketing and transport and cost escalation. Operating costs including royalties, marketing and transport are estimated at US$32.0/lb.
The accuracy provision for the DFS is ± 10%. Figures are expressed in US$ in real terms assuming a base date of 1 January 2011 unless otherwise stated.
Extract has engaged with potential customers to assess demand for production from the Husab Uranium Project, and has identified several possible strategic contracting opportunities. Extract is confident that it will become an attractive supplier to end-users, as a result of the Husab Uranium Project’s ability to offer geographic diversification and long term security of supply.
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1.9 Exploration Status
The region of the Husab Uranium Project (Zones 1 and 2) Prospects represent an advanced exploration project with Indicated and Inferred Mineral Resources being defined at Zone 1 and at Zone 2; and Inferred Resources defined for Zones 3 and 4.
1.10 Conclusions and Recommendations
The August 2010 resource estimate (Table 1.4_1) represents a significant increase in Indicated Mineral Resources relative to the previous July 2009 Resources for Zones 1 and 2, and now incorporates maiden resource estimates for Zones 3 and 4.
Potential remains to expand the Mineral Resource inventory at the Husab Uranium Project through extension of known deposits such as Zones 1, 2, 3 and 4 and definition of resources at prospects such as Zone 5, Middle Dome and Salem.
Coffey Mining has reviewed the drilling, sampling, assaying and field procedures used by Extract and consider them to be of high quality.
Extract has defined a base case mine plan and process plant design, including plans for delivery of the infrastructure necessary to support the project. The DFS has demonstrated the technical and economic viability of developing Husab into one of the largest uranium mines in the world, and supports a maiden reserve estimate for Zones 1 and 2 of the Husab Uranium Project.
The DFS defines a base case mine plan and process plant design, including plans for delivery of the infrastructure necessary to support the project. Several opportunities to add further value have been identified, including the proposed update of the resource model, mine plan optimization, and processing enhancements. Extract Resources has commenced a Mine Optimisation and Resource Extension Programme (MORE) to investigate these opportunities.
A resource update is planned for Q2, 2011 incorporating drilling completed until the end of January, 2011. The main focus has been infill drilling of Zone 1 and Zone 2 to define Measured Mineral Resources and increase the quantum of Indicated Mineral Resources, both by upgrading the classification of Inferred Mineral Resources within the current mine plan (and therefore not included within the reserve estimate) and by definition of additional resources.
Extract Resources also intends to continue its exploration programme in Zones 3, 4 and 5, Middle Dome, Salem, Ida Dome, and Pizzaro areas, which are not included in the DFS. Definition of additional reserves would be expected to add additional value and mine life to the project.
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Several potential process enhancements are being investigated and include the following:
Finer grind process: Potential to result in reduced leach residue grade which would result in increased recovery and simplified solid liquid separation circuit.
Elevated temperature acid leach: Potential to result in reduced leach residue grade which would result in increased recovery.
Direct IX or SX to replace Eluex: Potential to lower plant capital expenditure.
On-site acid production: Procurement of sulphuric acid represents a significant component of the project’s operating cost. The possible upside from on-site production of acid, including the potential to generate electricity and provide heat for an elevated temperature leach process, is being assessed.
Work on these value adding areas is planned to continue and to feature, as appropriate, in the final mine plan to be initiated during the project development phase.
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2 INTRODUCTION AND TERMS OF REFERENCE
2.1 Scope of the Report
Coffey Mining has been commissioned by Extract to prepare a National Instrument 43-101 (NI43-101) compliant Technical Report to update the resource estimation studies undertaken at Zone 1 to 4 of the Husab Uranium Project, based on the August, 2010 resource update. These prospects are located within Extract’s Husab Project area in central western Namibia within EPL 3138.
The requirement to prepare this Technical Report was the result of the first time definition of Mineral Reserves at the Husab Uranium Project. These Mineral Reserves were announced contained in a market release, issued by Extract on 5 th April 2011, regarding the Husab Uranium Project DFS.
Only the Husab Uranium Project Zones 1 to 4 will be discussed in this report. Coffey Mining has previously prepared a Technical Report in August 2008 (Inwood, 2008) for the Garnet Valley, New Camp and Ida Dome prospects which are also located within the Husab Project area.
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 2004. 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).
All monetary amounts expressed in this report are in United States of America dollars (US$) unless otherwise stated.
2.2 Principal Sources of Information and Scope of Inspection
In addition to site visits undertaken to the Husab Uranium Project in 2007, 2008 and 2010 by the primary author, this report has relied extensively on information provided by Extract and extensive discussions with Extract technical personnel. 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 listing of the principal sources of information is included in the references section, Section 27 of this report and in Section 3.
Coffey Mining and the other authors have 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 Extract along with a written request to identify any material errors or omissions prior to lodgement.
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2.3 Participants, Qualifications and Experience
Coffey Mining was responsible for the preparation of all portions of this report apart from Sections 5.3, 16, 18, 19, 20, 23 and 24 and the associated text in the summary, conclusions and discussion. As per Section 3, Coffey Mining has relied upon the expert advice of others for Sections 21 and 22.
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 Neil Inwood, who is a professional geologist with 17 years experience in exploration and mining geology. Mr Inwood is a Principle Resource Consultant with Coffey Mining 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 Inwood visited the Husab Projects in August 2007, August 2008 and August 2010.
Mr Steve Le Brun is a professional geologist with 25 years experience in exploration and mining geology, prepared Section 17.2 and associated text in the summary, conclusions and discussion. Mr Le Brun is a Principle Resource Geologist with Coffey Mining 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 Corley has not visited the Husab Project site.
Mr Mike Valenta is a professional engineer as registered on the International register as defined by the Washington Accord with 20 years experience in metallurgical aspects of mining projects, prepared Section 16 and associated text in the summary, conclusions and discussion. Mr Valenta is The Managing Director of Metallicon Process Consulting and a Member of the South African Institute of Mining and Metallurgy (SAIMM), past president of the Metallurgical Mine Manager’s Association (MMMA) 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 Valenta has not visited the Husab Project site.
Mr Ross Cheyne is a professional mining engineer with 20 years experience in open pit mining, prepared Section 18 and the portions of Section 23 that relate to Mining costs, and associated text in the summary, conclusions and discussion. Mr Cheyne is a Director and Principal Consultant with ORElogy and a Member of the 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 Cheyne has not visited the Husab Project site. As per Section 3, Mr Craig has relied upon the expert advice of others for Section 24.
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Mr Steve Craig is a professional mining engineer with 24 years experience in mine planning, consulting and study management, prepared Section 18 and the parts of Section 23 that relate to Mining costs, and associated text in the summary, conclusions and discussion. Mr Craig is the Managing Director of ORElogy and is a Member of the 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 Craig has visited the Husab Project site in June 2009 and in May 2010. As per Section 3, Mr Craig has relied upon the expert advice of others for Section 24.
Mr Hugh Browner is a professional engineer, with 11 years experience in operations and 25 years in design and has, reviewed Sections 5.3, 19 and 20 and associated text in the summary, conclusions and discussion. Mr Browner is the Engineering Manager of AMEC Minproc South Africa and is a Fellow of the Southern African Institute of Mining and Metallurgy, and has the relevant experience to be generally considered as a “Qualified Person” as defined in the Canadian National Instrument 43-101. Mr Browner has not visited the Husab Project.
Mr Steve Amos is a Metallurgist, with 21 years experience in process engineering, design, management, commissioning and R&D, reviewed Section 23 relating to the processing costs only and associated text in the summary, conclusions and discussion. Mr Amos is Technical Manager of AMEC Minproc and is a Fellow of South African Institute of Mining and Metallurgy, and has the relevant experience to be generally considered as a “Qualified Person” as defined in the Canadian National Instrument 43-101. Mr Amos has not visited the Husab Project.
Certificates for the Competent Persons are located in Appendix 2.
2.4 Independence
The Authors of this report do not have or have had previously any material interest in Extract Resources Limited or related entities or interests. Our relationship with Extract 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.
2.5 Abbreviations
A listing of abbreviations used in this report is provided in Table 2.5_1 below.
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Table 2.5_1
List of Abbreviations
$ United States of America dollars hr hours P80 80% of sample passing “ inches HRD half relative difference Pa. Pascal µm microns IBC Intermediate Bulk Container Pd palladium 3D three dimensional ICP-AES inductivity coupled plasma atomic emission spectroscopy PGE Platinum Group Elements -75µm 75 microns ICP-MS inductivity coupled plasma mass spectroscopy PLS Pregnant Liquor Solution AAS atomic absorption spectrometry IRR Internal Rate of Return ppb parts per billion AC Alternating Current ISO International Standards Organisation ppm parts per million ADU ammonium diuranate IX Ion Exchange ppt. precipitate Ag Silver JORC Joint Ore Reserves Committee PQ size of diamond drill rod/bit/core ANSTO Australian Nuclear Science & Technology Organisation K Potassium psi pounds per square inch ASX Australian Stock Exchange kg kilogram PVC poly vinyl chloride AusIMM Australasian Institute of Mining and Metallurgy kg.d.s/m²hr kilogram droge stof (dry weight) per square metre per hour Q2 second quarter BBMWI Bond Ball Mill Work Index kg/h kilogrammes per hour QC quality control bcm bank cubic metres kg/t kilogram per tonne QEMSCAN Quantitative Evaluation of Minerals by Scanning Electron Microscopy BRMWI Bond Rod Mill Work Index km kilometres QQ quantile-quantile C Carbon km² square kilometres RAB Rotary Air Blast Ca Calcium ktonnes thousand tonnes Rb Rubidium CC correlation coefficient kV kilovolt RC reverse circulation CCD Counter Current Decantation kW kilowatts RL (Z) reduced level CIX Continuous Ion Exchange kWhr/t kilowatt hours per tonne ROM run of mine cm centimetre L litre RQD rock quality designation Co cobalt l/hr/m² litres per hour per square metre SABC SAG Mill + Ball Mill + Crushing COG Cutoff Grade LAS Log Ascii Standard SAG Semi autogenous grinding CRM certified reference material or certified standard lb pounds SAIEA South African Institute of Environmental Cu copper LOM Life of Mine SD standard deviation CV coefficient of variation M Million SDU sodium diuranate DDH diamond drillhole M Mole/molar SEA Strategic Environmental Assessment DFS Definintive Feasibility Study m³/t cubic metres per tonne SEMP Strategic Environmental Management Plan DGPS Differential Global Positioning System Ma thousand years SG Specific gravity DMS Dense Media Separation MET Ministry of Environment and Tourism Si silica DOB Damara Orogenic Belt Mg Magnesium SMU selective mining unit DTM digital terrain model mg/L milligrammes per litre Sr Strontium E (X) easting ml millilitre SX Solvent Extraction ECS Engineering Cost Study Mlb Million pounds t tonnes EDC Erongo Desalination Company mm millimetres t/h tonnes per hour EDM electronic distance measuring Mm³/a Million cubic metres per year t/hr tonnes per hour EIA Environmental Impact Assessment MME Ministry of Mining and Energy t/m³ tonnes per cubic metre EMP Environmental Management Plan Mn Manganese Ta Tantalum EPCM Engineering Procurement Construction Management MRF Mine Residue Facility Th Thorium EPL Exclusive Prospecting Licence Mt Million Tonnes tpa tonnes per annum
eU 3O8 equivalent U 3O8 Mtpa million tonnes per annum TSX Toronto Stock Exchange ft. foot mV milli volt U Uranium
g gram N (Y) northing U3O8 tri uranium octoxide G&A General and Administration N$ Namibian Dollars UCS Unconfined Compressive Strength g/L grammes per litre NaOH sodium hydroxide UF6 Uranium Hexafluoride g/m³ grams per cubic metre Nb Niobium UNDP United Nations Development Program g/t grams per tonne of gold Ni nickel UO4 Uranium tetroxide GDP Gross Domestic Product NNNP Namib-Naukluft National Park UTM Universal Transverse Mercator GPR Ground Penetrating Radar NPV net present value v/v volume concentration GRS Gamma Ray Spectrometer NQ size of diamond drill rod/bit/core VBF Vacuum Belt Filters
GWe Gigawatt electrical NQ 2 size of diamond drill rod/bit/core w/w mass fraction H2O2 hydrogen peroxide ºC degrees centigrade w:o waste to ore ratio H2SO 4 sulphuric acid OEM Original Equipment Manufacturer WAGE West African Gold Exploration HARD half the absolute relative difference OGR Old German Railway WGS84_33s World Geodetic System 1984 - zone 33 south HDPE high density poly ethylene OK Ordinary Kriging wrt with respect to HPGR High Pressure Grinding rolls ORP Oxidation Reduction Potential XRF X-Ray fluorescence
HQ size of diamond drill rod/bit/core P80 80% passing
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3 RELIANCE ON OTHER EXPERTS
The authors of this report are not qualified to comment on each of the issues detailed below and accordingly have relied upon the representations and judgements of the following parties for each of these areas of the report:
Husab Project
Document 1– EPL 3138 Transfer approval letter: Report identity: Republic of Namibia Ministry of Mines & Energy letter RE: Application for the Transfer of EPL 3138 from West Africa Gold Exploration (Pty) Ltd to TLP Investments Seventy One (Pty) Ltd, 09 February 2007.
Maker of report: Republic of Namibia Ministry of Mines & Energy, A. Iilende Acting Mining Commissioner.
Reliance: Signed letter from Ministry of Mines & Energy, scanned copy of original. Document 2 – EPL 3138 Renewal of Exclusive Prospecting Licence letter: Report identity: Republic of Namibia Ministry of Mines and Energy letter RE: Notice to applicant of preparedness to grant application for the renewal of exclusive prospecting licence 3138 to TLP Investments Seventy One (Pty) Ltd. Approval granted to renew License 3138, 1 June 2007. Includes stamped endorsement.
Maker of report: Republic of Namibia Ministry of Mines and Energy, A. Iilende Acting Mining Commissioner.
Reliance: Signed letter from Ministry of Mines and Energy, scanned copy of original. Document 3 – EPL 3138 Licence Grant/Renewal: Report identity: Republic of Namibia Ministry of Mines & Energy EPL 3138 Licence, Dated 5 April 2004, License Stamp (dated 07 October 2004), Transfer Stamp (dated 21 February 2007), and Renewal Stamp (dated 20 April 2011 Extending the liscence until 19 April 2013.
Maker of report: Republic of Namibia Ministry of Mines & Energy, Minister of Mines and Energy.
Reliance: Signed letter from Ministry of Mines & Energy, scanned copy of original stamped document. Document 4 – EP L3149. Notice to Grant Exclusive Prospecting Licence letter: Report identity: Republic of Namibia Ministry of Mines and Energy letter RE: Notice to applicant of preparedness to grant application for an exclusive prospecting licence 3439 to Extract Resources Namibia (Pty) Ltd. Dated 30 October 2006. Includes stamped endorsements.
Maker of report: Republic of Namibia Ministry of Mines and Energy, A. Iilende Acting Mining Commissioner.
Reliance: Signed letter from Ministry of Mines and Energy, scanned copy of original.
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Document 5 – EPL 3149 Transfer Approval letter: Report identity: Republic of Namibia Ministry of Mines & Energy letter RE: Application for the Transfer of EPL 3439 from Extract Resources (Namibia) (Pty) Ltd to TLP Investments Seventy One (Pty) Ltd, 20 February 2007. Includes stamped endorsement.
Maker of report: Republic of Namibia Ministry of Mines & Energy, A. Iilende Acting Mining Commissioner.
Reliance: Signed letter from Ministry of Mines & Energy, scanned copy of original. Document 6 – EPL 3439 Licence Grant/Renewal: Report identity: Republic of Namibia Ministry of Mines & Energy EPL 3439 Licence Grant (dated 30 October 2006), Transfer Stamp (dated 21 February 2007) and Renewal Endorsement stamp, (30 April 2008) and Renewal Stamp (dated 30 April 2007).
Maker of report: Republic of Namibia Ministry of Mines & Energy, Commissioner.
Reliance: Commissioner signed stamp from Ministry of Mines & Energy, scanned copy of original.
Document 7 – EPL 3138 Environmental Clearance Letter: Report identity: Republic of Namibia, Ministry of Environment and Tourism letter RE: Environmental Clearance for the Environmental Assessment and Management Plan for the West Africa Gold Exploration Project – License 3138. Dated 29 April 2005. Includes stamped endorsement.
Maker of report: Republic of Namibia Ministry of Environment and Tourism, Permanent Secretary.
Reliance: Permanent Secretary signed stamp from Ministry of Environment and Tourism, scanned copy of original.
Document 8 – EPL 3439 Environmental clearance Letter Report identity: Republic of Namibia, Ministry of Environment and Tourism letter RE: Environmental Clearance for the Environmental Impact Assessment of the Proposed Diamond Drilling for West Africa Gold Corporation Ltd, within the Namib Naukluft Park (EPL 3439). Dated 26 February 2007. Includes stamped endorsement.
Maker of report: Republic of Namibia Ministry of Environment and Tourism, Permanent Secretary.
Reliance: Permanent Secretary signed stamp from Ministry of Environment and Tourism, scanned copy of original.
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Document 9 – Announcement of Agreed restructure Report identity: ASX Release. Agreed Restructure – Extract Resources Limited and Kalahari Minerals plc. Dated 5 September 2008.
Maker of report: Rance Dorrington. Extract Company Secretary.
Reliance: Downloaded copy of original. Document 10 – Legal Opinion on the status of EPL 3138.
Letter from Theunissen, Louw and Vennote Partners
Attorneys, Conveyancers, Notaries
Emailed letter confirming the ownership and status of EPL 3138
These files all have legal, title, tenure, land acquisition and compensation, and permitting implications.
Portions of this report (specifically Chapters 21, 22 and 24), were prepared under the supervision of technical experts who may not currently qualify as Qualified Persons for the purposes of the NI43-101; although it is likely that this would not be the case under the revised Code due to take effect in July. The authors are not qualified to comment on these areas and have included these chapters using the ‘Reliance of other Experts’ provision of Item 5 of the Ni-43101 Form F1 (Rules and Policies). The source-authors for these Sections are outline in Section 2.2. The Specific documents referred to include:
Stobart, B. 2011 – Environment and Permitting. Document detailing the Environmental Studies, Permitting and social or community impact requirements for a mining of the Husab Project.
Davies, S. 2011 – Market Studies and Contracts. Document detailing the marketing and taxation for a mining of the Husab Project.
Bevan, J. 2011 – Economic Analysis. Document detailing the economic analysis for a mining of the Husab Project.
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4 PROPERTY DESCRIPTION AND LOCATION
4.1 Background Information on Namibia
4.1.1 Demographics and Geographic Setting
Namibia is an independent republic with a total surface area of 825,418km², situated north of South Africa, west of Botswana and south of Angola. It is bordered to the west by the Atlantic Ocean (Figure 4.1.1_1). Namibia forms part of the Southern African Region.
Figure 4.1.1_1 Geography of Namibia
Windhoek, with a population of 230,000, is the capital of Namibia and is located in the Khomas Region in the centre of the country. The largest harbour is located at Walvis Bay, on the central west coast, south of Swakopmund. The country is mostly arid or semi-arid, comprising a high inland plateau bordered by the Namib Desert along the coast and the Kalahari Desert to the east. The Namibian coastline is swept by the cold Benguela current.
Namibia has a population of approximately 2 million people. The population comprises approximately 87.5% indigenous people, 6% of European descent and 6.5% of mixed origin. About 50% of the population belong to the Ovambo tribe and 9% to the Kavango tribe. Other ethnic groups include the Herero (7%), Damara (7%), Nama (5%), Caprivian (4%), Bushmen (3%), Baster (2%) and Tswana (0.5%).
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The official language is English, however Afrikaans is the commonly spoken language of most of the population. Indigenous languages include Oshivambo, Herero and Nama. According to World Bank standards, 84% of the population are literate.
4.1.2 History and Political Status
South Africa occupied the German colony of South West Africa during World War I and administered it as a mandate until after World War II when it annexed the territory. In 1966, the South West Africa People's Organization (SWAPO) launched a war of independence for the area that was subsequently named Namibia, but it was not until 1988 that South Africa agreed to end its administration in accordance with a UN peace plan for the entire region.
Namibia gained independence from the South African mandate on 21 March 1990 following multi-party elections and the establishment of a constitution. President Sam Nujoma served for the first three terms and was succeeded by President Hifikepunye Pohamba in March 2005 following a peaceful election.
4.1.3 Infrastructure
Namibia is serviced by a network of sealed highways connecting Windhoek in the central plateau region of Namibia with the coast at Walvis Bay, and with Botswana, Angola and South Africa. Generally unsealed but well-maintained access roads provide regional access throughout Namibia. Power is available via local extensions to an extensive regional electricity grid originating in South Africa. A railway line extends from the port of Walvis Bay to Tsumeb, where a copper smelter is in operation.
Water for the Husab Uranium Project is expected to be sourced from a yet to be built coastal desalination plant.
Areas within the Namib-Naukluft National Park, which includes the Husab Project, are granted for exploration, subject to appropriate environmental commitments.
4.1.4 Industry
The economy is heavily dependent on the extraction and processing of minerals for export. Mining accounts for approximately 20% of GDP. Namibia also has an important traditional subsistence agricultural sector.
In 2010, the estimated GDP (purchasing power parity) was US$14.64 billion and the per capita income was US$6,900 with a real growth rate of 4.1%. Mining of zinc, copper, and silver and increased fish production led growth in 2003-05. However, more recently, poor fish catches, a dramatic fall in diamond prices and higher costs of producing metals has undercut growth.
A high per capita GDP relative to the region hides the great inequality of income distribution. The UNDP’s 2005 Human Development Report indicated that 55.8% of the Namibian population live on $2 per day.
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Namibia is estimated to have earned US$1.28 billion from the export of diamonds, copper, gold, zinc, lead, uranium, cattle, and processed fish and karakul skins during 2010.
The Namibian economy is closely linked to South Africa, with the Namibian dollar pegged to the South African Rand.
Rich alluvial diamond deposits make Namibia a primary source for gem-quality diamonds. Namibia is the fourth-largest exporter of non-fuel minerals in Africa, the world's fourth-largest producer of uranium, and a producer of copper, lead, zinc, tin, silver, gold and tungsten. The mining sector employs only about 3% of the workforce, while about half of the population depends on subsistence agriculture for its livelihood.
Namibia normally imports about 50% of its cereal requirements. In drought years, food shortages are a major problem in rural areas.
Privatisation of several enterprises has been undertaken to stimulate long-term foreign investment.
4.1.5 Mining
Artisanal workers exploited copper mineralisation within Namibia prior to re-discovery during the 1800s following the exploration and development of the Southern African region. In Namibia (formerly German South West Africa), this exploration was completed largely by German interests. Modern mining began during 1840 when the Matchless Mine, southwest of Windhoek, was developed by the Walwich Bay Copper Mining Company.
Major operating metalliferous mines are present at Rössing (uranium), Skorpion (zinc), Navachab (gold). The Kombat Mine (copper-lead-zinc) closed in 2008.
4.2 Mineral Tenure
In Namibia, all mineral rights are vested in the State. The Minerals (Prospecting and Mining) Act of 1992 regulates the mining industry in the country. The Act has been designed to facilitate and encourage the private sector to evaluate and develop mineral resources. The Mining Rights and Mineral Resources Division in the Directorate of Mining is usually the first contact for investors, as it handles all applications for and allocation of mineral rights in Namibia.
Several types of mining and prospecting licenses exist, as outlined briefly below.
4.2.1 Non-Exclusive Prospecting Licenses (NEPL)
Valid for 12 months, these licenses permit prospecting non-exclusively in any open ground not restricted by other mineral rights. Prospectors must furnish the Mining Commissioner with details on all samples removed from the NEPL area.
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4.2.2 Reconnaissance Licenses (RL)
These licenses allow regional remote sensing techniques, and are valid for 6 months (renewable under special circumstances) and can be made exclusive in some instances. A geological evaluation and work plan needs to be submitted to the Mining Commissioner.
4.2.3 Exclusive Prospecting License (EPL)
Individual EPLs can cover areas not exceeding 1,000km² and are valid for three years, with two renewals of two years each. Two or more EPLs can be issued for more than one mineral in the same area. A geological evaluation and work plan (including estimated expenditure commitments) is a pre-requisite prior to issuing of the licenses.
4.2.4 Mineral Deposit Retention Licenses (MDRL)
These allow successful prospectors to retain rights to mineral deposits that are uneconomic to exploit immediately. MDRLs are valid up to five years and can be renewed subject to limited work and expenditure obligations.
4.2.5 Mining Licenses
Mining Licenses (ML) can be awarded to Namibian citizens and companies registered in Namibia. They are valid for the life of mine, or an initial 25 years, renewable for successive periods of up to 15 years. Applicants must have the financial and technical resources to mine effectively and safely.
With the exception of NEPLs and RLs, prior to licenses being issued, all applicants are required to complete an environmental contract with the Department of Environment and Tourism. Where relevant, environmental impact assessments must be made with respect to air pollution, dust generation, water supply, drainage/waste water disposal, land disturbance and protection of fauna and flora. An overview of Namibian mineral exploration and mining rights is provided in Table 4.2.5_1.
4.3 Project Locations and Land Area
The Husab Uranium Project Zones 1 to 4 is located within EPL 3138 which is itself situated within the broader Husab Project area. The Husab Project area consist of EPL 3138 and EPL 3439, which are located in central west Namibia and cover an area of approximately 637km² within the Namib Desert (Figure 4.3_1).
The project has dimensions of approximately 20km east-west by 35km north-south, and is bounded by the Swakop River in the south and the Khan River to the west. The Husab project area is characterised by low ridges of rock and wide expanses of sand and colluvial cover, interspersed with calcrete pans and traversed by deeply incised river valleys.
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Table 4.2.5_1 Namibian Mineral Exploration and Mining Rights
Namibia Mineral Exploration and Mining Rights Mining Act : Minerals (Prospecting and Mining) Act, 1992 State Ownership of Minerals : Yes Negotiated Agreement : No Mining Title / License Types Reconnaissance Tenements : Exclusive Reconnaissance License Exploration Tenements : Exclusive Prospecting License Non-exclusive Prospecting License Mining Tenements : Mining License Retention Tenements : Mineral Deposit Retention License Special Purpose Tenements : - Small Scale Mining Tenements : Mining Claim Reconnaissance Tenement Name : Exclusive Reconnaissance License Purpose : Permits regional prospecting by aerial or surface ground means or other remote sensing techniques Maximum Area : Two 1" x 1" graticules Duration : 6 months Renewals : 1 x 6 months (in special cases) Area Reduction : No Procedure : Application to Ministry of Mines and Energy Granted by : Ministry of Mines and Energy Prospecting Tenement Name : Exclusive Prospecting License Purpose : Confers exclusive prospecting rights for specified mineral(s) Maximum Area : 1000km² Duration : 3 years Renewals : 2 x 2 years Area Reduction : Yes. Relinquish at least 25% on first renewal and at least an additional 50% on second renewal. Can apply to Ministry of Mines and Energy for relaxation on these requirements. Procedure : Application to Ministry of Mines and Energy Granted by : Ministry of Mines and Energy Mining Tenement Name : Mining License Purpose : Confers the exclusive rights to extract specific minerals Maximum Area : Depending on deposit size Duration : 25 years or the life of mine, whichever is shorter Renewals : Multiple 15 year extensions to the end of mine life Procedure : Application to Ministry of Mines and Energy Granted by : Ministry of Mines and Energy
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Figure 4.3_1 Project Location Map
The projects are wholly-owned by Swakop Uranium (Pty) Ltd, a wholly-owned subsidiary of Extract Resources Limited. Table 4.3_1 shows the tenement schedule for EPL 3138 and EPL 3439, both of which are renewable on a bi-annual basis.
All the location references are based on the UTM WGS84 Zone 33 South map projection. In the Ida Dome area, a virtual grid has been established for the purposes of drillhole planning, and the bearing correction is 30 degrees (local grid north = 030 degrees from true north). The origin for the grid transformation is shown below:
UTM 503,110E 7,486,520N Local 20,000E 50,000N
The property boundaries were surveyed by a Differential Geographical Positioning System (DGPS).
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Table 4.3_1 Husab Uranium Project Tenement Schedule
Minimum Tenement Tenement Application Renewed Renewed Area Annual Rent Holder Expenditure Type No. Date From To km² (N$) (N$) EPL 3138 01.04.2011 20.04.2011 19.04.2013 Swakop Uranium (Pty) Ltd ^ 416 8,000 540,000 EPL 3439 03.11.2009 03.11.2009 02.11.2011 Swakop Uranium (Pty) Ltd ^ 221 4,000 1,500,000 ^ A wholly owned subsidiary of Extract Resources Pty Ltd
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4.4 Agreements and Encumbrances
A third party, Meercat Minerals Pty Ltd, retains a royalty of 1.75% of gross production revenue from base metal and precious metals revenue over EPL 3138 only.
4.4.1 Environmental Liabilities and Permits
The Husab Project area falls entirely within the Namib-Naukluft National Park. The environmental aspects of exploration are governed by the Ministry of Environment and Tourism. Coffey Mining understands that the ministry allows exploration and mining (as demonstrated by production commencing at the Langer Heinrich Mine in late 2006) within such areas, and is unaware of any current regulations that may significantly restrict access to this area for exploration.
Swakop Uranium lodged an Environmental Impact Assessment (EIA) and Environmental Management Plan (EMP) for a future mining operation at Husab with the Ministry of Environment and Tourism in November, 2010. Environmental approval for the mine and plant was received from Namibia’s Ministry of Environment and Tourism in January 2011. A further application will be made in respect of linear infrastructure required to support the plant. A Mining Licence application was lodged with the Ministry of Mines and Energy in December, 2010. The Mining Licence application area covers the main extent of granite hosted uranium mineralisation, at the northern end of Exclusive Prospecting Licence 3138.
A thorough overview of environmental liabilities and permits is shown in Section 21, Environmental Studies, Permitting and Social or Community Impact.
Permitting required to carry out the current exploration and feasibility study activities at the Husab Project has been sought and granted by the relevant Namibian Government departments:
Site Borehole water abstraction permits.
Exclusive Prospecting Licenses (EPL3138 & 3439).
Exploration Environmental Management Plan.
Accessory Works permits to operate camps in the NNNP.
Park permits for all employees and contract staff employed on the Husab Project.
Permits to allow for blasting and removal of a bulk test sample.
Exploration Radiation Management Plan.
Additional permits will be required prior to, and during, development of the proposed Husab Mine. Some have been applied for or granted, however at this stage, most have not. These permits include:
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Mining License (applied for 13 th Dec 2010).
Mining License Environmental Impact Assessment and Management Plan (approved 25 th Jan 2011).
Power Agreements (under negotiation).
Site water use permits.
Continuous Operation permit.
Application to exceed overtime.
Storage of explosives permit.
Transport of explosives permit.
Blasting vehicle license.
Necessary accessory works.
Construction period EMS.
Construction period EMP.
Access road junction with main road.
Telecoms Namibia service.
Radio frequency license.
Supply route of explosives from South Africa to Husab.
Magazine forklift.
Construction and Mining Radiation Management Plans.
Plant relocation/removal permits.
Archaeological site destruction permits.
Linear Infrastructure EIA and EMP (roads/pipes/powerlines).
Power usage.
Construction and operations – expatriate work visas.
Negotiate surface use with adjacent EPL holders.
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5 ACCESSIBILITY, CLIMATE, PHYSIOGRAPHY, LOCAL RESOURCES AND INFRASTRUCTURE
5.1 Access
The main access to the project area is gained by travelling approximately 40km east from Swakopmund or Walvis Bay along unsealed regional access roads and then 10km north along the Welwitschia Flats road. A second method of access is via an unsealed track from the northeast that joins with the main Windhoek-Swakopmund highway. Access within the project area is via unsealed roads established by previous explorers.
5.2 Climate
The project area is characterised by an arid temperate desert climatic regime, characterised by low sporadic rainfall averaging around 10mm per year, although high rainfall events (over 20mm) occur periodically. The project areas lie within the outer range of the Atlantic coastal fog zone. Temperatures average over 25 degrees Celsius in the summer months, and can peak at over 40 degrees. Winter temperatures are somewhat milder, with cool mornings and daily maxima around 20 degrees. Winds are generally from the south and west, but occasional strong north-easterly winds from the interior bring hot dusty conditions.
The climate has little or no effect on the length of the operating season, and exploration activities can be carried out all year.
5.2.1 Topography, Elevation and Vegetation
The area of the Husab Uranium Project is typified by flat desert plains with sparse bushy vegetation (Figure 5.2.1_1). The plains are characterised by surface sand and grit with overburden cover of up to up to 60m. To the west of the Project, the desert sand cover dies out abruptly along a north-east trend to expose rugged gully terrane.
More broadly, the Husab Mountains in the east of EPL 3138 are typical of the rugged chain of hills representing Swakop Group lithologies extending from Witpoortberg in the south, through Husaberg, to Marmor Pforte and Chuosberg to the north. These ranges of steep hills define the eastern edge of an extensive inter-montane plain lying between the Khan River and the mountains. The plain shows little relief apart from occasional hills, and is poorly vegetated in the Namib Park, although reasonably well-grassed over farmland to the north. The plain shows a typical desert (Figure 5.2.1_1) deflation surface composed of calcretised surface sand and grits with minor sub-crop.
Within a broad swathe on both sides of the Khan and Swakop rivers, water erosion has formed a typical badlands environment showing elements of extreme topography cut by deep river valleys.
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Figure 5.2.1_1 Desert Plains in the Husab Uranium Project Area
Vegetation is sparse, comprising generally scattered low shrubs and herbage on the plains, and more abundant plant growth within the seasonal drainage channels. Trees, such as the Camel Thorn, are restricted to the larger drainages. Exotic trees and shrubs, including some invasive species, are present in the Swakop Gorge, where they have been introduced by previous inhabitants or derived from seed dispersal down the river. Overall the topography varies in height between 200m and 710m.
5.2.2 Local Resources and Infrastructure
The project areas lie adjacent to well maintained gravel roads that service the National Park, and are close to the main Swakopmund -Windhoek highway. High-voltage power supply lines cross the tenement areas. Land in the area is dominantly used for game farming and national parks.
Water for industrial and pastoral activities is provided via subterranean resources, major watercourses and numerous su b-artesian bores. Water for drilling purposes is currently drawn from water bores.
Swakopmund is the closest major centre, comprising excellent support services such as transport, earthmoving, construction, commercial and banking. Namibia’s main port cit y, Walvis Bay, is located 30km south of Swakopmund.
Non technical exploration personnel are currently sourced out of Windhoek and Swakopmund, with management personnel currently sourced mainly from Australia. Specialist contractors also originate from Sou th Africa and Australia, whereas junior technical staff are mainly Windhoek based.
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5.3 Future Mining Operations
The electrical power utility provider in Namibia, NamPower, commands sufficient generated and/or imported energy sources to provide both the temporary and permanent power needs for the Husab mine. Capital, to be supplied by Swakop Uranium, is however required to fund the extensions to the transmission network that runs parallel to the B2 highway between Windhoek and Swakopmund. The solutions for both temporary and permanent power supply have been developed around this arterial supply route and fits well with the timeline of the Husab mine’s development and production schedule. Discussions with NamPower have taken place and are ongoing.
Existing fresh water resources in the area do not have adequate capacity to supply the projected regional demand and desalination of seawater is considered to be the only viable solution for permanent water supply. During development, water will be supplied through a temporary water supply pipeline to be constructed from the NamWater reservoir near the Rössing Mine. During operation, water is expected to be sourced from the proposed desalination plant at Mile 6, to be constructed and operated by third parties.
The company is a member of the Erongo Mining Water Users Group (EMWUG) which is working with the National Desalination Task Force (NDTF) to investigate and implement a strategy to deliver water to the project in line with the envisaged development timetable. After recommendations were tabled to the Namibian Cabinet in February 2011, a Public-Private- Partnership (PPP) has been approved for the structuring of a new desalination plant. However, to date there is no commitment from any party to build a desalination plant, and the company continues to assess potential temporary or fall back solutions in line with the project’s development timetable.
It is envisaged that the majority of workers employed at the Husab Mine will come from, and be resident in, the nearby towns of Arandis (pop. 4,500), Swakopmund (pop. 25,500) and Walvis Bay (pop. 60,000).
The proposed site layout plan, situated entirely within EPL3138, for the Husab Mine (Figure 1.6_1) shows the open pit mines together with the process plant and combined tailings/waste rock landform.
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6 HISTORY
6.1 Ownership History
The ownership history of the Husab Project (EPLs 3138 and 3439) is discussed in Section 6.2.
6.2 Exploration History
Prospecting for uranium mineralisation within the area captured by EPL 3138 took place following the discovery of large resources of low grade granite (alaskite) hosted mineralisation at Rössing. Elsewhere, further discoveries were made of primary uranium mineralisation at Valencia, Goanikontes and the Ida Dome. Only Rössing was developed as a mine, and the other occurrences became victims of the severe downturn in the uranium market in the early 1980s. This downturn lasted until around 2004 when the world uranium market began to revive.
Exploration activities by Kalahari Gold and Copper Pty Ltd during the period 1996-2002, included the collation and assessment of previous exploration data, and the acquisition and interpretation of detailed aeromagnetic and satellite imagery. Considerable multi-element partial leach soil geochemistry was carried out over the Husab area, and also the Von Stryk and Von Stryk South magnetite-barite skarns (to the south of EPL 3138), although no significant anomalous areas were generated from this work.
Exploration on the current EPLs commenced with the granting of EPL3138 to West Africa Gold Exploration (Pty) Ltd (WAGE) – a subsidiary of Kalahari Minerals plc – in April 2004. EPL3439 was granted to Extract’s wholly-owned Namibian subsidiary Swakop Uranium (Pty) Ltd.(SUPL) in November 2006. The Husab joint venture on EPL3138 between WAGE and Extract was signed in May 2005 and approved by the Mines Minister in October 2005. By December 2006 the Husab joint venture was terminated and EPL3138 transferred to SUPL and Kalahari Minerals plc became a major shareholder of Extract.
In 2006, Extract undertook a geological and structural review of the Husab project area and formulated a targeting plan for deposits located under surface cover. Extract used magnetic data to target the contact positions of the Khan Formation with follow up drilling through the cover sequence to sample the underlying bedrock. In 2007, a traverse of three angled RC drillholes targeted mineralisation beneath a low-level RAB anomaly. The RC drilling intersected significant uraniferous pegmatitic alaskites, with follow up drilling defining the Husab Uranium Project Zone 1 and Zone 2 prospects (Spivey and Penkethman, 2009).
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6.3 Resource History
The previous resources for Husab Uranium Project Zones 1 and 2 are discussed in Inwood (2009). Inwood (2008) discusses the estimated resources for the Garnet Valley, New Camp and Ida Central prospects which are located within the broader Husab Project Area. These estimates were carried out by Coffey Mining in August 2008 and are summarised below in Table 6.3_1.
Table 6.3_1 Husab Uranium Project, Namibia August 2008 Resource Estimate Reported at Various Cutoffs using a Bulk Density of 2.65t/m³ Ordinary Kriged Estimate based upon 3m (Garnet Valley and Ida Central) or 2m (New Camp) cut U 3O8 Composites
Tonnes Above Cutoff U O Contained U O Lower Cut 3 8 3 8 (Mt) (ppm) (M lb) Garnet Valley (Blocks 40m NS by 20m EW by 10m RL) 100 0.6 246 0.31 Indicated 200 0.5 259 0.26 100 43.5 224 21.39 Inferred 200 25.6 263 14.77 New Camp (Blocks 40m by 20m by 10m RL - Block Model Rotated about 0600) 100 4.0 156 1.4 Inferred 200 0.4 234 0.2 Ida Central (Blocks 40m N by 10m E by 20m RL) 100 5.2 170 1.96 Inferred 200 1.1 238 0.6 All Deposits 100 0.6 246 0.31 Indicated 200 0.5 259 0.26 100 52.7 213 24.8 Inferred 200 27 261 15.6
Inwood (2009) discusses the estimated maiden resource for Husab Uranium Project Zone 2 and an updated resource estimate for Zone 1. The Zone 1 updated resource estimate was undertaken by Extract in July 2009 and the Zone 2 maiden resource estimate was undertaken by Coffey Mining in August 2009. These resources are summarised in Table 6.3_2.
.
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Table 6.3_2 Husab Uranium Project, Namibia Zone 1 and Zone 2 Resource Estimates - Superseded August 2010 Reported at Various Cutoffs using a Bulk Density based on lithology Ordinary Kriged Estimates based upon 3m cut U3O8 Composites
Tonnes Above Cutoff U O Contained U O Lower Cut 3 8 3 8 (Mt) (ppm) (M lb) Zone 1 100 20.7 527 24.0 200 19.7 546 23.7 Inferred 300 16.2 608 21.7 400 12.5 684 18.8 500 8.8 782 15.2 Zone 2 100 102 543 122 200 96 565 120 Inferred 300 82 620 112 400 68 676 101 500 51 749 85
6.3.1 Production History
No historic uranium production has been recorded on any of the current project tenements.
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7 GEOLOGICAL SETTING
7.1 Regional and Local Setting
The regional geological setting of Namibia is dominated by the Damara Orogenic Belt, a major northeast trending belt of Mesoproterozoic to earliest Palaeozoic rocks that formed within a major intracontinental rift basin (Figure 7.1_1). The rift was deformed by closure of the basin during the late Neoproterozoic and early Palaeozoic. The Husab Project is located in the central (Swakop) zone of the Damara Orogen. Mineralisation is associated with structural and intrusion-associated settings formed during the major thrust deformation that closed the orogen.
Figure 7.1_1 Damara Orogenic Belt : Regional Geological Setting
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The Central Zone is generally thought of as a palaeo-arch, where Damara and pre-Damara rocks are complexly intermixed in a dome and basin topology following the collision of the Congo and Kalahari Cratons. The Central Zone is bounded by two north-east trending structures known as the Omaruru Lineament on the northern margin, and the Okahandja Lineament to the south (Spivey and Penkethman, 2009).
The regional setting features a northeast trending anticlinal hinge zone, with extensive inliers of Palaeoproterozoic rocks of the Abbabis Metamorphic Complex commonly present as elongated dome-like features surrounded by younger folded Damara Sequence lithologies.
The Abbabis Metamorphic Complex lithologies typically underwent medium to high grade metamorphism, typically amphibolite to granulite facies, with the succession dominantly comprising metasediment, paragneiss, orthogneiss and ortho-amphibolite, all of which are extensively invaded by pegmatites.
The Damara Orogenic Belt is a major Pan African age mobile belt of sedimentary and volcanic rocks that extends from the Namibian coast northeast into Botswana and Zambia. The orogen was initiated via rifting between the Kaapvaal and Congo Cratons between 1,000Ma and 900Ma, and closed at approximately 500Ma during a period of major thrust faulting and granitoid intrusion. Structures developed during initial rifting have continued to influence the tectonic development of the Namibian region up until the Jurassic period, when the African and South American continents rifted apart to produce the South Atlantic Ocean.
The DOB is broadly divided into three successions (Figure 7.1_2), as follows:
The Nosib Group comprising metamorphosed sandstone, quartzite and minor conglomerate, and representing sediments deposited due to erosion from uplifted basement during the initial rifting of the orogen. The Nosib is further sub-divided into Etusis Formation quartzites, arenites and arkoses, and the Khan Formation , consisting mainly of calc-silicate rocks with lesser clastic components.
The lower Swakop Group , which overlies the Nosib, comprises an alternating succession of dolomite, marble, schist and schistose diamictite. The succession includes, from lower to upper stratigraphic levels, the Rössing Formation (mainly carbonates, wackes, quartzites and mica-schists), the Chuos Formation (mica schists, calc-silicates and carbonates), and the Arandis and Karibib Formations (mica-schists, calc-silicate rocks and marbles). The lower Swakop Group is interpreted to include the thermal sag phase of rifting, which coincides with regional sag of the crust due to cooling of the underlying mantle.
The Kuiseb Formation , which comprises the upper part of the Swakop Group, is developed predominantly within the southern portion of the Damara Orogen. The Kuiseb Formation (mainly Flysch-type sediments) is represented by a thick succession of metamorphosed sedimentary rocks, generally comprising biotite schist with subordinate calcsilicate rocks and carbonaceous schist. A unit of metamorphosed mafic volcanic rocks, referred to as the Matchless Amphibolite Member, shows compositions consistent with those observed in mid-ocean ridge basalts.
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Figure 7.1_2 Stratigraphic Column of the Damara Orogen Group Subgroup Formation Max Thickness Lithology Pelitic and semi-pelitic schist and gneiss, migmatite, calc-silicate rock, quartzite. Kuiseb >3000 Thinkas member: Pelitic and semi-pelitic schist, calc-silicate rock, marble, para-amphibolite. Khomas Marble, calc-silicate rock, pelitic and semi-pelitic Karibib 1000 schist and gneiss, biotite amphibolite schist, quartz schist, migmatite. Swakop Diamictite, calc-silicate rock, pebbly schist, Chuos 700 quartzite, ferruginous quartzite, migmatite. Discordance Marble, pelitic schist and gneiss, biotite- Ugab Rössing 200 hornblende schist, migmatite, calc-silicate rock, quartzite, metaconglomerate. Discordance Migmatite, banded and mottled quartzo- feldspathic clinopyroxene-amphibolite gneiss, Khan 1100 hornblende-biotite schist, biotite schist and gneiss, migmatite, pyroxene-garnet gneiss, amphibolite, Nosib quartzite, metaconglomerate. Quartzite, metaconglomerate, pelitic and semi- pelitic schist and gneiss, migmatite, quartzo- Etusis 3000 feldspathic clinopyroxene-amphibolite gneiss, calc-silicate rock, metaphyolite. Major Unconformity Gneissic granite, augen gneiss, quartzo-feldspathic Abbabis Complex gneiss, pelitic schist and gneiss, migmatite, quartzite, marble, calc-silicate rock, amphibolite. Note: taken from Schneider and Seeger (1993) and modified after Jacobs et Al, 1986
All the above rock types have been intruded by a wide variety of syn-tectonic and post- tectonic intrusive rocks, including the Goas Diorite Suite, which includes a series of small disconnected plutons in the area south of Karibib. Other late to post-tectonic intrusives include the Salem type granites, commonly represented by grey biotite granites to granodiorites locally intruding Damara rocks. The so-called “red” granites, which are a heterogeneous group of intrusives composed of foliated and massive varieties, often occurring as small bodies intruding the Swakop Group marbles and schists. Other, mainly post-tectonic intrusive types include the leucogranites and alaskites, some of which are uraniferous. They typically appear below and around the lowest Damara marbles, and are characterised by vein to dyke-like or anastomosing forms, although massive or plug-like occurrences are known. A body of grey granite at 7492300mN 505000mE has itself been intruded by a uranium bearing alaskite, and the Rössing Mine exploits low grade uranium mineralisation within one such intrusive system.
The Damara Orogen was closed by a major south-directed thrust deformation, resulting in a pervasive north-northwest to northwest dipping foliation. Thrusts are interpreted to have transported portions of the sequence up to 200km to the south over basement rocks.
Rocks of the DOB are covered by the Kalahari Sands in northeast Namibia, but are continuous through Botswana into Zambia, where the orogen is represented by the Muva Group and Katanga Group of the Lufilian Arc.
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7.2 Project Geology
7.2.1 Husab Project
The Husab Project is dominated by a series of north-northeast to northeast trending regional- scale antiforms and synforms, which make up the main structural architecture of the entire Central Zone of the Damara. These meta-sedimentary folds or dome-like structures of the DOB are cored by gneissic and metasedimentary rocks of the Abbabis Formation (Figure 7.2.1_1). Figure 7.2.1_2 shows a geological cross section along northing 7506000 – the Husab Uranium Project discovery line. The basement rocks are covered to the northeast and south by stranded cover sequences of flat-lying calcrete and alluvial deposits, which are associated with a broad northeast trending valley marginal to the Khan River.
Figure 7.2.1_1 Husab Project : Geological Setting
The basement gneisses outcrop as a series of semi-ovoid features within the Central Zone of the Damara, in general forming somewhat poorly exposed extensions to the basement rocks exposed in the Swakop River Gorge on either side of the Ida Dome, and in the Khan River Valley, immediately south of the Khan Copper Mine, to the junction of the Khan and Swakop Rivers. Flanking Damara Sequence sediments show a complex pattern of folding and faulting, and the whole sequence is extensively invaded by syn- and post-tectonic granitoids and pegmatite swarms. Cross-cutting Mesozoic dolerite dykes are also evident locally.
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Figure 7.2.1_2 Husab Uranium Project Zone 1 Geological Cross Section (7506000N)
Basement domes to the east and west of the Husab project area, the latter along the Khan River, are predominantly comprised of metasedimentary rocks; however, basement associated with the cores of the Ida and Husab mine domes are gneissic. Regional magnetic data indicate that the regional structural history is complex (Figure 7.2.1_3).
Figure 7.2.1_3 Husab Project : Tenement Outline over 1st vertical derivative of magnetics
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8 DEPOSIT TYPES
The Husab Uranium Project uranium mineralisation is associated with intrusive alaskite rocks and is informally referred to as a Rössing Type uranium occurrence.
The best known example of alaskite hosted primary mineralisation and typical of intra- intrusive type mineralisation is the Rössing SJ deposit.
Exploration at the Husab Uranium Project has focused on identification of favourable stratigraphic units, particularly the Rössing and Khan Formations, that may host uranium mineralisation. Initial exploration targeting is based on surface mapping and interpretation of aeromagnetic imagery, which is followed up by relatively shallow, broad-spaced reconnaissance drilling and sampling.
The Rössing Mine is located close to the town of Arandis, 65 kilometres inland from Swakopmund, in the Namib Desert. Rössing is the world’s largest open-pit uranium mine, which started operations in 1976. The deposit comprises an extensive low-grade
(300ppm U3O8) alaskite-hosted deposit. Uranium occurs mostly as uraninite (55%) and beta- uranophane (40%) in the form of interstitial grains and crystal inclusions in minerals; betafite makes up for the remaining 5%. Secondary uranium minerals predominate in the weathering
profile. Average ore grades at Rössing commonly vary between 300ppm and 400ppm U3O8.
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9 MINERALISATION
9.1 Primary Uranium Mineralisation
The EPLs 3138 and 3439 contain occurrences of uraniferous granite s (alaskites) within the highly prospective Central Zone of the Damara Orogeny.
According to Spivey and Penkethman (2009), alaskite mineralisation is associated with “a zone of abnormally high heat flow which has produced leucogranite melts from pre -Damara basement, and preferentially enriched uranium over thorium with respect to source rocks. Magmas appear to be mainly passively emplaced as sheeted bodies and, more infrequently, as stocks. They occur in an area of high metamorphic grade, characterized by high temperature- low pressure metamorphism.”
Mineralisation is thought to be related to dilatational zones which have occurred due to regional deformation, with preferential emplacement of alaskites around the first occurrence of carbonates within the se quence. The change from terrigenous to carbonate sedimentary setting may have acted to change local redox conditions, and/or to trap and focus mineralised fluids beneath marble caps (Spivey and Penkethman (2009) ).
At the Husab Uranium Project , mineralised alaskite occur proximal to the contact between the Khan Formation and calc -silicates of the Rössing Formation. The majority of the uraniferous alaskites are located within the Rössing Formation. The host Rössing Formation can vary from 50m to over 200m wide and mineralised intercepts can vary from a few metres to over 100m wide. The average width of mineralised horizons is approximately 20m. The mineralisation at Zones 1 to 4 is a near-continuous mineralised area extending across approximately 8km stri ke length and up to 2 km across strike. Individual mineralised units can be traced for several hundred metres along strike and down dip. Metallurgical testwork indicates that uraninite is the main uranium mineral. Elevated uranium grades (for example
9,585ppm U3O8) associated with beta -Uranophane, an oxidation product of uraninite, have been identified in drillhole RDD002 (Figure 9.1_1).
Figure 9.1_1 Beta -Uranophane Mineralisation in Hole RDD002 (near 183m)
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The mineralisation is further described in Sections 11 and 17.
9.2 Secondary Uranium Mineralisation
In addition to primary alaskite hosted uranium mineralisation, the tenement holding hosts occurrences of secondary uranium vanadate (carnotite), similar to that comprising mineralisation at the Langer Heinrich development project. The best examples of secondary uranium mineralisation are found at the Husab Uranium Project, north of the Husab fluorite mine. A single line of drilling completed on this target zone has intersected calcrete hosted uranium mineralisation in a flat lying paleochannel.
An occurrence of calcrete uranium mineralisation is also reported in the Welwitschia Flats area exposed in erosion channels, and could be classified as a potential calcrete type deposit.
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10 EXPLORATION
This section will concentrate on the exploration work undertaken by Extract at the Husab Uranium Project area. Inwood (2008) discusses the exploration work performed in the broader Husab Project area in detail. Unless otherwise stated, all the current exploration activities within the project area is carried out by Extract.
Drilling is undertaken by contractors under the supervision of Extract personnel.
10.1 Husab Uranium Project, Zones 1 to 4
Extract announced a new uranium discovery at the Husab Uranium Project in February 2008. Following further resource definition drilling and exploration (which remain ongoing) a study of the feasibility of developing the project was concluded in March 2011. The Husab Uranium Project is located under Namib Desert sand cover with the northern licence boundary shared with the world-class Rössing uranium province.
The Husab Uranium Project target is interpreted as being an extension of the same stratigraphy that hosts the Rössing Mine, located 5km to the north and striking 15km onto the Husab Project. Airborne magnetic data clearly indicates the Rössing stratigraphy folds around and trends into EPL 3138. Southern strike extensions of the same stratigraphy that host the Rössing deposits trend under desert sands, which partly explains why there has been no previous exploration in the Husab Uranium Project area.
The initial drilling at the Husab Uranium Project commenced in April 2007 and was aimed at identifying the Khan and Rössing Formation contact - an unconformity surface that has been preferentially intruded by the leucogranites (alaskites) containing primary uranium mineralisation at the Rössing Mine to the north and at Ida Dome to the south. To test this exploration model, four 1.6km spaced lines, with vertical drillholes on 80m centres, were completed over the north east trending magnetic lows interpreted to represent the extensions of the Rössing stratigraphy.
The exploration model was verified in the field in late 2007 when three holes on line 7,506,000mN, intersected anomalous uranium mineralisation associated with zones of smoky quartz in altered leucogranite, beneath approximately 40m of overburden and leached
saprolite. All three holes recorded uranium values of at least 100ppm U3O8. Follow up angled RC drilling in late 2007 intersected wide zones of alaskite with significant zones of uranium mineralisation.
Chemical assay results in early 2008 confirmed the new alaskite hosted uranium discovery.
Discovery intersections included: 100m at 265ppm U3O8 and 40m at 240ppm U3O8 with both holes ending in mineralisation. Follow up drilling, down dip of the discovery holes, has intersected wide zones of high grade uranium mineralisation with results including 149m at
474ppm U3O8.
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When the drillhole database was handed over to Coffey Mining for the current Husab Uranium Project resource model (August 2010) a total of 117 lines of exploration and resource definition drilling had been completed over 9.4km of the 15km target zone. The majority of these holes are RC with some NQ and HQ core holes. Drillhole line spacing ranges from 1,600 to 25m, with holes on 25 to 100m centres. From this drilling, four zones have been defined (Zones 1 to 4). Zone 1 has been defined over 2,500 metre of strike (7504900 to 7507400mN), Zone 2 over 2,500 metres of strike (7502400 to 7504900mN), Zone 3 over 2,000m of strike (7500400 to 7502400mN) and Zone 4 over 1,200m of strike (7490400 to 750600mN). All Zones are locally open at depth. Zone 1 is open along strike to the north and there is significant potential to discover further mineralisation to the south of Zones 3 and 4.
The August 2010 resource estimate was based on approximately 400,000 metres drilled at the Husab Uranium Project, defining four significant zones of uranium mineralisation, with a cumulative strike length of approximately 9.4 kilometres.
Resource definition drilling over Zone 1 on a nominal 100m by 100m spacing was completed in December 2008 and provided the data used in the maiden Inferred Mineral Resource estimate in January 2009. An updated resource estimate was completed in July 2009 which included, among other additional drilling, some 50m by 50m spaced holes. This closer- spaced drilling enabled some Indicated resources to be defined at Zone 1. Drilling at Zone 1 since July 2009 has concentrated primarily on further infill drilling to increase the level of confidence in the resource estimate. In the August 2010 resource update, over 75% of the Zone 1 resource was classified in the Indicated category.
Resource definition drilling over Zone 2, on a nominal 100m by 100m spacing, was completed in June 2009 and provided data used in the maiden Inferred Mineral Resource estimate in July 2009. After July 2009, additional infill drilling and extensional drilling continued at Zone 2. Drilling is now on 50m sections with 50 to 100m spacing between drillholes. In the August 2010 resource update, over 80% of the Zone 2 resource was classified in the Indicated category.
The August 2010 resource update for the Husab Uranium Project contained maiden estimates for Zones 3 and 4. Drill spacing is on sections 200m apart, with 100m spacing between drillholes.
Details of the Zone 1 to 4 drilling and statistics are discussed in Section 17. Mineralisation remains open along strike to the south and at depth. Figure 10.1_1 displays the location of significant drilling results at the Husab Uranium Project.
Numerous significant zones of uranium mineralisation have been intersected at Zones 1 to 4, with chemical assay results summarised in Tables 10.1_1 and 10.1_2. The true thickness of the uranium mineralised zones ranges from 60% to 100% of the downhole thickness.
The higher grade zones are generally contained within zones of leucogranite with abundant smoky quartz and coarse grained biotite booklets (Figure 10.1_2). Some mineralisation has also been observed in the calc-silicates and schists within and adjacent to the mineralised granites.
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Figure 10.1_1 Husab Uranium Project: Significant Drilling Intercepts
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Table 10.1_1 Husab Uranium Project Zone 1 and Zone 2 Significant Drilling Intersections
Northing Easting Azimuth Dip From To Width Grade Hole ID (WGS84 33S) (WGS84 33S) (deg) (deg) (m) (m) (m) (g/t U 3O8) RDD145 7505150 506350 270 -60 371 444 73 2243 RDD193 7505250 505970 90 -60 253 336 83 968 RDD193 7505250 505970 90 -60 261 328 67 1182 RDD034 7505300 506400 270 -60 352 475 123 737 RDD009 7505500 506050 270 -60 228 294 66 1846 RDD030 7505500 506150 270 -60 84 194 110 736 RRC103 7505700 506000 270 -60 102 207 105 1198 RRC811 7505750 505900 90 -60 75 145 70 1112 RRC595 7505750 506000 270 -60 101 247 146 639 RRC114 7505800 506300 270 -60 66 124 58 1372 RRC114 7505800 506300 270 -60 66 124 58 1372 RRC181 7505800 506400 270 -60 100 258 158 660 RRC750 7505850 505875 90 -60 148 222 74 1042 RRC160 7506600 506650 270 -60 74 175 101 974 RDD037 7506600 506950 270 -60 410 522 112 799 RRC286 7506750 506900 270 -60 193 236 43 2197 RRC042 7506800 506680 270 -60 86 181 95 780 RSRC001 7506800 506690 0 -90 184 272 88 4491 RRC289 7506800 506800 270 -60 234 301 67 2469 RDD218 7507150 507150 270 -60 328 374 46 1700 RRC382 7503000 505700 270 -60 128 203 75 1445 RDD084 7503050 505350 270 -60 128 175 47 2741 RRC374 7503100 505800 270 -60 161 264 103 821 RRC528 7503150 505750 270 -60 136 227 91 854 RRC576 7503250 505750 270 -60 146 211 65 1125 RRC489 7503300 505550 270 -60 135 185 50 1618 RRC297 7503400 505600 270 -60 122 219 97 970 RDD022 7503600 505960 270 -60 112 270 158 552 RDD022 7503600 505960 270 -60 144 226 82 942 RRC610 7503600 506090 270 -60 208 359 151 540 RRC584 7503650 505950 270 -60 107 289 182 402 RRC266 7503700 505500 270 -60 249 270 21 3920 RDD046 7503700 506200 270 -60 297 391 94 896 RDD046 7503700 506200 270 -60 297 351 54 1450 RRC692 7503750 505950 270 -60 179 334 155 760 RDD005 7503800 505900 270 -60 232 305 73 1060 RDD005 7503800 505900 270 -60 247 305 58 1302 RDD102 7503900 505900 270 -65 237 366 129 1415 RRC304 7503900 505900 270 -60 260 342 82 1092 RDD072 7504000 506020 270 -60 351 361 10 9373
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Table 10.1_2 Husab Uranium Project Zone 3 & Zone 4 Significant Drilling Intersections
Northing Easting Azimuth Dip From To Width Grade Hole ID (WGS84 33S) (WGS84 33S) (deg) (deg) (m) (m) (m) (g/t U 3O8) RRC481 7500000 504600 270 -60 169 174 5 1766 RRC482 7500000 504700 270 -60 210 224 14 737 RRC486 7500000 505500 270 -60 265 317 52 207 RRC429 7500800 505100 270 -60 128 140 12 678 RRC430 7500800 505200 270 -60 189 238 49 172 RRC413 7501200 505300 270 -60 207 239 32 323 RRC413 7501200 505300 270 -60 264 292 28 528 RRC414 7501200 505400 270 -60 294 376 82 600 RRC425 7501600 505600 270 -60 284 360 76 145 RRC731 7502200 505450 270 -60 131 157 26 318 RRC774 7499600 503600 0 -90 184 201 17 396 RRC776 7499600 503800 0 -90 88 131 43 468 RRC776 7499600 503800 0 -90 241 250 9 2114 RRC788 7500000 503860 0 -90 167 194 27 733 R3RC0001 7500000 504000 270 -60 157 212 55 1474 RRC789 7500000 504060 0 -90 119 181 62 134 R3RC0002 7500000 504100 270 -60 104 157 53 1616 RRC771 7500200 503800 0 -90 185 195 10 1665 RRC690 7500200 503900 0 -90 160 174 14 642 RRC688 7500200 504000 0 -90 158 181 23 811
Figure 10.1_2 Husab Uranium Project Mineralisation from Hole RDD005
(Note: High grade mineralisation is associated with dark smoky quartz and biotite)
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11 DRILLING
11.1 Introduction
Data collection methods applied by Extract have been reviewed in the field by Coffey Mining and, as such, have been directly assessed. The following sections are a summary of the Extract/WAGE drilling and collection a pproach.
11.2 Diamond Core Drilling
The diamond drilling completed at Husab Uranium Project was undertaken by Major Drilling with a skid-mounted Sulvan and Longyear 44 and 2 track-mounted LF90s. The number of rigs on site at any one time varied between one and ten.
Between January 2008 and July 2009, 209 NQ size diamond drillholes were completed for a total of 84,736 m. The deepest hole to date is 703 m. Drilling was completed in run lengths of up to 3 metres. Figure 11.2_1 show an example of remaining ¼ core from hole RDD002.
Figure 11.2_1 Husab Uranium Project Core – Hole RDD002 Showing the Contact of an Uraniferous Pegmatitic Alaskite with Biotite Schist
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The drilling companies’ performances are generally satisfactory, with acceptable daily productivity rates, acceptable sample recovery and safety standards being achieved.
Care is taken by the site geologists to align the drill-rigs appropriately prior to commencing each hole. The rig alignment (i.e. the azimuth and dip) is also re-confirmed using a compass prior to commencement of drilling.
All holes have been downhole surveyed by a single-shot survey tool (either a Sperry Sun or Eastman instrument) or a Reflex-eze Multishot survey tool. The single shot surveying is currently completed by the drilling contractors at the completion of the hole. Namibian based geophysical contractor, Terratec Geoservices, also completed downhole deviation surveys on the majority of holes drilled with data collected on one centimetre increments. To avoid excessive data collection, Extract choose to use one reading every 5 m which is considered sufficient for accurately plotting a drillhole trace.
Core orientation of all holes with a dip between -45º and -75º has been undertaken since 2007 using either the ACE Reflex orientation tool or a downhole spear. Some holes still remain only partially orientated, due to the rotation of core ends during drilling and discontinuities in the core.
Geotechnical logging has routinely been completed at all target areas, with recoveries and RQDs recorded by a crew of field technicians under the supervision of staff geologists. Where holes could be reliably orientated, alpha and beta measurements have been taken on significant discontinuities such as geological contacts, veins, joints and faults. Twelve holes, six at Zone 1 and six at Zone 2, were drilled in 2009 with the primary aim of collecting geotechnical information. The holes were drilled by Major Drilling, with logging and data collection being supervised by Namibian company GeoLogic Solutions Ltd. Additional work, including external audit of work already completed and interpretation of downhole televiewer data was completed by Golder Associates, operating out of their Johannesburg office.
A review of the core by Coffey Mining showed high core recovery for the holes drilled at the Husab Uranium Project. Core recovery from the drilling database averages approximately 99%.
11.3 RC Drilling
Between April 2007 and June 2010, approximately 1,419 reverse circulation (RC) drillholes were completed at the Husab Uranium Project for 312,570m, with the deepest hole being 415m. The RC drilling has been completed by Major Drilling, Ferro Drill, Metzger Drilling, Wallis Drilling and RC Drilling Services Namibia. At the end of June 2010, RC drilling at the Husab Uranium Project was being carried out by 7 drill rigs.
The drilling companies performances are generally satisfactory, with acceptable daily productivity rates, acceptable sample recovery and safety standards being achieved.
Care is taken by the site geologists to align the drill-rigs appropriately prior to commencing each hole. The rig alignment (i.e. the azimuth and dip) is also re-confirmed using a compass prior to commencement of drilling.
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Figure 11.3_1 RC Drilling at Husab Uranium Project
All holes are downhole surveyed by Namibian based geophysical contractor Terratec Geoservices using the GRS tool with downhole deviation surveys collected on one centimetre increments. To avoid excessive data collection Extract choose to use one reading every 5 metres which is more than sufficient for accurately plotting a drillhole trace.
11.4 Drilling Orientation
Most of the drilling is undertaken normal to the plane of the principal mineralised orientation (where practical). The dominant drill direction at the Husab Uranium Project has been -60º towards 270º (true azimuth – Projection: UTM WGS 84 Z one 33 South). As a result, the bulk of the drillhole intersections will reflect the true thickness of mineralisation.
Coffey Mining considers the drilling directions are appropriate.
11.5 Drilling Results
A summary of the more significant drilling results ret urned from the Husab Uranium Project is provided in Tables 10.1_1 and 10.1_2 of this report. These results are derived from diamond drillholes and RC drilling. The drilling results for the resource areas are more specifically discussed in Section 17.3.
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11.6 Drilling Quality – Drillhole Database Verification
Extract have internal systems and procedures for checking their main database and the loading of data. Extract utilises the services of an expert contractor to supervise the data loading and upkeep of the resource database. Internal validation checks are also completed by Extract personnel operating from Extract’s head office in South Perth.
Coffey Mining considers the use of a robust commercial data management system, such as DataShed, and the ongoing checking of such a database to be of high industry standard. During the resource estimation process for the Husab Uranium Project, Coffey Mining checked the top 500 assays from the prospect against the original laboratory assay files. No material errors were identified during this check. Coffey Mining considers that the Extract database is of high quality and suitable for use in the resource estimation.
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12 SAMPLING METHOD AND APPROACH
12.1 Diamond Core Sampling
The portions of holes to be sampled are selected at the discretion of the geologist completing the logging. In general all zones of alaskite and internal schist are sampled. Hand spectrometer surveys on all drill core is also used to help identify zones of anomalous uranium mineralisation for sampling. The drill core is logged in detail, intervals for sampling selected and an appropriate sampling form completed. The core is then cut evenly down the middle using a diamond saw. The two halves of each piece of core are placed back in the core tray in the original position.
The drill core is sampled in 1m intervals by trained and supervised technicians. Each metre is sampled by taking the left-hand half of each piece of core for that metre and placing it into an appropriately labelled sample bag, leaving the remaining half core in the tray for reference purposes.
Calico sample bags with draw-strings are used for core sampling, identified by sample tickets. One half of each ticket (identical halves), which has a printed sequence of six digit sample numbers, is placed into the calico sampling bag. The technician completing the sampling annotates the hole number and the sample interval on the remaining portion of the sample ticket. As part of the quality control protocols, the technician verifies that the metre interval marked on the core matches the metre interval written on the sample ticket, and also matches the metre interval on the sample form. The technician also verifies that the corresponding sample number on the sample form for that interval matches the sample number on the ticket and also matches the sample number written on the sample bag.
Sample intervals are generally one metre in length, although sample intervals can be reduced to 25cm or multiples thereof within suspected mineralised zones. Once the entire metre, or fraction thereof, has been sampled and placed in the calico bag along with the sample ticket, the bag is tied firmly.
Samples for each hole are placed into large polyweave bags, with approximately 12-20 samples per bag. The bags were then numbered and labelled with the enclosed sample numbers and then taped closed.
In batches of approximately 1,000 samples, the polyweave bags are loaded onto a truck and sent with a dispatch sheet to the Genalysis Laboratory Services Pty. Ltd. (Genalysis) preparation laboratory in Johannesburg, South Africa.
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12.2 RC Sampling and Logging
12.2.1 Resource Drilling
For the resource definition drilling, 1m samples are taken. The RC samples are collected from the cyclone into a large plastic bag. The sample is then split using a three tiered splitter (with approximately 38mm wide riffles) to obtain a ⅛ split (approximately 5kg). The samples are poured evenly across the full width of the riffles (Figure 12.3.1_1).
Figure 12.3.1_1 Husab Uranium Project RC Drilling
The 5kg split is then re-split using the lower tier of the riffle to obtain an approximately 2kg to 2.5kg sample which is then sent to the laboratory for analysis (laboratory sample). The remaining sample is kept as a reference sample. A handheld spectrometer is used to obtain
an empirical eU 3O8 assay from each sample after splitting.
The sample bags are labelled in permanent marker pen. Extract uses a system of pre- numbered sample books with removable tags to ensure that the correct sample numbering is undertaken. The removable tag bearing the sample ID is placed inside each plastic laboratory sample bag and then secured with a cable tie. The stub left behind in the sample book is filled out with the hole and interval details for the sample, ensuring that there is a permanent record kept.
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Sampling consistency, accuracy of the drillhole metre intervals and sample ID information is monitored by the RC site supervisor and the rig geologist on a regular basis throughout the drilling of a hole. The mass of each sub-sample for submission to the laboratory is checked using a set of kitchen scales. The sampling observed by Coffey Mining was of a high recovery.
Prior to August 2008, sample recovery was visually checked by the supervising geologist but not routinely recorded. Visual estimates of the sample bags on site indicate a consistent high recovery. Since August 2008, Extract has commenced weighing of the sample bags for recovery estimates and recorded these results.
12.2.2 Exploration Drilling
For exploration RC or RAB drilling and where significant mineralisation is not expected to be found (e.g. alluvial cover), 5m composite samples are taken. These samples are taken using a spear. Each metre sample is homogenised, by holding the bag shut and kneading or massaging the bag and the spear is inserted through the full depth of the bulk sample.
12.2.3 QAQC Sampling
Check samples in the form of standards, field duplicates and blank samples are inserted into the sampling stream with their sample IDs forming part of the general sample sequence. Standards are currently inserted at a rate of 3 in every 100 samples such that sample numbers ending in “00”, “30” and “70” are standards. Field duplicates are taken at a rate of 3 in every 100 samples such that sample numbers ending in “36”, “66” or “96” are duplicates of intervals ending in “35”, “65” and “95”. Blank samples are inserted at a rate of 3 in every 100 samples such that sample numbers ending in “10”, “40” or “80” are blanks.
12.3 Sampling Quality
Coffey Mining has reviewed Extracts sampling procedures and observed the drilling and sampling practices in the field and found that overall they were of a high standard. All technicians appear well trained and are effectively supervised by Extract staff.
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13 SAMPLE PREPARATION, ANALYSES AND SECURITY
13.1 Sample Security
Drill samples are produced at the drill sites under the direct supervision of Extract personnel. All sampling is carried out by Extract’s field staff, under the supervision of Extract’s geological staff. All drilling samples are kept under supervision of Extract staff at their exploration camp until dispatch. Samples are currently transported via Windhoek, to the Genalysis preparation laboratory in Johannesburg. Due to the remoteness of the exploration camp and the supervision by Extract personnel, Coffey Mining considers that there is little opportunity for sample tampering by an outside agent.
13.2 Analytical Laboratories
Genalysis has been used as the principal analytical laboratory. The sample preparation is completed in Johannesburg, South Africa, and the analytical laboratory in Perth, Australia, assay the pulps.
The National Association of Testing Authorities Australia (NATA) has accredited Genalysis Laboratory Services Pty Ltd, following demonstration of its technical competence, to operate in accordance with ISO/IEC 17025 (1999) which includes the management requirements of ISO 9002:1994. This facility is accredited in the field of Chemical Testing for the tests, calibrations and measurements shown in the Scope of Accreditation issued by NATA.
Set Point Laboratories has been accredited to operate in accordance with ISO/IEC 17025 by the South African National Accreditation System (SANAS), which is responsible for the accreditation of laboratories (testing and calibration).
13.3 Sample Preparation and Analytical Procedure
The primary sample preparation and analysis is completed at the Genalysis preparation laboratory in Johannesburg, South Africa, and the Genalysis analytical facility in Perth, Australia. Coffey Mining has reviewed the Johannesburg facility and considers it to be well organised and operated. The sample preparation includes crushing and pulverizing, and a Flowsheet showing the complete preparation undergone by every sample is shown Figure 13.3_1.
The primary sample pulps are sent to Genalysis in Perth for analysis for U. Other elements that have been assayed on occasions in the past are Au, Ag, Cu, Ni, Zn, Nb, Ta, Th, C, and S. Specific gravity determinations (pycnometry) are carried out on sample pulps at a rate of one sample in 10. Some umpire pulp duplicates from very early drilling at the Husab Project have been analysed by Set Point in South Africa for uranium only by pressed pellet XRF. Umpire samples from the Husab Uranium Project have been submitted to Ultratrace, Perth. The umpire samples were analysed for U only, using a peroxide fusion followed by ICP-MS.
The majority of the U assaying has been completed using inductively coupled plasma mass spectroscopy (ICP-MS) while inductively coupled plasma atomic emission spectroscopy (ICP-AES) methods with appropriate collectors, for example 30g fire assay for gold and nickel sulphide collection for PGEs have also been applied.
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Figure 13.3_1 Flow Diagram of Sample Preparation at the Genalysis Sample Preparation Facility in Johannesburg, South Africa
Between September 2008 and February 2009, Extract included pressed pellet XRF. The change was initiated primarily to get an increase in the sample turn around time and to reduce the backlog of samples currently awaiting analyses at the lab. The pressed pellets is produced at the Perth laboratory and analysed for U, Th, Sr and Rb. U is reported to a detection limit of 5ppm.
A brief summary of the analytical methods used by Genalysis is provided below:
Au (detection limit=1ppb) by aqua regia digest, analysed by inductively coupled plasma (ICP) mass spectrometry (B/MS), and repeats analysed by graphite furnace atomic absorption spectrometry (B/ETA) with a detection limit of 1ppb.
Ag (0.01ppm) and U (0.01ppm) by ICP-MS following an aqua regia digest (B/MS).
Cu (1ppm), Ni (1ppm) and Zn (1ppm) by flame atomic absorption spectrometry following an aqua regia digest (B/AAS).
Cu (0.01%) by modified (for higher precision) multi-acid digest including hydrofluoric, nitric, perchloric and hydrochloric acids, and analysed by flame atomic absorption spectrometry (AX/AAS).
Cu by flame atomic absorption spectrometry following a four acid digest (AT/AAS).
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Uranium by ICP-MS following a four acid digest (AT/MS).
Uranium by ICP-MS following a sodium peroxide fusion (DX/MS).
Uranium by XRF (Pressed Pellet).
SG by pycnometry. The pycnometer measures the volume of powdered and porous materials as well as solid objects. From this the density of the material can be calculated. Precision is enhanced by using as large a sample as possible.
Table 13.3_1 summarises the analytical techniques used for the Husab core.
Table 13.3_1 Summary of Assaying by Laboratory for Au, Ag, Cu, Ni, U, Zn and SG
Laboratory Analytical Detection Total Element Comment Code Method Limits Samples Genalysis B/MS 1ppb 0 Au Genalysis B/ETA 1ppb 0 Ag Genalysis B/MS 0.01ppm 288 Genalysis B/AAS 1ppm 0 Cu Genalysis AX/AAS 0.01% 0 Genalysis AT/AAS 1ppm 288 Standard Method used Ni Genalysis B/AAS 1ppm 288 Genalysis B/MS 0.01ppm 0 Genalysis AT/MS 0.01ppm 362,460 Standard Method used U Genalysis DX/MS 0.1ppm 0 Genalysis XRF 5ppm 54,567 Set Point XRF 7ppm 0 Pressed Pellet Zn Genalysis B/AAS 1ppm 288 SG Genalysis /GPYCN 0.01g/cm³ 42,692 Gas Displacement pycnometer
13.4 Bulk Density Determinations
Extract currently uses the water immersion method for determining density of core billets. The bulk density measurements are discussed in Sections 17.1.5 (Zone 1) and 17.2.3 (Zone 2).
Specific gravity (SG) readings are also taken, determined by pycnometry at Genalysis. In practice, these readings can be expected to be 3% to 10% higher than the relevant bulk density in solid competent rock (e.g. a granitoid) as pycnometer determinations do not take into account porosity in the rocks or minerals.
A review of the site density data in 2010 indicated that sample recording and procedural errors were resulting in some spurious high and low density data being recorded. The site density procedures have recently been reviewed and revised recording practices have been suggested to mitigate future recording errors.
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13.5 Radiometric Downhole Assaying
Radiometric assays are collected from a Gamma Ray Spectrometer 42 (GRS) tool operated by Terratec Geophysical Services Namibia (TGS). The GRS tool uses a 250 to 500 channel Na(TI) scintillator with a crystal of 25mm diameter and 500mm length. The dead time of the detector is 4 s. Readings are provided in an electronic format in a .LAS file.
The .LAS file contains information such as Hole ID, location (latitude and longitude), date,
count per second data( total, K, U and Th channels), eU 3O8 (calculated from the total count
channel), U 3O8 (calculated from the U data), and ThO 2 (calculated from the Th channel).
Based upon the recommendation of TGS, the eU 3O8 values (derived from the total count data)
were used in the estimation studies. The U 3O8 data (source from the U channel) was not used as it was considered to be too noisy due to the relatively lower sample count.
13.6 Adequacy of Procedures
The analytical methods and laboratories adopted are considered appropriate. The sampling methods, chain of custody procedures, sample preparation procedures and analytical techniques are all considered appropriate and are compatible with accepted industry standards.
In light of the recent disequilibrium study by Extract (Culpan, 2008), it is suggested that a water correction factor be considered to any future radiometric results.
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14 DATA VERIFICATION
As the primary focus to the Husab Uranium Project is uranium exploration, the assessment of the quality control data will focus solely on the uranium data. Coffey Mining has previously analysed the QAQC data for the Garnet Valley, Ida Dome and New Camp prospects (Inwood, 2008) and found the data to be of good quality.
The following discussion is based upon QAQC report compiled by Extract in July 2010 (Wilson, 2010); however all opinions are from Coffey Mining. The results of the analysis are in line with the QAQC audit of Extract’s assay data undertaken by Coffey Mining in August 2008 (Inwood, 2008) and in the December 2008 review of QAQC data. The following review deals with the QAQC data collected from between 15 th July 2009 and 7th July 2010. Coffey Mining has reviewed the underlying data used to create the report and considers the results to be reliable. Summary plots from the QAQC report are reproduced in Appendix 1.
14.1 Standards and Blanks
14.1.1 Extract Submitted Standards and Blanks
For the RC drilling, Extract is currently inserting blank and standard samples at the rate of 3 blanks and 3 standards per 100 samples. Additionally, field duplicates are inserted at a rate of 1:33. The Extract standards have been sourced from African Mineral Standards (AMIS), a reputable laboratory.“Sample numbers ending in “00”, “30” and “70” were allocated to standards. Sample numbers ending in “36”, “66” and “76” were allocated to duplicates of the previous samples. Sample numbers ending in “10”, “40” and “80” were allocated to blanks. In addition extra blanks, up to a maximum of 5 per hole, were targeted on high grade intervals as indicated by a hand-held spectrometer. These were placed 2 samples after the high grade sample because the laboratory uses two mills and processes alternate samples through each. "
Summary control plots from these standards are in Appendix 1.
Some of the Husab Uranium Project data analysed exhibited instances of probable sample misnumbering; these outliers were excluded from the analysis. Wilson (2010) reported the standards using combined ICP and XRF data against the expected value of the standards when measured by ICP.
The results of the standards analyses are shown in Figure 14.1.1_1 and indicate overall good accuracy and precision for all of the standards. The results of standards AMIS0046, 154 and 156 exhibits positive bias of between 6.1% to 7.7%, however this bias is not seen in other standards of similar value and there is a low sample count for Standards AMIS0154 and AMIS0156.
The 5,206 blanks generally report well, with some contamination (up to 71ppm U 3O8, but
typically less than 15ppm U 3O8) evident for intervals reportedly near high-grade mineralisation.
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Figure 14.1.1_1 Extract Submitted Standards and Blanks
(Wilson, 2010)
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14.1.2 Laboratory Blanks and Standards
A total of 6 standards inserted by Genalysis were identified in the database. Summary control plots from these standards are in Appendix 1.
The results of the standards analysis are shown in Figure 14.1.2_1. Overall, the results indicate acceptable accuracy and precision for the bulk of the standards present, with a 6.5% bias indicated for the very low grade sample ORESA-45P (E.V. 2.4ppm U).
Figure 14.1.2_1 Laboratory Submitted Standards and Blanks
(Wilson, 2010)
The results of the laboratory control blanks exhibit low contamination.
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14.2 Duplicates
Extract submitted field duplicate samples at the rate of 3 duplicate per 100 samples for the RC drilling. Up to November 2009, field duplicates were taken as a split of the first split that was taken for the laboratory sample (coded as RC_SPLIT). Post November 2009, the method was changed so that the entire reject sample was re-split (coded as RC_DUP). A total of 2,267 RC_SPLIT and 2,389 RC_DUP data pairs were available for analysis.
The duplicates data was filtered for samples above 10ppm U for the Genalysis data. This was done to limit the undue relative effect that small grade differences will have on low grade samples. Summary plots of the duplicates data are located in Appendix 1.
Scatter plots of both field duplicate datasets shows a good correlation (Appendix 1) for both scatter and Quantile-Quantile (QQ) plots. Precision plots and statistics of the duplicate data indicate very high precision with 95% of the data within a 20% Rank Half Absolute Relative Difference (HARD, also referred to as Mean Absolute Paired Difference – MAPD). The RC field duplicates exhibit good precision, with between 82% and 84% of the samples returning within a 10% Rank HARD precision level. Likewise the laboratory pulp repeats exhibit very high precision levels with 97% of the data within a 10% Rank HARD precision level.
A total of 2,946 internal laboratory pulp duplicates were also analysed. A total of 97% of the data fell within a 10% rank HARD. Summary statistics for the duplicates data is shown in Table 14.2_1.
Table 14.2_1 Husab Uranium Project Summary of Uranium Precision Data
Number of % Within RANK HARD Limits Comparative Means (ppm) Sample Type Data Pairs (10%/20%) (Original Lab./Duplicate Lab.) RC Field Duplicates – Old 2,267 84/95 78/80 Method – RC_SPLIT RC Field Duplicates – New 2,389 82/95 68/68 Method – RC_DUP Umpire Pulp Duplicates 910 88/97 708/734 Laboratory Pulp Repeats 1 2,946 97/99 77/77 1 Genalysis Perth
A total of 910 pulps from Genalysis were sent to Ultra Trace for umpire analysis purposes using ICP-MS after a peroxide fusion digest. The 910 samples represented approximately
10% of the +75ppm U 3O8 samples collected during the study period. The Ultra Trace analysis reported 3.5% higher mean grade than the Genalysis analysis, with the effect being most
evident for samples between 700ppm U3O8 and 1,300ppm U3O8. A total of 88% of the samples reported above a 10% rank HARD limit.
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14.3 Data Quality Summary
The analysis of the standard and duplicate data indicates industry acceptable levels of precision and accuracy. The Umpire pulp duplicated data reports well, indicating good repeatability between laboratories. Similar quality trends are seen as for previous results analysed by Coffey Mining.
Although the screen size testwork indicates that the pulverisation is producing less than the ideal 85% of the sample passing a 75 micron sieve, the high precision levels demonstrated by the pulp and field duplicates do not indicate that this is negatively impacting with the assaying process.
The Extract assaying is considered to meet industry acceptable levels of quality and is suitable for resource estimation studies.
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15 ADJACENT PROPERTIES
The highly prospective Central Zone of the Damara Orogeny contains the Rössing uranium mine (which has produced approximately 260 million pounds of uranium since the start of operations in 1976), the Langer Heinrich uranium mine (owned by Paladin Energy), the Valencia deposit (owned by a subsidiary of Forsys Metals Corporation) and the Etango Project – (formerly referred to as Goanikontes, owned by Bannerman Resources Limited) (Figure 15_1).
Figure 15_1 Location of Uranium Deposits on Properties Adjacent to the Extract Tenement Holdings
All of the information pertaining to the adjacent properties is disclosed publicly by the owner or operator of the adjacent properties through their websites. The qualified persons and authors of this report have not been able to verify all of the information with respect to the adjacent properties contained in this report. The information on the adjacent properties is not necessarily indicative of the mineralisation at the Husab Uranium Project that is the subject of this report.
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15.1 Rössing Uranium Mine
The Rössing Mine, operated by a subsidiary of Rio Tinto, is located approximately 6 kilometres from the northern boundary of the Husab Project. The uranium bearing stratigraphy is interpreted to trend into the project areas under superficial cover (Husab Uranium Project). The bulk of the economic mineralisation associated with the Rössing deposit is hosted by alaskite on the northern limb of the “mine” synclinorium. The alaskite is preferentially emplaced into the pyroxene-hornblende gneiss and biotite-amphibole schist units of the Khan Formation in the northern ore zone, and into biotite-amphibole schist/lower marble/lower biotite-cordierite gneiss of the Rössing Formation in the central ore zone.
The alaskite is widely distributed beyond the limits of the open pit, but is not uniformly uraniferous. Portions are entirely barren or only slightly mineralised and only a few restricted sections are sufficiently mineralised to support exploitation. Alaskite hosts all primary and most of the secondary uranium minerals.
15.2 Etango Deposit
The Etango Uranium project is owned by Bannerman Resources Ltd (Bannerman) and contains Measured, Indicated and Inferred uranium resources hosted within alaskites. The deposit is located 6.5 kilometres west of the Husab Project on the western margin of the Palmenhorst Dome. The eastern extension of the Palmenhorst Dome trends into the region of Extract’s Hildendorf Prospect.
In October 2010, Bannerman publicly released resources for Etango consisting of a Measured
Resource of 62.7Mt at 205ppm U 3O8, an Indicated Mineral Resource of 273.5Mt at 200ppm U3O8
and an Inferred Mineral Resource of 164.6Mt at 176ppm U3O8 above a 100ppm U3O8 lower cutoff.
15.3 Langer Heinrich Deposit
The region also hosts a number of secondary uranium occurrences, the best documented of which is the Langer Heinrich calcrete-hosted uranium deposit, which is owned by Paladin Energy Ltd (Paladin), located approximately 30km southeast of Husab. Paladin commenced mining on the deposit in December 2006.
Calcrete cemented sediments of Tertiary age are widely distributed throughout the region and represent viable exploration targets, including in areas of alluvial cover where radiometric geophysical surveying is not effective.
In October 2010, Paladin publicly released resources for Langer Heinrich consisting of a
Measured Resources of 46.7Mt at 0.053% U3O8, an Indicated Mineral Resource of 77.6Mt at
0.055% U3O8 and an Inferred Mineral Resource of 18.5Mt at 0.06% U3O8 above a 250ppm U3O8 lower cutoff.
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16 MINERAL PROCESSING AND METALLURGICAL TESTING
16.1 Introduction
The metallurgical testwork commenced with laboratory scale batch testwork conducted at a scoping level in July 2008 and continued through to a hydrometallurgical and comminution pilot plant testwork phase that commenced in April 2010 and concluded in November 2010.
The ore body exists in two defined areas, namely Zone 1 and Zone 2. The two zones differ in terms of grade and mineralogy and have largely been tested separately. The current mining plan is to commence production on Zone 1 ore in years’ one and two, with Zone 2 being introduced in year three. The feed to the process plant will be a blend of Zone 1 and 2 ore which will vary in proportion depending on the mining plan.
The following laboratory scale batch testwork was conducted on selected drill cores that were deemed to be representative of the ore body, albeit with emphasis on samples representing the early years of production.
Head assays
Mineralogy
Comminution
Dilute leach testwork
Agitated leach testwork
Flotation
Heap leach amenability
Radiometric ore sorting
Ion exchange
Solvent extraction using ammonium hydroxide strip
Ammonium diuranate precipitation
The following aspects of the selected process were tested at pilot plant scale:
Leaching
Dilution and screening
Fines thickening
Thickener underflow re-leach
Pressure filtration of leach residue
Continuous ion exchange
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Solvent extraction using sodium carbonate strip
Sodium diuranate precipitation
Uranyl peroxide precipitation
During the hydrometallurgical pilot plant phase, selected equipment vendors were invited to conduct equipment specific tests to develop design data.
FL Smidth – filtration and thickening
Larox – filtration
Delkor – filtration
RPA Filtres Phillipe – filtration
Mixtec – agitation
Huntsman – coagulation
Flottweg – centrifuge
Comminution pilot plant testwork was conducted to confirm the selection of single stage semi- autogenous (SAG) milling as the preferred milling option due to simplicity and the ability to satisfy the minimum fines generation requirements of down-stream processes. Three circuits were selected based on comminution modelling and were configured to produce a milled product of P80<780µm:
Single stage SAG milling and a secondary pebble crushing stage.
Primary SAG milling, secondary ball milling and a secondary pebble crushing stage.
Primary SAG milling, secondary rod milling and a secondary pebble crushing stage.
16.2 Metallurgical Batch Testwork
This section summarises testwork completed in Sydney and in Perth on diamond drillholes representing various grades and rock lithologies across the length, width and depth of the Husab Zone 1 and Zone 2 deposits of the Husab Uranium Project located in Namibia. This testwork programme commenced in March 2009. The sample preparation and comminution testwork of this programme was completed at SGS Oretest. The majority of Zone 1 leaches were completed at ANSTO and Zone 2 leaches at SGS Oretest. This report also summarises the scoping level metallurgical testwork results completed on composites prepared from a single diamond drillhole, RDD002, from Zone 1. The testwork was completed at Ammtec in Perth between July and October 2008.
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16.2.1 Sample Selection and Composite Formation
Drillholes were selected based on their location, uranium grades and rock lithology to provide sufficient coverage of all parameters. There was also an emphasis on samples representing the first few years of production. The majority of composites represent the major rock type in the Zones 1 and 2 resource of uraniferous leucogranite (Alaskite and Pegmatite), representing 70% of the ore tonnes. The RDD002 composites tested in the Ammtec scoping programme represent one Zone 1 diamond drillhole whereas the Feasibility Study programme includes 13 drillholes from Zone 1 producing 44 Drillhole Composites and 15 drillholes from Zone 2 producing 62 Drillhole Composites. Drillhole Composites have been selected based on rock lithology categories as determined by field logging of the drill core. The rock lithology categories selected are alaskite (>80% alaskite or pegmatite), biotite schist (>50% biotite schist), calc silicate (>50% calc silicate), gneiss (>50% gneiss) and mixture (includes all other rock lithology distributions not matching the categories).
16.2.2 Head Assays
One of the issues encountered during the composite formation process was that over 60% of the Zone 1 Drillhole Composites selected from cross sections had higher assay grades than expected. This results in a large proportion of the composites with grades higher than the
estimated Zone 1 global resource grade of 442ppm U3O8.The assay grades of Ca, K, Mg and Si can be used to confirm the rock lithology with calcsilicate containing higher distributions of Ca, K and Mg replacing Si whereas alaskite is lower in Ca, K, Mg and higher in Si. All rock lithologies contain a range of sulphur head grades, determined to be mostly sulphide sulphur, ranging from 0 to 3,0% and has proved to have a significant impact on the leach reagent consumptions. The weighted average sulphur grades for Zone 1 composites tested is 0.44% and Zone 2 0.38%.
16.2.3 Mineralogy
Uraninite is the predominant uranium mineral present with the major occurrence as discrete uraninite grains with minor occurrences on margins, rims, veins and inclusions. Varying amounts of coffinite were identified with trace to minor brannerite and trace thorite identified (Townend, 2008-2010). No refractory uranium mineralisation (such as betafite) was identified in the samples at ANSTO Minerals (Prince & Kelly, 2009) and only traces of betafite were identified in two of the seven samples analysed by the QEMSCAN (quantitative evaluation of minerals by scanning electron microscopy) work at the University of Witwatersrand in Johannesburg, South Africa (Freemantle, 2009 & 2010). Mineralogy of leach Test L1RS -38µm feed and residue of the Alaskite composite (Prince & Kelly, June 2009) concluded that the uraninite, coffinite and thorite had leached and the brannerite showed signs of chemical attack but had not leached at the conditions of the test. Follow up work determined that the brannerite, which accounts for approximately 2% of the total uranium in the sample could be leached at higher acidity and temperature. Analysis of leach residues (Prince & Kelly, August and November 2009) confirmed the majority of uranium present in the form of brannerite with inclusions of uraninite in gangue accounting for a high distribution of the remaining losses. The three ore type composites are all comprised of the major rock forming minerals microcline, albite, quartz, mica (muscovite and/or biotite/phlogopite) with some chlorite (as clinochlore) with the distribution varying for each ore type. Minor phases observed were a Ca-Mg-Al-Fe-silicate (ferroandiopside), pyrite, sphene, apatite and ilmenite.
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Zircon is also evident in the calc-silicate and biotite schist composites. Since project initiation the focus of the mineralogical studies has shifted from the identification of uranium mineral phases and their mineralogical distribution and associations in the ore, to a need to understand differences in the distribution, composition and abundance of the gangue minerals present, and how these differences may be impacting upon the leach chemistry and reagent consumptions and the physical characteristics of the leached slurry.
16.2.4 Comminution
Extensive comminution testwork has been completed on drillhole composites and pilot plant feed composites (Hill, 2010) with the outcome that:
Two thirds of the samples are classified as having a medium-strong unconfined compressive strength (UCS) and does not impact on the selection of the type of crusher;
The average crushing work index (CWI )of ~7 kWh/t is relatively low and does not impact on the selection of the type of crusher;
SAG mill amenability tests indicate the samples are in the softest quartile of all samples tested using this method (>8,000 in total). The average A x b is 65-78 and the Dwi 3.6-4.2kWh/m³;
The abrasion index results are above average at 0.34-0.36 and it would be expected that a significant amount of abrasion wear on crushing and grinding wear components and grinding media will occur through the comminution circuit, however the impact on wear on crushing and grinding components can be reduced during the process plant design stage;
The bond rod mill work index (BRMWI) is moderate, averaging 12.4 to 13.4 kWh/t;
The BRMWI and the bond ball mill work index (BBMWI) results show an increasing work index for finer grinds below about 400µm but the work index is similar for grind sizes above 400µm;
The comminution characteristics are related to grain size with lower values produced from the coarse grained pegmatite samples compared to the alaskite samples. The biotite schist is harder than both uraniferous leucogranites.
16.2.5 Dilute Leach Testwork
Dilute leaches were completed on pulverised samples at very low densities and were aimed at determining the maximum extraction under expected plant conditions and also the maximum extraction under more severe, but still realistic conditions. The recoveries at the expected plant conditions were 96.5% for the alaskite composite, 94.5% for the biotite schist composite and 98.5% for the calcium silicate composite and the recoveries increase by between 0.4% and 1.7% at the more severe conditions.
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16.2.6 Agitated Leach Testwork
Encouraging results have been produced from the agitated leach testwork programme. The
composites appear to be relatively insensitive to grind size between grind sizes of P 80 355 to 710µm, with finer grinds yielding approximately marginal increases in recoveries. Variations in pH and/or free acid, Oxidation Reduction Potential (ORP) and ferric concentration similar uranium results, which is indicative of a robust ore where fluctuations in process plant parameters are not expected to impact on plant recoveries. The test conditions impact more on reagent consumptions than uranium recovery.
The optimum parameters selected from the ANSTO testwork programme on the alaskite composite for variability testing of the drillhole composites were:
P80 710µm;
pH 1.5;
500mV ORP with pyrolusite;
1.0g/L ferric addition;
40ºC;
55% solids.
Subsequent testing at SGS Oretest refined the conditions to:
4-5g/L free acid instead of pH control;
2.0g/L ferric concentration instead of ORP control.
A minimum of 0.5g/L ferric is required at the commencement of the leach to produce rapid leach kinetics, however the recirculation of process solutions back to the leach will provide a source of iron and iron addition to the leach may not be required. The use of a synthetic process solution for leaching appears to reduce the acid consumption by about 10%. The comparative tests of 55% and 70% w/w solids leach densities produced the same results albeit with reduced agitation speeds at 70% solids to minimise abrasion of the solids. Forty four Zone 1 drillhole composites and sixty one Zone 2 drillhole composites have been leached in the variability programme. It was noted during the ANSTO variability programme (Prince & Kelly, Nov. 2009) that there is a strong correlation between sulphur head grade and total iron in solution. It was apparent that this iron was present as ferrous and due to the leach control mechanism at ANSTO of maintaining an ORP the majority of the ferrous is oxidised to ferric, which consumes pyrolusite, thus there is a strong correlation between sulphur head grade and pyrolusite consumption. ANSTO have established from testing of the Groote Eylandt pyrolusite that 2kg/t of acid is consumed for every 1kg/t of pyrolusite consumed. As a consequence there is a correlation between sulphur head grade and acid consumption.
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Modification of the control at SGS Oretest to maintaining a ferric concentration instead of maintaining an ORP impacted positively on pyrolusite consumption and subsequently acid consumption. The sulphur grade association led to formation of sub categories under each lithology category. Sulphur grades were split at 0.16% with values equal to or less than this categorised as low sulphur and grades >0.16% categorised as high sulphur. The acid consumption is greater for all high sulphur categories (compared to low sulphur) except for Rössing Formation calcsilicate where the acid consumption is driven more by calcium and magnesium content. The increased dissolution of iron for the higher sulphur head grade samples was not directly due to the sulphides but indirectly related due to an association of iron dissolution from minerals associated with these sulphides. It appears that for Zones 1 and 2, leach performance is driven by a residue grade which remains relatively constant with varying head grades. The recommended residue grades to be used in the mining and financial models are:
55ppm U 3O8 for Zone 1;
44ppm U 3O8 for Zone 2.
Leach recoveries can then be calculated from these residue grades and head grades.
The acid consumption comparison of batch leach tests and continuous leach tests (pilot plant) concluded that no factors could be consistently applied to the variability leach tests to scale batch tests to continuous tests. However water type comparisons concluded that acid consumption when leaching in re-circulated process solution is about 10% less than leaching in Tap Water (either Sydney or Perth) thus a factor can be applied to the average acid consumption:
Zone 1 and 2 low sulphur Rössing Formation alaskite 14.6kg/t acid consumption;
Zone 1 high sulphur Rössing Formation alaskite 18.7kg/t acid consumption;
Zone 2 high sulphur Rössing Formation alaskite 27.5kg/t acid consumption;
Zone 1 low sulphur Rössing Formation Rössing 21.8kg/t acid consumption;
Zone 2 low sulphur Rössing Formation Rössing 18.2kg/t acid consumption;
Zone 1 high sulphur Rössing Formation Rössing 29.9kg/t acid consumption;
Zone 2 high sulphur Rössing Formation Rössing 22.0kg/t acid consumption;
Zone 2 low sulphur Chuos Formation Chuos 16.1kg/t acid consumption;
Zone 2 high sulphur Chuos Formation Chuos 21.8kg/t acid consumption.
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The acid consumption of the Rössing Formation calcsilicate is a function of Ca + Mg grade and is not influenced by sulphur content. Without Ca + Mg resource data the acid consumption for this category has been determined by calculating the weighted average acid consumption of the 7 x drillhole composites (Zone 1 and 2 combined) of 46.0kg/t and application of the water type factor reduces this to 41.4kg/t. This consumption can be applied to all calcsilicate from both zones irrespective of sulphur grade.
16.2.7 Flotation
Flotation regimes at pH 2.5 and natural pH have produced high recoveries with lower flotation tail grades than leach residue grades on a single composite, however the flotation programme was plagued with poor repeatability. The option of flotation maybe negatively impacted by the requirement to grind finer than for leaching, higher unit costs for flotation reagents compared to acid.
16.2.8 Heap Leach Amenability
Heap leach amenability testing in the form of bottle rolls have recovered between 53 and 81% of the uranium at crush sizes of -12.5mm on a number of drillhole composites ranging in rock lithology and grade. The recovery is crush size dependant with the size fraction recoveries decreasing significantly above about 4.75mm. Column leach testwork completed for over 100 days on a resource grade drillhole composite, RDD018A, recovered 68% and 64% at crush sizes of -6.3 and -12.5mm respectively. This results in column residue grades of
138 and 157ppm U3O8 respectively and compares to 35ppm from a P 80 500µm agitated leach. Heap leach amenability recoveries are not high enough to continue testing and no further heap leach amenability testing will be completed.
16.2.9 Ion Exchange
A commercial strong base gel type I resin, Amberjet 4400Cl, widely used in the industry for the uranium recovery from sulphate solutions, was used for the testwork.
The main findings were as follows:
High equilibrium loadings were achieved on the resin (77g/L U3O8 [wet settled resin] wsr) under standard loading conditions (pH 1.5 and 40°C) and a feed concentration of
1.4g/L U3O8 in the PLS. The equilibrium loading for a projected PLS tenor of ~400mg/L
U3O8 in the PLS is also high at 65gg/L U 3O8wsr);
Loading rates were as expected for this type of resin;
The impurity load on the fully loaded resin was low, notably with respect to Fe. The Fe/U
ratio on the loaded resin was 0.18% and this was improved with 0.05 M H 2SO 4 scrubbing (3BV) prior to elution;
The concentration of Si on the loaded resin from a single column load cycle was measured at ~3g/L. With repeated cycles of loading and elution, the impact of increased Si concentration needs to be monitored in continuous trials to determine whether periodic caustic regeneration is required;
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Vanadium was effectively rejected by the resin;
The elution rate with 1M sulphuric acid was suitable and <10BV were required for column
elution to achieve <1g/L wsr U 3O8 in the eluted resin;
The eluate produced was relatively clean, but due to the sulphate content (90g/L) it required sulphate removal by gypsum precipitation prior to direct precipitation of uranium with hydrogen peroxide;
Uranium precipitation with H 2O2 resulted in nearly quantitative metal recovery, with 1.6mg/L uranium left in barren;
The uranyl peroxide produced experimentally was of sufficient quality to meet the less stringent converter specifications. In this precipitation testwork, the lime used was not of sufficient high quality and caused contamination of the uranium product, notably with respect to Fe. With improved lime quality, it would be expected that a uranyl peroxide precipitate could be produced that would be of sufficient quality to meet all converter specifications.
16.2.10 Solvent Extraction
Some challenges were experienced with the stability of the composite leach solution. It was found that on standing, the filtered solution became turbid, possibly due to bacterial or algae growth or the formation of the insoluble Si compounds.
The laboratory batch solvent extraction testwork was performed using a mixture of 5 vol.% Alamine 336 and 2 vol.% Isodecanol in Shellsol 2046. This solvent composition is widely used in industrial processes for the concentration and purification of uranium.
The main findings were as follows:
Uranium loading was very favourable with a maximum equilibrium loading of up to 5.7g/L achieved in the batch tests;
The physical characteristic of the loaded solvent in the batch tests were not found to be favourable, with emulsion formation observed at >3g/L uranium loading;
The impurity load on the solvent was very low and it was demonstrated in batch tests that a scrub circuit would be of limited benefit;
Stripping of solvent with ammonium sulphate/ammonia was tested and found to be very favourable;
A batch bulk strip followed by precipitation of ammonium diuranate with ammonia addition produced an ADU product meeting the most stringent converter specifications.
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Overall, the ion exchange testwork was very encouraging with high equilibrium loadings, reasonable uranium selectivity, suitable loading and elution kinetics and the possibility to control the product purity with respect to iron with the introduction of a scrubbing step. In a plant situation, the actual loading is expected to be considerably lower than the equilibrium loading. In addition, the resin loading will also be lower than tested here due to lower projected uranium concentrations in the PLS. The batch data generated in this testwork can be used to model potential IX processes, either fixed bed, semi-continuous or fully continuous. The predictions will then need to be experimentally confirmed. From the batch laboratory testwork, it was recommended that an IX process be considered as the relatively low concentrations of uranium in the PLS (350mg/L) combined with the issues associated with PLS stability and filtration difficulties make it an ideal candidate for IX technology. This preferred process configuration was tested continuously in a pilot plant in association with a leaching circuit to provide representative leach liquors. Chemically, a solvent extraction process would perform well, with the production of ADU expected to be well within converter specifications. However, taking into account the observed issues associated with PLS stability and solvent emulsification, the relatively low projected concentrations of uranium in the PLS and the aggressive development timetable, it is not recommended that a direct solvent extraction process be considered as a baseline option. From the batch work, it was noted that continuous testing for solvent extraction would be imperative, in order to determine whether these operational issues can be overcome for consideration in an eluex (IX/SX) circuit configuration.
16.2.11 Precipitation
The IX eluate was used for precipitation with H2O2 and the SX strip solution was used for ADU precipitation. Analysis of the IX eluate specifically with respect to iron indicated that there had been contamination of the uranium product introduced via the lime used for gypsum precipitation. When the estimated contribution of the impurities derived from lime is deducted from the total amount of contaminants, the estimated purity of the final uranium product, shown in Table 16.2.11_1, is well within the specification range.
16.2.12 Radiometric Ore Sorting
Analysis of the radiometric ore sorting testwork completed by UltraSort determined that the
radiometric sorting process is inefficient below about 250ppm U3O8.
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Table 16.2.11_1 Uranium Product Analysis
Estimated IX SX Cameco Converdyn Comurhex Element Composition (%) (%) (% wrt U) (% wrt U) (% wrt U) of IX Ppt.* (%) Al 0.10 0.002 0.05 0.10 Ag <0.003 <0.003 0.010 As <0.003 <0.003 0.010 0.010 1.00 Ba <0.03 <0.003 0.010 Ca 0.37 <0.03 3.00 0.050 1.00 Cd <0.003 <0.003 0.010 Cr <0.003 <0.003 0,010 Fe 0.25 0.001 0.01 1.00 0.15 K 0.03 0.05 0.20 Mg <0.03 0.02 3.00 0.020 Mn 0.004 0.004 Mo <0.003 <0.003 0.10 0.10 0.050 Na 0.28 0.006 0.03 0.50 1.00 Na + K 0.31 0.007 0.07 1.00
PO 4 0.34 0.11 0.08 0.20 0.10 1.00 Pb <0.003 <0.003 0.010 S 0.22 1.00 SO4 0.66 1.31 1.00 3.00 Se <0.003 <0.003 0.010 Si 0.13 0.01 0.05 0.50
SiO 2 0.28 0.02 0.10 0.50 0.50 Th <0.003 <0.003 0.50 0.10 Ti 0.02 0.0008 <0.003 0.50 0.10 U 79 71 75
V2O5 <0.01 <0.005 0.30 Zr <0.003 0.013 0.10 0.01 0.20 * Excludes contamination from lime addition
16.2.13 Other Testwork
Float/sink heavy liquid separation testwork completed on three size fractions from a P80 1mm grind at varying SG’s produced an improved separation with increased liberation, but the light fractions contained significant uranium values and heavy or dense media separation is not likely to be effective in upgrading the Zones 1 and 2 ore. There was some evidence of silica polymerisation in the IX/SX testwork and the vendor filtration testwork which was supported by microfiltration tests. Following these outcomes and poor phase separation in the Pilot Plant SX circuit trials with Polysil coagulants concluded that 60-80mg/l of RM1250 was required to minimise crud formation in the SX circuit. The RM1250 was added to the CIX eluant prior to feeding into the SX Extract circuit.
Pyrolusite activity tests were completed on six pyrolusite sources with the best results from:
UF74 – sourced from Morocco and the current source of pyrolusite used at Rössing;
The manganese grades of the samples vary and there is a correlation between Mn:Fe ratio and available MnO2 but for UF74, due to its very low iron content and high Mn:Fe ratio.
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The preferred product from this testing was UF74 and this was used during the Pilot Plant programme and has been recommended as the preferred supply to site. A sample of Assmang Hematite was tested to determine the dissolution of iron for addition to the leach circuit. The test confirmed that Assmang Hematite was a possible source of ferric for the leach circuit, although the Pilot Plant programme concluded that the leach circuit was self sufficient in iron. In the early stages of acid leach testing and process flowsheet development there was an expectation that neutralisation of some acid maybe required. Surface samples of “limestone” sourced from in and around the Husab Uranium Project were collected for testwork. Activity tests determined that 1.26t of limestone was required to neutralise 1.0t of acid at P80 75µm grind size. The process flowsheet does not include neutralisation but if neutralisation is required at any stage of the Project life the near mine samples could be used for this process.
16.3 Conclusions from Batch Laboratory Testwork
Comments on the outcomes of the study testwork programme include the following:
SAG mill amenability tests indicate the samples are in the softest quartile of all samples tested.
High uranium recoveries with >60% of the drillhole composite leach recoveries >90%.
The leach acid consumptions are generally below 25kg/t with excursions above this due to high Ca and Mg ore or high Fe releasing ore, which contributes to oxidation of excessive ferrous to ferric and a resultant increase in pyrolusite and acid consumptions.
The ore is relatively insensitive to grind size with incremental increases in uranium recovery at finer grind sizes.
Reducing the leach temperature from 35ºC to 40ºC to 25ºC to 30ºC impacts on the leach kinetics but has minimal impact on the final leach recovery.
IX testwork shows good selectivity of uranium from other impurities with low levels of contaminants feeding downstream processing.
High quality precipitate produced from either H 2O2 precipitation or ADU.
Heap leach amenability testing produced moderate recoveries of 64% to 68% at crush sizes of -6.3mm and -12.5mm.
Radiometric ore sorting and dense media separation (DMS) or heavy media separation (HMS) were determined to be ineffective in upgrading the ore.
Pre-concentration by flotation showed some promise, but the testwork data was not reproducible.
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16.4 Metallurgical Pilot Plant Testwork
The pilot plant metallurgical testwork programme consisted of two campaigns, namely a confirmatory comminution at Mintek in South Africa (Mokwena et al, 2010) and hydrometallurgical confirmatory batch testwork campaign with the main pilot plant campaign at SGS in Perth (Jayasekera et al, 2010) consisting of three separate pilot plant runs totalling approximately 800 hours of operation.
16.4.1 Introduction
In the pilot plant tests, run one was a five-day reliability run while runs two and three (14 days each) processed ore representative of the early mine production years. Throughout the run, various samples were provided for selected vendor testwork. Off-line metallurgical testwork was performed by SGS, F.L. Smidth Minerals, Larox Pty Ltd, Huntsman, Turnbery RPA / Filtres Philippe and Flottweg.
Husab PQ drill core (85mm core diameter) samples were prepared to -5mm (top size) using a jaw crusher, high pressure grinding rolls and cone crusher prior to commencement of run 1. The sample was later re-crushed using a cone crusher to -2mm (top size) to better target the leach feed size distribution. Run three of the continuous hydrometallurgical piloting campaign conducted at SGS included the following stages:
Re-pulping;
Leaching;
Dilution and screening;
Fines thickening;
Thickener underflow re-leach (latter part of the run);
Filtration;
Continuous ion exchange (CIX), including a regeneration stage;
SX;
Recovery and refinery - SDU precipitation, sulphation, peroxide precipitation.
16.4.2 Ore Preparation
PQ core samples (ca. 9.6 tonnes) from the Husab deposit were received at SGS Lakefield Oretest Pty Ltd (Malaga, WA) during February, 2010. All core samples were weighed, logged and split into composites. Ore was separated into composite samples based on the mining plan, spatial location, lithology and down hole spectrometer readings. Sufficient representative material from selected composites comprising core samples and samples crushed to a nominal top size of -35mm, was set aside for physical parameter testing (comminution testwork). All remaining material from these composites was stage-crushed to -16mm top size.
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The remaining core was separated into assay composites and non-assay composites. Assay and non-assay composites were stage-crushed to a -16mm top size. The assay composites were rotary blended and a 1kg sub-sample was pulverized and assayed. Based on spatial location, lithology, down hole spectrometer readings and composite uranium assays, composite samples were grouped into yearly composite samples (reliability run, Zone 1 – year 1, Zone 1 – year 2, Zone 1 – year 3 and Zone 2). All yearly composites were then individually crushed through laboratory high pressure grinding rolls (HPGR). The HPGR product was screened at 5mm with the oversize being stage-crushed (laboratory jaw and cone crusher) to -5mm top size. Each yearly feed type was individually blended and split into 20kg bags. Prior to run three, all Zone 1 – year 2 and Zone 1 – year 3 material were crushed to a -2mm top size and split into 10kg charges to accommodate the batch dry milling process.
16.4.3 Batch Dry Milling
Each composite sample was crushed at -2mm top size and batch dry milled in 10kg charges for 20 minutes in a ball mill to provide feed for the leach and subsequent pilot plant testing.
16.4.4 Re-pulp
Dry ore (10kg/h) and CIX barren solution were mixed in a 6.5L agitated tank to achieve 71% (w/w) solids. The re-pulp discharge was an overflow arrangement that fed the leach circuit.
The Re-pulp circuit operated for 334 hours during run three with >99% availability.
16.4.5 Leaching
The leach circuit consisted of six tanks arranged in a cascade allowing gravity flow between tanks via an overflow launder fitted with a pneumatic vibrator to aide slurry flow. The leach tanks were fitted with water jackets to control the slurry temperature in the range 35°C to 40°C. Dilution of the slurry was achieved with SX raffinate (added to tank 2) and de-ionised water (to tank 3). Target leach discharge slurry density was 69% (w/w) solids. The leach tanks were fitted with down comers prior to the start of run three. Initially, concentrated (98%) sulphuric acid was dosed to tank 1 (single dosing point) at a target flow rate equal to 21kg/t to 23kg/t of feed ore, in order to improve extraction due to the higher free acid concentration. Sulphuric acid addition was controlled to target a free acid (sulphuric acid) concentration of 5.5g/L. On 15 May 2010, deionised water addition to tank 3 was replaced with 41g/L sulphuric acid which was added at the same rate to target an acid addition of 1kg/t to tank 5. On 16 May 2010, acid addition to the leach was modified. Acid was added to tanks 1 and 3 to target 5g/L free acid in tanks 2 and 4. The target (tank 6) discharge free acid was 2g/L to 3g/L free acid. The lower free acid concentration was designed to minimise the dissolution
of undesirable ions such as iron. Finely ground (P 80 -75µm) UF74 pyrolusite was manually added every 15 minutes with 77% to 80% of the dose being added to tank 1 and 20% to 23% to tank 3. Pyrolusite was dosed at a rate of 2kg/t to 3kg/t to target a ferric to ferrous ratio (Fe 3+ :Fe 2+ ) of 1.5 to 1.8.
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Provision was made to add ferric sulphate (Ferriclear) solution to tanks 2 and 3 using individual dosing pumps. However as in run 2, this was not required during run three. Agitation was provided by a double, rubber coated, 4-bladed axial flow turbine impellers. The impellers were removed at the termination of run 3 and there was noticeable wear on the leading edge of all impellers. The volume of each leach tank was approximately 17 L and the circuit retention time was approximately 12 hours. A tracer tests was undertaken in the leach circuit during run three. Leached slurry was discharged to duty and standby surge tanks (one filling and one emptying) each located on a weigh scale which fed the d ilution and screening circuit. The Leach circuit operated for 334 hours during run 3 with >99% availability.
16.4.6 Dilution and Screening
Leach discharge was pumped to a dilution tank (2,7 L) where it was diluted with CIX barren, filtration liquors and thickener overflow solution. CIX barren flow to the dilution tank ceased on 13 May 2010, in order to increase the uranium tenor in the thickener overflow (CIX feed). The dilution tank discharge was pumped to a vibrating 106µm Kason screen where the stream was split into coarse and fine fractions simulating a commercial cyclone operation. Screen oversize was collected on an hourly basis and fed to the batch filtration circuit. Screen undersize fines were fed to the fines thickening circuit.
16.4.7 Fines Thickening
Screen undersize slurry was pumped from the surge tank to the fines thickener pre-mix tank where it was diluted with thickener overflow to <5% w/w solids. Magnafloc 351 flocculant (hydrated in deionised water to 2.5g/L and then diluted in CIX barren to obtain 0.5g/L) was added in-line as the slurry was pumped from the pre-mix tank to the thickener feed well. Target flocculant addition for the entire circuit was 50g/t of pilot plant feed. The thickener overflow was fed to a settling tank which was subsequently fed to the CIX circuit. The underflow slurry was batch transferred hourly to the filtration circuit. De-ionised water was added to the thickener pre-mix tank at a rate of 120mL/h to simulate loss from evaporation. The screening and thickening circuits operated for 326 hours for an availability of >97%.
16.4.8 Underflow Re-leach
The re-leach circuit was operated from 18 May 2010 until the end of run three. A bleed stream of the thickener underflow, roughly one third of the total flow, and CIX barren (1:1 w/w ratio) was fed to a 17L agitated leach tank. The re-leach was operated at 40 °C. Pyrolusite was added to target an ORP of 500mV. The product slurry overflowed into a bucket which was fed to the filter in hourly batches.
16.4.9 Filtration of Leach-End Residue
The fines thickener underflow and screen oversize were combined and diluted to approximately 60% with CIX before gentle blending and flash flocculation. At the end of run three, the re-leach product was also fed to the filter. The combined slurry was fed to two LaroxPF01 filters. Filter operation was adapted to simulate belt filter efficiencies which are to be used in the commercial plant.
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Three separate washing stages were undertaken during run three. Equal volumes of CIX barren liquor was used for washes 1 and 2 and Perth tap water was used for wash 3. Wash 3 volume (Perth tap water) was reduced during the run by ~75% to aid the bleeding of undesirable salts in the filter cake moisture. The formate and all washates were combined and fed to the screen as dilution to target a screen undersize density. Moist filter cakes were weighed, sampled and stored.
16.4.10 Continuous Ion Exchange
Pregnant leach solution (PLS) from the fines thickener overflow was pumped to a surge tank mounted on scales ahead of the CIX adsorption circuit. CIX barren (cell 11 discharge) was recycled to the re-pulp, screening and filtration stages. The adsorption circuit consisted of 11 stages and the resin moved counter current to the PLS in nitrogen agitated cells, setup in a cascading train. The volume of each cell was ~11.5 L. Humidified nitrogen was sparged at 2 to 3 L/min to each cell. The solution overflowed from one cell to the other whilst resin was retained in the cell using a 425µm screen placed in the opening of the overflow. The elution circuit consisted of 8 stages of counter current elution with solvent extraction raffinate. The volume of each cell in the stripping circuit was ~1.5L. The resin employed was the used resin left at the end of run 2 and had reduced in volume from 300mL per stage to approximately 280mL per stage. There were a total of 20 resin batches in the CIX circuit (adsorption and elution). Both circuits were operated at ambient temperature.
There was one additional standby cell placed adjacent to the adsorption train. The loaded resin (topmost cell in the adsorption train) was separated from the solution by passing through a 350µm screen and then transferred to the last cell in the elution circuit. Likewise, the fully eluted resin (topmost cell in the elution train) was recovered and placed in the stand-by cell in the extraction circuit. They were then moved up to the next position in the train. The indexing procedure involved some back mixing when the contents of the top most cell in the adsorption and elution trains were drained back into their respective feed tanks while recovering resin. Back mixing required the CIX adsorption and elution feed rate to be increased proportionally. Indexing of resin was typically performed at 1½ to 4 hourly intervals, depending on the uranium flux and resin loading.
Concentrated eluate from the CIX elution circuit was collected in a surge tank mounted on scales. This was then transferred in batches to the SX PLS tank on scales to be fed into the SX circuit. Sulphuric acid concentration in SX raffinate was adjusted to 100g/L, prior to being
fed in batches to the CIX eluant feed tank. Towards the end of the run, sulphuric acid (H 2SO 4) concentration was increased to 130g/L. Scrub product solution was batch transferred to the CIX elution stage and fed to elution cell 2, due to the residual uranium tenor. At times during run 3, scrub product was not fed to the elution stage due to high uranium concentrations in the product solution. In these instances scrub product was recycled to the SX PLS tank.
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All resins were regenerated between 13 and 15 April 2010. The regeneration process entailed removing the eluted resin after it had been removed (indexed) from the S1 position. The resin was then washed in de-ionised water and placed sequentially in 20g/L sodium
hydroxide (NaOH), 60g/LNaOH, 20g/LNaOH, de-ionised water and 2g/L H2SO 4, respectively, for 15 minutes each. Once the regeneration process was complete the resin was placed into the standby position in the CIX extraction circuit and put back into operation once the next index was performed. It took roughly 2½ days to complete the regeneration process on all resins. Each resin was only regenerated once during run three.
The CIX circuit operated for a total of 326 hours with >97% availability during run three.
16.4.11 Solvent Extraction
The solvent extraction circuit included four extraction stages, three scrubbing stages and three stripping stages.
All mixer-settlers and after-settlers were constructed from clear PVC with welded joints, rather than glued. All cells were of the same dimension; mixers were 50mm wide, 50mm long and 40mm deep (100mL live volume) whilst settlers were 50mm wide, 160mm long and 50mm deep (400mL live volume). After-settlers comprised a mixer box and three compartments with baffled weir arrangements. The settlers housed a heat exchanger to maintain the circuit temperature of ~40°C.
In the mixer boxes organic and aqueous were mixed using slotted disc impellors 30mm in diameter. Each disc contained 4 slots, 4mm wide and 5mm deep, cut across its diameter. The agitators were operated between 1100 to 1300 rpm.
All circuits were operated in a counter-current mode, with aqueous and organic from the previous stages being introduced to the mixer under a slotted disc impellor. The impellor operated as a pump mixer, drawing the organic and aqueous into the mixer and the resulting mixture then overflowed a weir into the settler where the phases were disengaged. The organic and aqueous overflowed the settler via individual weirs and advanced to the next stage.
In all mixers, O:A ratios were maintained close to 1:1 with internal recycling of the aqueous from the settlers to compensate for varying advance O:A ratios. The aqueous(PLS) and organic (barren organic) flow rates were measured hourly and adjusted as necessary to achieve the required advance O:A ratio. Initially, all mixers were run with organic continuous phase continuity but towards the end of the run, E1, S3 and SC3 operated aqueous continuous whilst E4 and S1 ran organic continuous.
The extractant used was 6%(v/v) Alamine 336 (supplied by Cognis) together with 3%(v/v) isodecanol as a phase modifier in the diluent Shellsol 2325.
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The CIX eluate (SX feed) was pumped to a header tank, to supply a constant head pressure to the pump suction, and then on to the extraction circuit E1. Two types of Huntsman coagulants (Polysil RMB 2050 and RMB 1250) were dosed to the CIX eluate tank at varying concentrations throughout run 3 in an attempt to control stable emulsion or crud formation in the extraction circuit. The raffinate (E4 aqueous) reported to an after-settler and passed through an activated carbon column to remove any entrained organic, prior to returning to CIX stripping circuit as the eluant. A bleed of the raffinate was also returned to the leach circuit. The packed carbon columns were 30mm in diameter and 400mm in length.
The loaded organic from extraction (E1 organic) reported to an after-settler to lower aqueous entrainment. The circuit incorporated three scrub stages for impurity removal. The scrub circuit configuration was altered a number of times during run 3. In the initial set up, the
loaded organic (E1 organic) was contacted with water acidified with H 2SO 4 to pH 2.0 in the first scrub stage (SC1) and the product was collected separately. The aqueous scrub feed to the scrub stage 2 (SC2) was SDU barren solution from the refinery adjusted to pH 3.0 in a pre-mix tank. The aqueous product from SC2 allowed back to the pre-mix tank, essentially setting up a re-cycle of aqueous flow between the SC2 and the pre-mix tank at a flow rate equal to organic flow (i.e. O:A 1:1). Scrub 3 aqueous feed was also from the pre-mix tank at a flow rate equal to the SDU barren solution (pH adjusted) feed to the pre-mix tank. Scrub 3 aqueous product (SC3) was collected and mixed with SC1 product as the scrub product to be re-cycled to CIX strip circuit. Due to elevated uranium concentrations in the scrub product this scrub product was recycled back to the SX PLS tank. In another version of this configuration, SC1 aqueous product was collected in the pre-mix tank and the flow rates of SDU barren solution and SC3 feed solution were adjusted to compensate for additional flow into the pre- mix tank.
Above configuration was altered again on 18 May 2010. The aqueous feed to the SC1 was
changed from pH 2.0 water to uranium oxide (UO 4) barren washes 1 to 4 from the refinery.
The aqueous scrub feed to the SC3 was UO 4 primary barren solution. The aqueous product from SC3 flowed counter currently to SC2 and the SC2 aqueous product was collected. The aqueous product from SC1 was collected separately and combined with SC2 aqueous product. This was combined with SX raffinate (E4 aqueous) and acidified to obtain 130g/L free acid and fed to CIX strip 1.
The scrubbed organic (SC3 organic) reported to an after-settler before entering the strip circuit. Stripping was conducted with two solutions; a 1.5 M sodium carbonate solution was fed to the S3 stage while pH adjusted SDU barren solution (pH between 10.1 and 11.0,
adjusted with concentrated H 2SO 4) was fed to the S1 stage. S3 aqueous product flowed counter currently to S1. The strip product (S1 aqueous), was collected and transferred to the refinery as the SDU feed solution.
At the commencement of run 3, the SX circuit was filled (all the mixers and settlers) with solutions (aqueous and organic) removed from the circuit at the completion of run 2. The organic extractant (Alamine 336) concentration of 10%(v/v) used in run 2 was diluted with Shellsol 2325 to achieve an extractant (Alamine 336) concentration of 6%(v/v) prior to the commencement of run 3.
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The SX circuit operated for 320 h from 8-21May 2010 with a >95% availability.
16.4.12 Recovery and Refinery
The recovery and refinery stages were semi-batch operations which were operated daily (day shift only).
Sodium Diuranate (SDU) Precipitation
SDU (Na 2U2O7-7H 2O) was precipitated from the solvent extraction strip product at 35°C in a 10L flat-bottomed agitated vessel. A 400% (1:4) seed recycle was employed to promote precipitate growth. Strip product solution and a calculated amount of 250g/L NaOH were dosed to the SDU seed material over a period of 2 hours. The product slurry was then allowed to agitate for 4 hours. Slurry pH, ORP and temperature were monitored throughout the batch tests. A portion of the SDU product slurry was placed in a 1L cylinder and a settling test was performed. The SDU slurry was allowed to settle overnight and the supernatant decanted. The thickened slurry from the settling test was recombined with the settled slurry in the reactor vessel. A 20% portion of the settled SDU slurry, with 80% remaining as seed, was then put through a 5stage de-ionised wash water cycle to prepare the solids for the sulphation stage. The SDU barren solution was divided, with 80% being acidified to a pH between 10.1 to 11.0 and advanced to SX strip (S1 mixer) while 20% was fed to the solvent extraction scrub pre-mix tank.
Sulphation
The washed SDU solids in wash 5 solution were agitated in a 2 L flat-bottom flask agitated by a magnetic stirrer and equipped with a nitrogen sparge and 2 dreschel bottles acting as carbon dioxide scrubbers. The slurry was maintained at 60°C and acidified to approximately pH 1,5 with 98% w/w sulphuric acid. An appropriate amount of deionised water was added to the system to target 50g/L uranium in the product solution. The system was sparged with nitrogen as a carrier gas for four hours to transport evolved carbon dioxide through the 2 scrubber vessels which contained 200g/L sodium hydroxide. The resulting solution containing uranyl sulphate was then advanced to the peroxide precipitation stage. The product solution was filtered prior to the peroxide precipitation stage.
Peroxide Precipitation
Subsequent to the sulphation stage, the slurry temperature was decreased to 38°C. Hydrogen peroxide at 30% (w/w) and 200g/L sodium hydroxide were added alternately to maintain a target pH of 1.5. Upon reaching the hydrogen peroxide end point, sodium hydroxide addition continued until a pH of 3.0 to 4.0 (batch dependent) was reached. The sample was then allowed to agitate at temperature for three hours. The solids were collected by centrifuge before being washed twice via re-pulping and either filtering or centrifuging. A portion of the solids were dried and sent for assay with the remainder being collected and stored in deionised water.
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16.4.13 Vendors
All vendor testwork was performed on-site at SGS during run three. F.L. Smidth Minerals, Larox Pty Ltd, Mixtec, Huntsman, Turnbery RPA / Filtres Philippe and Flottweg performed thickening, filtration and centrifuge testwork on Zone 1 – year 2 and Zone 1 – year 3 material produced during run three. SGS performed rheological testwork and particle sizing, as per the project proposal, on selected streams.
F.L. Smidth Minerals performed thickening testwork on pilot plant feed material (pre- leach thickening). Thickening tests were also conducted on SDU product and fines thickener feed slurries (-106µm screened leach discharge). Filtration testwork was performed on combined screen oversize and fines thickener underflow slurries. The tests were performed at SGS from 13-21 May 2010.
Larox performed filtration testwork on leach discharge slurry and combined screen oversize and fines thickener underflow slurries from 11 to 21 May 2010.
Mixtec performed agitation tests on Zone 2 material (repulped filter cakes from run 2) from 6 to 8 May 2010. Testwork was also performed on Zone 1 - year 2 material (re-pulped filter cakes 137-163) at the completion of run three.
Huntsman evaluated the use of Polysil coagulants in the SX extraction stage on crud formation, phase disengagement and silica concentrations in the aqueous streams. Various pilot plant streams were monitored to determine soluble and colloidal silica concentrations.
Turnbery RPA / Filtres Philippe performed filtration testwork on leach discharge slurry and combined screen oversize and fines thickener underflow slurries from 12-17 May 2010.
Flottweg performed centrifuge testwork on peroxide precipitation product (UO4) and SX stable emulsions and crud from 18-21 May 2010.
SGS performed rheological testwork and particle sizing on various pilot plant streams (Zone 1 – year 2 and Zone 1 – year 3 material) throughout Run three.
16.4.14 F.L. Smidth – Filtration
F.L. Smidth Minerals (FLSM) conducted sedimentation and vacuum filtration testing on a sample generated by the pilot plant operated at SGS Laboratory in Perth, Australia. Testing was conducted on site at SGS in April and May 2010 to determine operating parameters for a commercial horizontal belt filter.
SGS provided the feed sample for the leach discharge stream. The majority of the testwork was done on samples by replicating the sending of leach discharge through a hydrocyclone. The fine slurry and the coarse slurry were re-blended at various ratios, mixing techniques, flocculant dosages, and percent solids to determine the optimal operating characteristics. Samples of the leach discharge prior to separation into fine and coarse particles were also tested in order to compare filtration rates for a system without hydrocyclones to a system with hydrocyclones.
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The Husab vacuum filtration results are shown in Table 16.4.14_1 below.
Table 16.4.14_1 Husab Uranium Project Husab Vacuum Filtration Results
Zone 1 Uranium Feed Zone 2 Year 1 Year 2 Year 3 Coarse to Fines Ratio (dry weight basis) 75/25 75/25 70/30 70/30 Feed Solids Concentration (wt%) 60% 65% 65% 65% Flocculant 920SH 920SH 920SH 920SH Floc Dosage (grams per ton of dry fines) 200 200 250 250 Floc Dosage (grams per ton of dry solids) 50 50 75 75 Sizing Basis (kg/hr-m²) 2200 3000 3700 2500
It was concluded that the F.L. Smidth bench-scale filtration equipment successfully dewatered the Husab uranium leach discharge sample in Perth, Australia in April and May 2010. The objectives of the bench-scale testing were:
Determine impact of coarse to fines ratio on filtration rates
Determine impact of flocculant dosage on filtration rates
Determine impact of slurry percent solids on filtration rates
16.4.15 F.L. Smidth – Thickening Testwork
F.L. Smidth Minerals (FLS) conducted thickening work to establish the thickener underflow density that could be achieved on a laboratory scale, and to size and design the optimal thickening system.
Various thickening technologies were tested so as to offer the best cost competitive solution with optimal results.
Four different ore types were tested for leach feed and leach stages. SDU and UO4 samples were also tested.
Two types of FLS technology were evaluated:
Hi-Rate Thickening
Hi-Density Thickening
Four different ore types were tested for both pre-leach and post-leach. The various slurries and dilution liquors were obtained from SGS personnel and they were taken to be representative of the material to feed the thickeners. The settling flux and settling tests were performed using 500mL and 4000mLcylinders respectively.
Flocculant screening was done using three non-ionic flocculants. Floerger SNF FA 920 SH (non-ionic) proved to be the best flocculant for Husab samples in terms of settling rates, bed compaction and overflow clarity.
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The summary of results obtained from the testwork is shown in Table 16.4.16_1 and Table 16.4.16_2.
Table 16.4.16_1 Husab Uranium Project Thickening Testwork Summary
Hi-Rate Thickeners Leach Feed (-106µm) Ore ID Zone 1 Zone 2 Year 1 Year 2 Year 3 Solids SG 2.7 2.7 2.7 2.7 Liquor SG 1.076 1.076 1.076 1.076 Optimum Solids wt% 10 10 10 6 Flocculant Dosage g/t 40 60 60 60 Initial Settling Rate m/hr 19.6 22 15 23.7 Settling Flux (t/m²h) 0.267 0.235 0.255 0.310 O/F Clarity g/m³ 1330 1195 4820 1045 Underflow Solids wt% 64 64 62 60 Yield Stress (Pa) 25 20 20 25 Leach Discharge (-106µm) Ore ID Zone 1 Zone 2 Year 1 Year 2 Year 3 Solids SG 2.7 2.7 2.7 2.7 Liquor SG 1.086 1.086 1.086 1.089 Optimum Solids wt% 3 4 4 3 Flocculant Dosage g/t 120 140 140 120 Initial Settling Rate m/hr 35 18 19 18 Settling Flux (t/m²h) 0.154 0.121 0.152 0.135 O/F Clarity g/m³ 1840 2075 5180 1990 Underflow Solids wt% 50 50 52 45 Yield Stress (Pa) 20 20 20 25 Hi-Density Thickeners Leach Feed (-106µm) Ore ID Zone 1 Zone 2 Year 1 Year 2 Year 3 Solids SG 2.7 2.7 2.7 2.7 Liquor SG 1.076 1.076 1.076 1.076 Optimum Solids wt% 10 10 10 6 Flocculant Dosage g/t 40 60 60 60 Initial Settling Rate m/hr 19.6 22 15 23.7 Settling Flux (t/m²h) 0.596 0.660 0.572 0.572 O/F Clarity g/m³ 1330 1195 4820 1045 Underflow Solids wt% 67 69 70 63 Yield Stress (Pa) 80 70 80 80 Leach Discharge (-106µm) Ore ID Zone 1 Zone 2 Year 1 Year 2 Year 3 Solids SG 2.7 2.7 2.7 2.7 Liquor SG 1.086 1.086 1.086 1.089 Optimum Solids wt% 3 4 4 3 Flocculant Dosage g/t 120 140 140 120 Initial Settling Rate m/hr 35 18 19 18 Settling Flux (t/m²h) 0.346 0.271 0.342 0.304 O/F Clarity g/m³ 1840 2075 5180 1990 Underflow Solids wt% 55 58 58 49 Yield Stress (Pa) 90 80 80 80
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Table 16.4.16_2 Husab Uranium Project Thickening Testwork Summary
Ore ID SDU UO 4 Zone 1 Repulp Solids SG 3.95 4.35 2.7 Liqour SG 1.17 1.039 1.09 Solids wt% As is As is 4 Flocculant Dosage g/t None None 100 Initial Settling Rate m/hr 0.19 0.02 15.6 Settling Flux (t/m²h) N/A N/A 0.176 O/F Clarity g/m³ 4195 N/A 3180 Underflow Solids wt% 62 (As received) 54 (As received) 56 Yield Stress (Pa) 80 30 80
The unit areas were calculated on the thickener having an equivalent mud bed of 1m for Hi-Rate and 3m for Hi-Density thickeners, and are inclusive of a 20% scale up factor.
The underflow solids wt% of SDU and UO 4 shown in Table 16.4.16_2 were received ‘as is’ from SGS personnel and the samples were tested for rheology by FLS.
The following conclusions and recommendations were made:
Sample Characterisation:
Solids specific gravity (SG) for Zone 1 and Zone 2 samples of 2,7t/m³ and was used for all calculations
All leach discharge samples tested were sampled from the pilot plant -106 m screen undersize slurry and were taken to be representative
Leach feed samples were wet-screened at -106 m
All samples tested were sent to SGS laboratories for laser sizing (fine particle size distribution).
Feed Dilution:
The optimal settling flux was found to be in the range of 6% to 10% feed well solids concentration for the leach feed samples and 3% to 4% feed well solids concentration for leach discharge samples
Flocculant Dosage:
Floeger SNF FA 920 SH flocculant (non-ionic) was selected as the best flocculant for the Husab samples, in terms of settling flux and overflow clarity, and was used for all thickening testwork
The optimal flocculant dosage ranged from 40g/t to 60g/t for leach feed samples and 120g/t to 140g/t for leach discharge samples
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Settling Rates:
Initial settling rates of 15m/hr to 23,7m/hr were achieved on leach feed samples
Initial settling rates of 18m/hr to 35m/hr were achieved on leach discharge samples
Overflow Clarity:
Good overflow clarities as visually compared to process barren liquor were achieved across the tests
Overflow samples for different tests were analysed by SGS
Underflow Solids Concentration:
Leach feed: underflow solids concentrations of 60 wt% to 64 wt% were achieved from the 4 L batch tests after two hours residence time
Leach discharge: underflow solids concentrations of 45% to 52% were achieved from the 4 L batch tests after two hours residence time
Thickener selection and operating window:
The thickening rates selected are based on the 4 litre settling tests results. 16.4.16 Larox
Larox conducted filtration tests on the four different feed materials being run through the pilot plant to determine:
Flocculant selection
Cake thickness
Maximum filtration capacity
Moisture content of the cake
cake handling
cake washing efficiency.
The filtration tests on the Husab uranium samples showed that the material can be successfully dewatered and washed using Larox filtration technologies.
The Husab uranium pilot plant was run over 2 campaigns each of 2 weeks (runs 2&3).
Each campaign had 2 different batches of feed material.
For run 2 they were Zone 2 (T41–46) and Zone 1 - Year 1 (T47-51);
For run 3 they were Zone 1 – Year 2 (T52-62) and Zone 1 - Year 3 (T63-66);
A similar series of tests were conducted for each of the 4 feed types.
All of the results are based on the samples as tested.
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Zone 2
The recombined products filtered well after washing, resulting in a filtration rate of about 577kg d.s./m²hr with a moisture content of around 17%, the resultant filter cake was 18mm and came off the cloth well. The flocculant dose (M351) was around 107g/t.
The barren wash ratio was 0.18m³/t, with a combined wash ratio of 0.45m³/t.
The barren wash efficiency at this ratio was 56%, with a total wash ratio of 93%.
Zone 1 – Year 1
The recombined products filtered well after washing, resulting in a filtration rate of about 1934kg d.s./m²hr with a moisture content of around 20.8%, the resultant filter cake was 39mm and came off the cloth well. The flocculant dose (M351) was around 169g/t.
The barren wash ratio was 0.47m³/t, with a combined wash ratio of 0.66m³/t.
The barren wash efficiency at this ratio was 76.1%, with a total wash ratio of 91.7%.
Zone 1 – Year 2
The recombined products filtered well after washing, resulting in a filtration rate of about 1538kg d.s./m²hr with a moisture content of around 20.2%, the resultant filter cake was 38mm and came off the cloth well. The flocculant dose (M351) was around 167g/t.
The barren wash ratio was 0.59m³/t, with a combined wash ratio of 0.78m³/t.
The barren wash efficiency at this ratio was 93.8%, with a total wash ratio of 94.7%.
Zone 1 – Year 3
The recombined products filtered well after washing, resulting in a filtration rate of about 1520kg d.s./m²hr with a moisture content of around 19.8%, the resultant filter cake was 38mm and came off the cloth well. The flocculant dose (M351) was around 167g/t.
The barren wash ratio was 0.50m³/t, with a combined wash ratio of 0.69m³/t.
The barren wash efficiency at this ratio was 87.2%, with a total wash ratio of 90.5%.
Note: Flocculant doses are only for flocculant added to the filter, and they do not include any flocculant added previously to the thickener.
16.4.17 Delkor
Delkor conducted laboratory scale Buchner funnel testwork was conducted at SGS labs in Perth during March and April 2010. Samples tested were taken from the Husab pilot operation at SGS. Tests were run on both unclassified leach residue and leach residue which had been classified into a coarse and fine fraction and recombined for filtration.
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It was found that the samples generally filtered rapidly when conditioned with flocculant as would be expected from the relatively coarse grind.
Flocculation
For all samples, filtration without the addition of flocculant was slow. The addition of flocculant produced a homogenous mixture of coarse and fines throughout the filter cake and resulted in rapid cake formation and cake washing rates. Delkor’s tests focussed on the use of Guar to flocculate the slurry based on historical data from acid leach uranium belt filter, costs per ton of ore treated and good cake washing characteristics;
Guar powder was hydrated to a strength of 3g/L, with optimum dosages on most samples varying between 170 and 220g/t. Consistently there was a ± 50g/t (25%) range of optimum dosage; above which, filtration rates declined;
Filtration Rates
Some variation in filtration rate was observed between the different samples. This correlated with the % of fines in the sample and had to be compensated by higher Guar dosage. Cake formation rates are between 7t/m²h and 8t/m²h.
Cake Washing Results
The filter cakes were washed with measured quantities of synthetic barren solution immediately after cake formation. The washed cakes were sampled, re-pulped with acidified water and analysed for uranium and Lithium in solution. The wash efficiencies were very close to theoretical models and consistent with wash results on similar acid leach uranium ores.
A factor 1 displacement wash (liquor content of cake before wash) produced an 80% removal of the original liquor. These results were consistent for both uranium and lithium analysis and for all samples of leach residue tested. Based on a one wash system, the soluble recoveries shown in Table 16.4.17_1 can be achieved.
Table 16.4.17_1 Soluble Recoveries - Factor 1 Displacement Wash
Wash Ratio (m³/ton) Displacements Soluble Recovery 0,33 1,0 80% 0,5 1,5 97,1% 0,83 2,5 99,1%
The single stage wash results were used to model a 2 stage counter-current case where the last wash (barren solution) is recycled to the 1 st wash of the belt filter. Based on a 2 stage counter- current wash system, the soluble recoveries shown in Table 16.4.17_2 can be achieved.
Wash rates varied significantly between the various samples from 1,9 to below 1,0m³/m²h in tests where the fines content was highest.
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Table 16.4.17_2 Soluble Recoveries - 2 Stage Counter-Current Wash
Wash Ratio (m³/ton) Displacements Soluble Recovery 0,37 1,15 97,4% 0,58 1,8 99%
Cake Drying Results
Residual cake moisture on most samples averaged 19% to 21% (w/w). In all cases the cake was competent and discharged easily from the filter cloth. There was no free moisture evident and the cake should be amenable to conveying over long distances.
Filter Sizing Calculation
Final filter sizing will depend on, amongst other factors, the wash ratio and wash configuration. Using a single wash of 1,5 displacements and based on the best test sample results achieved will result in:
Form Zone =23mm cake in 15 seconds
Wash Zone =1,5 IBV’s in 40 seconds
Drying Zone =15 seconds
Total Cycle Time =70 seconds
23mm Cake SG =33kg/m²
Sizing Flux =1700kg/m²h
16.4.18 RPA Filtres Philippe
Filtres Philippe performed filtration, washing and dewatering tests on leached uranium slurry of varying coarse and fines ratios, using a 50cm² Philippe vacuum funnel and a 3.2m³/h membrane vacuum pump.
The process conditions expected were set at:
Coarse / fines ratio =variable
Nominal output =2000t/h
Concentration =65 to70% solid
Density =1,6 SG
Temperature =18°C (room temperature)
Process temperature =40°C
Washing =Barren and tap water
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The objective is to produce:
Cake moisture as low as possible
U3O8 wash recovery >99%
The following recommendations were made:
Filter cloth recommended is either calendared, 85 micron double layered weave or a 6185 PES monofilament in either polypropylene or polyester. These clothes are expected to last for 6-7 months.
The filtrate quality is expected to be less than 1g/L of solids.
The two preferred flocculants are SNF FA 920 SH and SNF FA 912 BPM, to be prepared at 5g/L, matured for 4 hours and dosed at a maximum rate of 180g/t solids. The flocculant is to be diluted to between 0.125 and 0.25g/L and injected either in multiple in line points or using a series of flocculator tanks to control the contact time.
Belt speed to be set at 33m/min.
A cake thickness of 25mm is to be used for leach slurry. A thickness of between 25mm and 30mm is to be used for the blended fines and coarse material.
A wash ratio of 0.5m³ to 1.2m³/ton dry solid has been used, but no specific number is given.
A cake moisture between 17 and 20% is achievable after 15 to 20 s of drying time.
An average filterability of between 1600 and 1700kg/h/m² is achievable.
16.5 Comminution Pilot Plant Testwork
16.5.1 Introduction
Mintek conducted three semi-autogenous (SAG) pilot milling tests as part of the study to determine which circuit will be best suited for the optimal processing of the material (Mokwena et al, 2010). The circuit to be designed will be required to produce a leach feed material with
a grind of 80% passing 780 microns. Previous leach tests indicated that the U 3O8 extraction in excess of 90% can be achieved at this grind.
The amount of fines in the milling circuit product, as indicated by the mass percentage of -106 micron material, should not exceed 30% as this would present downstream materials handling problems.
The pilot plant tests conducted on the material included:
SAG mill in closed circuit with a pebble crusher and rod mill (SRC circuit).
SAG mill in closed circuit with a pebble crusher (SSSAG circuit).
SAG mill in closed circuit with a pebble crusher followed by a ball mill (SABC circuit).
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Bench scale comminution tests were also conducted to characterise the ore and support the pilot plant data generated. The tests conducted included SAG mill comminution (SMC), bond ball mill and rod mill grindability, bond crushability, uni-axial compressive strength (UCS), bond abrasion, grind mill and the JKTech abrasion test for self-breakage tests.
16.5.2 Sample Preparation
Approximately 140t of ore was delivered to Mintek for the intended testwork program. The P 80 of the sample as received was 212mm since the ore delivered contains 25t of material coarser than 212mm. The coarse material (+212mm) were broken with an excavator to 100% passing
212mm and the resulting P 80 of the ore was 150mm. Sample preparation was conducted on the ore to prepare a feed for the three piloting tests and a sub sample was taken to conduct comminution bench scale testwork in order to characterise the Husab uranium ore and support data generated during pilot tests.
The Husab uranium ore delivered to Mintek appeared to consist of two distinctive ore types. The difference between the two ore types was identified in texture and hardness and was classified as sample 1 and sample 2 or rock type 1 and rock type 2. Bench scale comminution testwork was conducted on both samples separately to determine if there is a significant difference in hardness between the two rock types. A visual observation of the bulk ore received indicates that 80% of the material received constitutes rock type 1 and 20% of rock type 2.
16.5.3 Summary of Results
The bond ball work index (BBWI) tests were conducted on a composite sample at 300 m and 850 m closing screens. The test results showed that the work index was slightly sensitive to the choice of the limiting screen size used. The BBWI tests results showed that the composite sample could be classified as being medium hard with the work indices of 13.50kWh/t and 12.78kWh/t tested at 300 m and 850 m respectively.
The bond rod work index (BRWI) conducted at a limiting screen of 1.18mm indicated that the Husab composite sample could be classified as being medium hard with the work index of 13.06kWh/, which is consistent with the BBWI data.
The average bond crushability work indices for sample 1 and sample 2 were 8.6kWh/t and 8.8kW/t (a 2% difference), respectively indicating that the samples could be classified as being very soft.
The average bond abrasion index values implied that both rock types are characterised by medium abrasion tendency with average values of 0.3285 for sample 1 and 0.2788 for sample 2.
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An average of UCS results for sample 1 and sample 2 were found to be 131.7MPa and 149.2MPa respectively indicating the material is categorised in the medium hardness class of UCS values. Although UCS values cannot be used directly for mill or crusher circuit design, when coupled with a database, they can sometimes serve as a diagnostic tool in identifying anomalous breakage behaviour.
Raw SMC data were sent to JKTech in Australia for analysis. The test was conducted on both samples using two different size classes namely -31.5+26.5mm and -22.4+19mm.
The product of the JKTech impact breakage parameters A and b was used to indicate the relative hardness (resistance to impact breakage) of the two samples. Sample 2 displayed more resistance to breakage at the coarser size with an Axb parameter of 56 while sample 1 was relatively less resistant to impact breakage with a Ax b parameter of 94.
The JKTech abrasion test results showed that sample 2 is slightly more resistant to abrasion
breakage than sample 1 with a t a value of 0.43 and 0.58 respectively.
The single stage SAG mill in closed circuit with pebble crusher was tested first. The SSSAG pilot plant campaign revealed that the amount of energy that will be required by the SAG mill to produce the desired grind would be 5.11kWh/t. This was achieved with a SAG mill running at 30% full with a ball charge of 15%. The total circulating load was only 93%. The material did not produce a significant amount of pebbles; only 12% (relative to the fresh feed) of pebbles were generated implying that the material is relatively soft.
16.5.4 Conclusion
From the three pilot plant testwork conducted, the single stage SAG mill with pebble crusher offers the best perspective due to:
the simplicity of the circuit;
the reduced quantity of equipment; and
the possibility of further expansion if required (conversion to SABC).
The SSSAG circuit results indicates that 5.11 kWh/t of energy would be required by the SAG
mill to obtain the desired grind (P 80 of 780 microns) with the correct proportion of fines (percentage passing -106 m) of 25% and a re-circulating load ratio of only 93%. The SRC circuit generates less proportion of fines (percentage passing -106 m) 20% in comparison to 25% for the SSSAG.
16.6 Development of Flowsheet Based on Testwork
The flowsheet used in the definitive feasibility study (DFS) is primarily derived from the flowsheet used in the hydrometallurgical pilot plant as described in Section 16.4 above, the Mintek comminution pilot plant and the findings from a PLS option study recommending the adoption of the strong acid strip process. There are three main differences to that flowsheet that are incorporated into the DFS flowsheet:
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The inclusion of a mineral sizer in the comminution circuit.
The replacement of the leach discharge screen and fines thickener with a coarse filtration and fines CCD wash circuit
The replacement of the SX carbonate strip process with the strong acid strip process.
The comminution circuit consists of a crusher with stockpile and two trains of closed circuit single stage SAG mill and pebble crusher. (SSSAG circuit). This circuit has been tested and compared with other configurations and is the circuit recommended by the comminution pilot plant testwork report described in Section 16.4 above. The choice of the SSSAG circuit is driven by the simplicity of the circuit, the reduced quantity of equipment and therefore the relative ease of further expansion or addition of more equipment if required, while still maintaining the required quality of grind product as required for leaching and subsequent process unit operations.
The solid/liquid separation and wash in the hydrometallurgical pilot plant was simulated using a screen to separate fine from coarse material, then thickening the fines and separating the PLS. The thickened fines material was flocculated, mixed with the screened coarse material, flocculated again and filtered. The filter cake was washed and washate was sent to the PLS. This configuration was not a consideration for a full scale plant, but simply a means of separating the PLS for downstream pilot plant testwork. The solids served as feed to vendor filtration and settling testwork. Filter tests have shown that considerable risk exists if either insufficient fines (incomplete washing and uranium losses), or excessive fines (severely reduced filtration rates) are present in the feed to the filters. This information has resulted in a conservative approach to the filtration plant design.
In order to create a feed material suitable for filtration, a fines fraction is separated by cyclone such that the amount of fines in the cyclone underflow is suitable for the filtration requirements. The excess fines are then treated separately in a CCD wash circuit, filtered by disc filtration and the relatively dry cake from both the belt filter plant and disk filter plant is conveyed to the tailings storage facility. In this configuration the need for minimising water losses to tailings (cost and environmental pressures), minimising uranium losses to tails while maintaining the filter feed quality in an operable regime are achieved.
The carbonate strip circuit was shown to be very expensive in terms of capital and operating cost. The proven strong acid strip process was adopted for the DFS due to its simplicity of design and operation and the significantly reduced operating costs. The process entails stripping the loaded organic with a strong acid strip solution. The excess sulphate contained in the loaded strip liquor which would otherwise inhibit the peroxide precipitation, is removed by precipitating gypsum using a high quality slaked lime. The gypsum precipitation process replaces the previous intermediate SDU precipitation and re-dissolution process. The final product remains uranyl peroxide.
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16.7 Key Technical Features
Mineral sizer (product top size 250mm) for primary crushing of the run-of-mine (ROM)
ore followed by semi-autogenous (SAG) milling to produce a P 80 of 780µm;
Atmospheric leach process with 14 hour residence time. Importing of sulphuric acid and pyrolusite as lixiviant and oxidant respectively;
Solid / liquid separation using belt filtration with option of counter current decantation for excessive fines. Leach residue deposited as filter cake for minimising water requirement;
Continuous ion-exchange using the NIMCIX technology and conventional solvent extraction (SX) uranium upgrading and refining process;
Production of uranyl peroxide by precipitation with hydrogen peroxide following a strong acid strip process.
16.8 Process Plant
The DFS flowsheet is shown in Figure 16.8_1. The process is best described by separating the process plant into areas.
16.8.1 Crushing
The ROM ore is received in a ROM bin which feeds via an apron feeder on to a mineral sizer. The crusher ore is transported to the coarse ore stockpile which has a twenty four hour live capacity. The coarse ore is drawn from the coarse ore stockpile onto two mill feed conveyors using two apron feeders per conveyor. Steel balls are added onto the mill feed conveyors.
16.8.2 Milling
The mill feed conveyors feed two semi-autogenous (SAG) mills. The milled slurry discharges from the mills over a screen. The oversize from the screen is returned to the mill via a pebble crusher. The undersize of the mill discharge screens advance to the fines screens. The coarser oversize fraction returns to the SAG mills for further milling. The finer underflow fraction advances to the leach circuit.
16.8.3 Leach
The milled slurry passes through a cyclone circuit to separate the coarse and fine fractions. The fine overflow fraction is thickened in a leach feed thickener. The thickened underflow slurry is recombined with the coarse cyclone underflow to achieve a slurry that is at an optimal relative density. The slurry then feeds into the leach circuit consisting of two trains of six leach reactors. Pyrolusite is added to the leach reactors as an oxidant to enable leaching and sulphuric acid, delivered to site, is added to maintain the uranium in solution. The leached slurry advances to the solid liquid separation circuit.
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Figure 16.8_1
DFS Flowsheet
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16.8.4 Solid-l Liquid Separation
The leached slurry passes through a cyclone circuit to separate the slurry into coarse and fine fractions. The coarse fraction advances to the vacuum belt filters (VBFs) where the filtrate is collected and recycled back to the fines treatment circuit. The fines fraction advances to the five stage counter current decantation (CCD) circuit. The recycled filtrate from the VBF is used as wash water in the CCD circuit. The liquid from the CCD circuit, now termed the pregnant leach solution (PLS), is clarified to remove suspended solids before advancing to the ion exchange circuit.
16.8.5 Ion Exchange
The PLS passes through eight parallel NIMCIX adsorption columns. The uranium is adsorbed onto the ion exchange resin, largely excluding contaminants. The loaded resin is transferred to elution columns where the resin is contacted with diluted sulphuric acid to extract the uranium from the resin. The eluate, which is now a much smaller volume than the PLS and contains a greater concentration of uranium, advances to the solvent extraction circuit for further concentration and purification. The eluate advances through an eluate filter to the solvent extraction plant.
16.8.6 Solvent Extraction
The eluate is contacted with organic in a four stage counter current extraction mixer settler circuit. Extraction raffinate is recycled back to the IX eluate tank, with a bleed stream going to leach. The loaded organic from the extraction circuit is scrubbed in a two stage counter current scrub mixer settler circuit using recycled uranyl peroxide barren liquor, supplemented with demineralised water and sodium hydroxide. The scrubbed organic is stripped in a five stage counter current strip mixer settler circuit using dilute sulphuric acid as a strip liquor. The loaded strip liquor advances to the precipitation circuit.
16.8.7 Precipitation
The loaded strip liquor is reacted with slaked lime in a four stage gypsum precipitation circuit. The gypsum slurry is thickened in a thickener. The thickener underflow advances to the gypsum belt filter where the slurry is washed, dewatered, repulped and returned to the leach circuit. A reseed stream from the thickener underflow reports back to the gypsum precipitation reactors. The thickener overflow advances to the uranyl peroxide circuit, which consists of four reactors in series. Peroxide is added to the reactors to cause the uranyl peroxide to precipitate out. Sodium hydroxide is added to the reactors to control the pH of the solution.
The uranyl peroxide slurry reports to a thickener. A reseed stream from the thickener underflow returns to the reactors. The remaining thickener underflow is washed and dewatered in a two stage centrifuge circuit. The dewatered uranyl peroxide is then dried and packaged, ready for shipping.
16.8.8 Tailings
Plant tailings are conveyed to the mine residue facility and disposed of alongside waste rock from the mine.
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17 MINERAL RESOURCE ESTIMATES
The August 2010 resource update was based upon work conducted by Neal Culpan of Extract, who completed 3D models of the geology and mineralisation of the Husab Uranium Project deposit. The supplied interpretations were then checked and verified by Coffey Mining and used in the estimation process with only generally minor changes made.
Two separate models were initially created for the Husab Uranium Project Mineralisation; Zone 1 was modelled using Surpac software; and Zones 2 to 4 were modelled using Datamine software. Subsequent to the completion of each model, the resulting block models were then combined into a single global Datamine block model for reporting and study purposes.
The Qualified Persons responsible for the resource estimate are Steve Le Brun, who is an Principal Resource Geologist with the consultancy Coffey Mining Pty Ltd. and Neil Inwood, who is a Principal Resource Geologist with the consultancy Coffey Mining. The details of the resource estimates undertaken are summarised in the following sections.
17.1 Zone 1 Resource Estimate
17.1.1 Resource Database and Validation
Database
The drillhole database in the vicinity of Zone 1 contains a total of 487 resource holes drilled by Extract between 2007 and 2010 (Figure 17.1.1_1). Early regional exploration holes were excluded from the database. The majority of the drillholes were drilled angled at 60° towards west (WGS84_33S grid). The database contains 95 diamond holes for 37,350m and 392 RC holes for 111,230m.
Both radiometric (128) and chemical (118,713) U3O8 assays were available in the database. Assay results from several drillholes were still pending at the time of the estimate and are intended to be incorporated in the next update. Approximately 25,855 individual samples from 2003 individual mineralised intersections were used directly in the resource estimate.
The drillholes in the database have DGPS collar pick up surveys accurate to ±10cm relative to the survey base station. All but a few 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.
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Validation
Prior to loading data into the database the following checks are carried out by Extract:
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. are not 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.
The following additional checks were also undertaken prior to estimation:
Downhole survey trends.
Overlapping intervals.
Missing intervals.
Checks of the top 1,000 assay intervals to the original laboratory files.
3D analysis of collar positions and downhole survey traces.
No significant validation errors were detected in the database and the database was considered appropriate for the use in the resource estimation.
17.1.2 Geological Interpretation and Modelling
The geological sequence at Zone 1 is a south-plunging antiform with gneisses of the Khan Formation overlain by a mixture of metasediments and altered calc-silicate lithologies of the Rössing Formation which is overlain by biotite schists and gneisses of the Chuos Formation. Alaskite bodies have generally intruded sub-parallel to foliation. This is observed both in core and in outcrop at the northern end of Zone 1. Separate DTM models for the main lithological units were generated by Extract for use in coding of the block model and for allocating density.
Modelling of the Zone 1 mineralisation consisted of an initial geological model of the broad lithological units (Khan, Rössing, Chuos). An alaskite model was then created using the underlying geological framework as a guide. Finally, a nominal +75ppm mineralisation wireframe was constructed that used these geological models to help guide the orientation of the mineralized outlines (Figures 17.1.1_1 and 17.1.2_1).
The downhole thickness of mineralised intervals ranged from 3 to 145m, with an average thickness of 23m. A total of 42 mineralised zones (grouped into 39 domains) were created that had strike extents ranging from 125m to 1100m. Due to the geometries of the mineralisation, the true thickness of the mineralisation ranges from 80% to 100% of the downhole thickness.
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Figure 17.1.1_1 Husab Uranium Project Zone 1 - Mineralised Zones and Drill Type
(Red – Diamond, Blue – RC)
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Figure 17.1.2_1 Husab Uranium Project, Zone 1 Sectional Interpretation (7,506,800mN)
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Alaskite Bodies
Due to the complexity of the alaskite lithologies exhibited in the drilling sections, it was decided to use an Indicator based probability model to generate an alaskite model. The alaskite indicator model was generated utilising Vulcan mining software and then imported into the Surpac Block model. For the purposes of the resource model, blocks having probabilities of >50% were coded as alaskite in the waste regions, and blocks having probabilities of >30% were coded as alaskite within the modelled mineralisation. This coding regime resulted in the proportions of mineralisation being coded to the main lithological units (e.g. Khan, Chuos, Alaskite, calc-silicate, etc) within the block mode being similar to that observed in the drillhole database.
Surface Cover and Weathering Profile
Alluvium covers most of the Husab Uranium Project. The depth of this alluvial cover tends to deepen to the south and towards the east. The thickness of alluvial cover at the south end of Zone 1 is up to 75m. The alluvium is comprised of an upper layer, typically about 20m thick, of loose sand and gravel underlain by carbonate cemented gravel/conglomerate. A model was created for the base of alluvial cover.
17.1.3 Radiometric Data
Factored downhole spectrometer eU 3O8 data from hole RDD130 was used in the resource as
assay results were pending. The raw eU 3O8 data was factored for the purposes of the estimate using the regression defined by Culpan (2008), being: Fact_eU3O8 = [eu3o8_raw]*0.73-45.
17.1.4 Statistical Analysis of Composites and Top Cuts
The data captured within the mineralisation wireframes was composited to a regular 3m downhole composite length, residual intervals of less than 1.5m were discarded. Table 17.1.4_1 summarises the composite statistics for the various zones within the deposit.
A statistical analysis was carried out on the composited data for each unit to determine appropriate top cuts to apply to the data. The approach taken included:
Review of the 3D grade distribution;
Review of the histogram and probability plots with significant breaks in populations used to identify possible outliers;
Ranking of the individual composites and investigating the affect of the higher grades upon the standard deviation and the mean of the data population.
The resulting top cuts applied (Table 17.1.4_1) resulted in a decrease of the naïve mean for the individual mineralised zones of between 1% and 23%, typically from the cutting of up to 9 composites. The mineralised zones which exhibited the largest changes in mean grade (e.g. Zones 2 and 34) were characterised by a strong positive distribution with a relatively few
number of high grade composites (e.g. 3 composites >2,000ppm U3O8 for Zone 2) supplying up to 13% of the contained relative metal for the individual zones. Figure 17.1.4_2 shows type examples of the graphs that were examined from each mineralised zone to assess the top cutting of outlier assays.
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Figure 17.1.4_2 Histogram Plot from Mineralised Zones 2 to 4
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Table 17.1.4_1 Husab Uranium Project Zone 1 Summary Statistics and Top Cuts Applied to the Various Mineralised Zones
Top Cut % # Lode Number Min. Max. Mean Median Data Variance C.V. Cut Mean Decrease Cut 1 160 14 2,649 308 149 407 165,796 1.32 2000 304 1 1 2 807 3 24,086 637 176 1,834 3,364,701 2.88 4400 503 21 20 3 775 2 11,330 426 173 783 612,718 1.84 3500 402 5 9 4 369 3 13,915 547 213 1,168 1,363,919 2.14 3,500 475 13 8 5 358 3 6,304 406 177 657 431,808 1.62 3,000 391 4 5 6 258 6 20,335 622 202 1,558 2,428,602 2.51 4,000 527 15 4 7 198 5 4,903 625 196 948 899,322 1.52 3,500 602 4 5 8 291 10 3,143 252 137 355 126,128 1.41 2,000 246 3 3 9 69 20 883 192 161 136 18,448 0.71 192
10 14 12 768 251 180 250 62,645 1.00 251
11 37 5 709 214 174 165 27,064 0.77 214
12 41 14 1,940 312 142 434 188,298 1.39 312
13 104 24 5,804 408 170 812 659,865 1.99 2,500 351 14 3 14 367 9 4,901 433 202 602 361,806 1.39 3,000 424 2 5 15 455 3 5,320 427 190 618 382,397 1.45 3,000 418 2 5 16 547 3 5,534 461 236 651 423,965 1.41 3,000 448 3 8 17 187 6 4,754 555 304 738 545,079 1.33 3,000 534 4 5 18 122 7 2,779 446 236 522 272,137 1.17 446
19 159 4 1,900 398 276 366 133,941 0.92 398
20 73 6 1,228 307 203 286 81,892 0.93 307
21 59 8 601 169 109 153 23,547 0.91 169
22 141 6 4,192 360 125 734 539,153 2.04 2,500 320 11 5 23 29 23 573 186 104 166 27,393 0.89 186
24 73 10 3,942 384 91 754 568,365 1.97 2,500 355 7 3 25 95 5 4,438 433 162 741 548,604 1.71 2,500 393 9 2 26 188 2 3,167 386 171 542 294,218 1.40 2,500 383 1 1 27 148 5 3,679 347 132 546 298,292 1.57 2,500 332 4 2 28 400 2 4,564 567 314 715 511,845 1.26 3,500 561 1 2 29 92 16 3,520 321 144 528 278,261 1.64 2,500 310 3 1 30 190 4 11,405 784 363 1,324 1,752,125 1.69 4,000 712 9 4 31 39 8 2,232 320 167 455 206,782 1.42 320
32 127 4 3,898 454 293 559 312,481 1.23 2,500 443 2
33 97 25 1,941 354 165 403 162,099 1.14 354
34 40 10 8,453 692 142 1,489 2,216,335 2.15 3,000 533 23 2 35 288 4 8,902 836 426 1,143 1,305,498 1.37 4,500 798 5 5 36 28 58 3,798 621 173 924 853,024 1.49 621
37 168 4 3,389 343 143 502 251,719 1.46 2,000 331 4 3 38 43 5 571 156 113 131 17,223 0.84 156
39 86 8 1,628 217 103 294 86,637 1.36 217
17.1.5 Bulk Density Data
Density readings taken from water immersion of whole core have been taken for 90 diamond holes drilled in the Husab Uranium Project area. After removal of obvious outlier data (i.e. values <1.5t/m³ and >3.5t/m³), 3,338 reading were grouped by lithology type and an average bulk density readings were determined. Table 17.1.5_1 summarises the density readings.
The site density procedures have recently been reviewed and revised recording practices have been suggested to mitigate future recording errors.
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Table 17.1.5_1 Density Readings taken from Drill Core at Husab Uranium Project (Filtered by >1.5tm³ and <3.5tm³)
Biotite Calc- Min Item Gneiss Alaskite Sediments Marble Cover Schist Silicates Zones Count 966 379 1,192 281 410 28 82 866 Minimum 1.55 1.61 1.58 2.07 1.53 2.63 1.85 1.62 Maximum 3.5 3.48 3.48 3.46 3.45 3.24 2.61 3.47 Mean 2.73 2.73 2.69 2.76 2.78 2.86 2.18 2.70 Median 2.71 2.72 2.65 2.7 2.77 2.81 2.67 Standard Deviation 0.23 0.23 0.21 0.2 0.31 0.19 0.19 Variance 0.05 0.05 0.05 0.04 0.1 0.04 0.04 Coefficient of Variation 0.09 0.08 0.08 0.07 0.11 0.07 0.07
17.1.6 Variography
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. Isatis geostatistical software was used throughout. Both traditional semi-variograms and
correlograms were used to analyse the spatial variability of the U 3O8 3m composites for the mineralised zones. Downhole variography was calculated and considered when determining the nugget for each of the zones. Variography was undertaken on individual zones and upon grouped zones to determine the optimal variograms for use in the Ordinary Kriging estimation process. The domains used for the variography are shown in Figure 17.1.6_1 and the resulting variograms are shown in Figure 17.1.6_2.
After assessing the variography of individual mineralised zones, then of grouped zones; it was decided to use the variography derived from 5 major zone groupings for the subsequent OK estimate. The resulting variography generally showed reasonably structured variography in the major directions, but poorly structured variography in the semi major directions; often with similar total ranges to the major directions. As no robust anisotropy could be determined in the plane of mineralisation, and after reviewing the omnidirectional variograms for the zones, it was decide to use similar ranges in the major and semi-major directions. The resulting variography is shown in Table 17.1.6_1. For the estimate, the variography was orientated according to the main geometry of the modelled mineralised zone and are shown in Table 17.1.6_1.
Table 17.1.6_1 Grouped Zone Variography – Husab Uranium Project Zone 1 (Relative variances shown)
Major Semi-Major Minor Major Semi-Major Minor Zone Grouping Nugget C1 C2 Range Range Range Range Range Range Main Zone 25% 60% 28 28 15 15% 130 130 70 North 25% 50% 30 30 8 25% 120 120 45 South Major & East 25% 47% 30 30 8 28% 88 88 20 West 25% 60% 55 55 25 15% 210 220 84 South Minor 25% 50% 20 20 10 25% 94 94 45
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Figure 17.1.6_1 Grouped Domains – Husab Uranium Project Zone 1
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Figure 17.1.6_2 Modelled Variography – Husab Uranium Project Zone 1 (Variogram Directions, Clockwise from Top Left - Major, Semi-Major, Minor) Main North
West (Major and Semi-Major Directions) South Main South Minor (Omnidirectional)
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17.1.7 Block Model
A block model was created in the national grid (WGS84_33S) using Surpac mining software. The model extents, block sizes and attributes are summarized in Table 17.1.7_1 below.
Table 17.1.7_1 Block Model Parameters – Husab Uranium Project Zone 1
Type Y X Z Minimum Coordinates 7504500 505500 0 Maximum Coordinates 7507500 507500 600 User Block Size 25 25 15 Minimum Block Size 6.25 6.25 3.75 Rotation 0.000 0.000 0.000
Attribute Name Type Decimals Background Description avdis Real 12 -99 Average distance to samples category Integer - 0 Classification (1=meas; 2=ind; 3=inf, 4=unclass) density Real 2 2.75 Insitu Dry Bulk Density zone Integer - 0 Numerical zone number (1-39), air = 0, waste = 99 kvkrigvar Real 3 -99 Kriging variance lithgeol Character - Rock types- Rössing, alaskite, Chuos etc. Lith_ind real 3 0 Indicator for alaskite nsamps_okcut Integer - -99 Number of samples estflag_okcut Integer - 0 Estimation flag (1,2,3) u3o8_cut Real 2 -99 Cut uranium grade unitsor IntegerReal 2 0 Slope of regression
The following DTMs and 3DMs used in the construction of the model:
10m_july10_model_topo1.dtm - Topography DTM;
zn1_min_naijuly8.dtm - U3O8 Mineralisation wireframe;
chuos_Rössing_khan_faults100719.dtm - definition of the Chuos, Rössing and Khan Formation lithologies;
karibib-chuos1007.dtm - definition of the Karibib Formation lithologies;
bob1007.dtm - Base of overburden 3D model defining the base of the cemented conglomerate and sand lithologies;
gneiss090603_zn1.dtm - Gneiss 3D model;
calsil_zn1-4_1007.dtm - Cals-silicate unit;
karibib-chuos1007.dtm - Definition of the Karibib and Chuos Formation lithologies;
greygran1007.dtm - eastern granite .
17.1.8 Grade Estimation
Ordinary Kriging (OK) Estimate
U3O8 grade was estimated into the block model using OK after top cuts had been applied to the original 3m composite data. Neighbourhood analysis was used to optimise the search distance and number of informing samples for the first pass.
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Table 17.1.8_1 summarises the estimation parameters for each of the modelled zones.
Table 17.1.8_1 Husab Uranium Project - Zone 1 Sample Search Parameters for OK Estimate
Search/Variography Orientation Domain Zones Major (°) Semi-Major (°) Minor (°) 1,3 00 210 60 120 30 300 North 2 00 220 60 130 30 310 4 00 210 50 120 40 300 5, 10, 22 00 210 60 120 30 300 6,7,9 00 210 50 120 40 300 East 8, 11, 38 00 180 50 090 40 270 23 00 190 40 100 50 280 31 ,32, 33 00 200 60 110 30 290 35 00 200 50 110 40 290 South Minor 34, 36 00 210 60 120 30 300 37 00 180 80 090 10 270 South Major 19, 20, 21 00 210 60 120 30 300 12, 15, 17, 18 00 210 60 120 30 300 Main 13, 16 00 200 60 110 30 290 14 00 200 50 110 40 290 24 00 190 40 100 50 280 25 00 210 50 120 40 300 West 26, 27, 30, 37 00 180 80 090 10 270 28, 39 00 180 50 090 40 270 Search Parameters Major Axis Semi-Major Axis Minor Axis Min Max. Pass Max/hole (m) (m) (m) Samples Samples 1 80 80 26.7 14 24 6 2 160 160 534 14 24 6 3 240 240 80 6 12 6
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:
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.
Comparison of average informing composite grade and average block model grade (see Table 17.1.8_2 below).
Comparison plots of average informing composite grade and average block model grade by mineralized unit (Figure 17.1.8_1).
Visual inspection of estimated blocks against the informing composites and original drillhole data.
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Table 17.1.8_2
Comparison of Model Grades and Informing Composite U 3O8 Grades for Mineralised Units
Zone Block Grade Naïve Grade Declustered Grade B.M. / Naïve Mean B.M. / Declustered Mean 1 294 308 319 5% 8% 2 437 503 438 15% 0% 3 382 412 390 8% 2% 4 465 492 442 6% -5% 5 392 401 381 2% -3% 6 513 533 549 4% 7% 7 591 621 663 5% 12% 8 235 252 262 7% 11% 9 185 192 203 4% 10% 10 286 251 281 -12% -2% 11 210 214 201 2% -5% 12 392 312 315 -20% -20% 13 310 395 355 27% 14% 14 391 432 420 10% 7% 15 481 425 442 -12% -8% 16 448 457 471 2% 5% 17 573 553 527 -4% -8% 18 419 446 428 6% 2% 19 436 398 402 -9% -8% 20 319 307 301 -4% -6% 21 167 169 168 1% 0% 22 311 360 383 16% 23% 23 171 186 207 9% 21% 24 474 384 346 -19% -27% 25 367 433 357 18% -3% 26 417 386 368 -7% -12% 27 381 347 371 -9% -3% 28 653 566 553 -13% -15% 29 322 321 318 0% -1% 30 660 719 711 9% 8% 31 371 320 298 -14% -20% 32 436 454 463 4% 6% 33 369 354 346 -4% -6% 34 561 591 630 5% 12% 35 804 796 756 -1% -6% 36 546 621 608 14% 11% 37 299 343 327 15% 9% 38 163 156 171 -4% 5% 39 212 217 219 2% 3%
Overall, there was a good comparison between the informing composite data input into the model and the resulting block grades. Several zones were identified with block grades which deviated significantly from the sample mean (e.g. Zone 12). These occurrences were typically found to be due to the influence of higher grade composites in the smaller zones which had only a few numbers of samples (e.g. 41 samples for Zone 12 and 73 for Zone 24). Areas of significant variance within the block model were also dealt with in the block model by limiting the amount of Inferred classification.
Density
The densities used in the resource model (Table 17.1.8_3) were based on the analysis reported on in Section 17.1.5, with some minor adjustments to take into account potential errors in the density data.
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Figure 17.1.8_1 Example Northing Validation Plots – Husab Uranium Project Zone 1
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Table 17.1.8_3 Husab Uranium Project - Zone 1 Density Values Applied to the Various Rock-Types within the Resource Model
Lithology Density t/m³ Rössing Fm* 2.75 Chuos 2.75 Karibib 2.73 Khan 2.73 Calc-silicates 2.78 Alaskite 2.65 Cover** 2.18 Unknown/uncoded geology 2.70 * Rössing Formation is a combination of sediments, biotite schist and gneiss lithologies ** Based upon combined sand and cover materials
17.1.9 Resource Reporting and Classification
The resource estimate for Husab Uranium Project Zone 1 update has been classified in accordance with the criteria laid out in the JORC Code (2004) and NI43-101. Indicated and Inferred Mineral 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.
The classification of the Zone 1 resource was based on the confidence levels placed in the key criteria listed in Table 17.1.9_1 below. Figure 17.1.9_1 illustrates the classification applied to the resource model.
Figure 17.1.9_1 Oblique View of the Classified Zone 1 Resource Model and Drillholes
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Table 17.1.9_1 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 majority of drillholes have been High Points surveyed downhole for deviation Data Density and Ranging from 50m x 50m to 100m x 100m. Low to High 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 geological Moderate 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 of N/A Assumptions support) has not been investigated.
The resource estimate for the Husab Uranium Project Zone 1 is reported below at a range of cutoff grades.
Table 17.1.9_2 Husab Uranium Project Zone 1 - August 6 2010 Resource Estimate
Reported at various cutoffs, Preferred cutoff : 100ppm U 3O8 Ordinary Kriged Estimate based upon 3m cut U 3O8 Composites Parent Cell Dimensions of 25m NS by 25m EW by 15m RL
Lower Cutoff Tonnage Grade Contained U 3O8 Contained U 3O8 (ppm U 3O8) (Mt) (ppm U 3O8) (MKg) (MLb) Indicated 100 122.2 450 55.0 120.1 200 104.5 490 51.2 113.9 300 78.0 580 45.3 99.3 400 56.2 670 37.6 82.6 Inferred 100 41.3 420 17.4 37.8 200 34.2 470 16.1 35.2 300 24.7 550 13.6 29.9 400 17.0 640 10.9 24.1 Note: Figures have been rounded.
17.2 Zones 2 to 4 Resource Estimates
17.2.1 Resource Database and Validation
The database contains a combination of chemical assaying (138,518 samples – 91%), factored radiometric data (12,981 1m composites – 9%). Approximately 16,976 individual samples were used directly in the resource estimate, of which 185 were radiometric assays.
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Table 17.2.1_1 summarises the number, metres drilled and type of hole.
Table 17.2.1_1 Husab Uranium Project (Zone 2 - 4) Number, Metres Drilled and Type of Hole
Zone Hole Type No. Holes Metres Drilled RC 400 121,329 2 DDH 88 36,250 RC 48 16,460 3 DDH -- -- RC 57 17,267 4 DDH 3 1,060
The drillholes were typically drilled due west (WG84/33S grid) with a dip of -60°.
Validation checks of the database included:
A detailed analysis was undertaken comparing the radiometric data for matching intervals against the chemical assays.
During the course of the 3D modelling of the resource, the database was checked for any gross survey and position errors. The resulting database was considered to be robust and appropriate for use in resource estimation.
As part of the validation process, Coffey Mining conducted checks on the top 1,000 chemical assays for the Husab Uranium Project (all zones). Checks of these assays in the supplied database against laboratory certificates files indicate no significant data related issues with the database.
17.2.2 Geological Interpretation and Modelling
To establish appropriate grade continuity, the mineralisation model for the Husab Uranium
Project deposit was based upon a nominal 75ppm U3O8 mineralisation halo. The mineralisation constraints were generated based upon sectional interpretation and three dimensional analyses of the available drilling data. The main lithological contacts (e.g. alaskite and sediments) were considered at the time of modelling and used to guide modelling of mineralisation shapes. Unless a strong geological model could be established, mineralised zones which did not have more than two drillhole intersections on two consecutive sections were not estimated.
The Husab Uranium Project - Zone 2 region (Figures 17.2.2_1 to 17.2.2_5) was modelled as 29 distinct zones (3m to 82m thickness, averaging 19m) with a NE trend. Individual zones were modelled to extend for up to 1,200m along strike and between 100m to 600m down-dip. Due to the geometries of the mineralisation, the true thickness of the mineralisation ranges from 80% to 100% of the downhole thickness. The Husab Uranium Project – Zones 3 and 4 regions (Figures 17.2.2_1 and 17.2.2_3) were modelled as eight distinct zones (3m to 79m thickness, averaging 16m) with a NE trend. Individual zones were modelled to extend for up to 1,300m along strike and between 100m to 400m down-dip. Due to the geometries of the mineralisation, the true thickness of the mineralisation ranges from 80% to 100% of the downhole thickness.
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Figure 17.2.2_1 Husab Uranium Project (Zones 1 - 4) Drillhole Location Plan
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Figure 17.2.2_2 Husab Uranium Project Zone 2 Mineralised Zones and Drill Type
(Red – Diamond Drillholes, Blue – RC)
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Figure 17.2.2_3 Husab Uranium Project Zones 3 and 4 - Mineralised Zones and Drill Type
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Figure 17.2.2_4 Husab Uranium Project Zone 2 - Sectional Interpretation (7,503,600mN)
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Figure 17.2.2_5 Husab Uranium Project Zone 3 - Sectional Interpretation (7,501,600mN)
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Figures 17.2.2_4 and 17.2.2_5 show example sectional interpretations from Zones 2 and 3 respectively.
Alaskite Model
A probablistic alaskite model was coded to the Zones 2 to 4 portions of the block model in the same fashion as discussed in Section 17.1.2.
Sulphide Model
A sulphide model was coded to all the Zones in the block mode using the nearest neighbour method, with samples that had logged sulphides coded as 1 and all other samples coded as 0. This model was not classified and was used solely to aid in identifying potential areas of sulphidic material.
Surface Cover and Weathering Profile
The area of the resource is covered by up to 100m of mainly sandy superficial cover (Figure 17.2.2_4). Extract supplied a 3D model of the surface cover (Sand and conglomerate overburden) and main lithological units based upon the lithological logging of the RC and diamond drillholes. The model was reviewed by Coffey Mining and was used to code overburden into the resource block model.
Radiometric Data
Downhole radiometric data was used for portions of 29 drillholes. The raw eU 3O8 data was factored as outlined in Section 17.1.1.
17.2.3 Statistical Analysis of Composites and Top Cuts
The data captured within the mineralisation wireframes was composited to a regular 3m downhole composite length, with residual intervals of less than 1.5m retained. Table 17.2.3_1 summarises the composite statistics for the various zones within the deposit.
A statistical analysis was carried out on the composited data for each unit to determine appropriate top cuts to apply to the data. The approach taken included:
Review of the 3D grade distribution;
Review of the histogram and probability plots with significant breaks in populations used to identify possible outliers;
Ranking of the individual composites and investigating the affect of the higher grades upon the standard deviation and the mean of the data population.
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Table 17.2.3_1 Husab Uranium Project Zones 2 to 4 Summary Statistics and Top Cuts Applied to the Various Mineralised Zones
Top Cut % # Lode Number Min. Max. Mean Median Std. Dev. Variance C.V. Cut Mean Decrease Cut 201 29 10 1,305 287 134 369 136,333 1.29 6,000 287 0.0
202 22 80 2,232 251 116 452 204,437 1.8 6,000 251 0.0
203 240 10 9,695 598 190 1,026 1,052,497 1.72 6,000 583 2.6 1 204 270 6 7,395 440 150 872 760,976 1.98 6,000 435 1.2 1 205 131 8 2,246 333 149 465 215,860 1.4 6,000 333 0.0
206 207 8 32,865 500 168 2,325 5,407,907 4.65 6,000 370 25.9 1 207 46 13 1,708 332 149 374 139,539 1.13 6,000 332 0.0
208 24 5 752 281 175 246 60,395 0.88 6,000 281 0.0
209 26 27 3,328 598 395 725 525,875 1.21 6,000 598 0.0
210 477 5 6,367 624 346 816 665,706 1.31 6,000 623 0.2 3 211 568 5 3,928 479 246 621 385,298 1.3 6,000 479 0.0
212 1058 5 15,010 593 242 1,056 1,115,393 1.78 6,000 572 3.6 5 213 140 5 2,888 369 133 518 267,857 1.4 6,000 369 0.0
214 363 5 7,187 536 246 832 691,448 1.55 6,000 533 0.6 1 215 27 24 2,102 288 121 437 191,016 1.52 6,000 288 0.0
216 860 5 14,874 732 264 1,328 1,762,612 1.81 6,000 695 5.0 9 217 116 5 4,204 528 260 665 441,887 1.26 6,000 528 0.0
218 82 7 3,025 456 231 598 357,326 1.31 6,000 456 0.0
219 76 5 4,041 698 306 896 803,259 1.29 6,000 698 0.0
220 155 5 1,900 258 173 309 95,737 1.2 6,000 258 0.0
221 51 6 2,030 315 157 479 229,552 1.52 6,000 315 0.0
222 79 5 1,192 237 132 258 66,412 1.09 6,000 237 0.0
223 15 23 1,058 256 188 265 70,204 1.03 6,000 256 0.0
225 38 5 2,853 416 190 597 355,884 1.44 6,000 416 0.0
226 78 11 4,747 635 190 1,013 1,025,640 1.6 6,000 635 0.0
227 213 5 7,671 527 222 947 895,869 1.8 6,000 519 1.5 1 229 162 5 8,996 434 191 843 710,458 1.94 6,000 416 4.3 1 351 21 26 1,052 316 189 283 80,234 0.9 2,000 316 0.0
352 14 45 425 183 121 126 15,788 0.69 2,000 183 0.0
353 203 7 3,048 242 124 410 167,743 1.69 2,000 230 5.2 3 354 27 39 368 148 126 89 7,987 0.61 2,000 148 0.0
355 108 14 1,040 241 150 221 49,031 0.92 2,000 241 0.0
356 10 20 1,548 328 156 457 209,132 1.4 2,000 328 0.0
357 28 10 3,663 541 182 874 764,278 1.62 2,000 454 16.1 3 440 185 5 6,074 637 260 1,018 1,037,279 1.6 4,000 602 5.4 5
The resulting top cuts applied (Table 17.2.3_1) resulted in a decrease of the naïve mean from 1% to 26%, typically from the cutting of up to 9 composites. The mineralised zones which exhibited the largest changes in mean grade (e.g. zones 206 and 357) were characterised by a strong positive distribution with a relatively few number of high grade composites
(e.g. 1 composite >6,000ppm U3O8 for Zone 206) supplying up to 26% of the contained relative metal for the individual zones. Figure 17.1.1_3 show type examples of the graphs that were examined from each mineralised zone to assess the top cutting of outlier assays.
Bulk Density Data
Bulk density was applied based upon lithological domains (see Section 17.1.1) and is shown in Table 17.2.3_2.
Mineralised domains were assigned the bulk density of the underlying host lithology. Figure 17.2.3_1 shows a cross-section of the final model and the lithological coding used to assign bulk densities.
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Table 17.2.3_2 Husab Uranium Project (Zone 2) Bulk Density Values Assigned by Supplied Lithological Wireframes
Assigned “GEOL” Field Lithology Bulk Density (in Final Model) Sand / Overburden 2.18t/m³ OVB Karibib Formation 2.73t/m³ KARB Chuos Formation 2.75t/m³ CHUS Rössing Formation 2.75t/m³ ROSG Calc-Silicates 2.78t/m³ CALC Khan Formation 2.73t/m³ KHAN Alaskite 2.65t/m³ ALSK Granite/Gneiss 2.65t/m³ GRN
Figure 17.2.3_1 Cross-Section of the Zone 2 Block Model at 7,503,500mN Showing the Lithological Coding Used for Bulk Density Assignment
Variography
In most cases, individual lenses contained too few composites to generate meaningful individual correlograms. Due to limited data distribution in the down-dip orientation, semi-major correlograms were poor; therefore, it was decided to apply the parameters from the major axis in the semi-major orientation. The variography of the combined mineralised lenses (201 to 229) from ZONE 2 were used as the basis for the estimation for all three zones with the data from each of the mineralised zones as inputs for the OK estimation for each of the mineralised zone. The major and semi-major axes of the variogram and search axes were orientated to be parallel to the main trend of the individual modelled mineralisation domains.
Table 17.2.3_3 summarises the resulting variogram model parameters from Zone 2 and Table 17.2.3_4 summarises the resulting variogram model parameters from Zones 3 and 4. Figure 17.2.3_2 shows example variography from Zone 2.
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Table 17.2.3_3 Husab Uranium Project (Zone 2) Variogram Parameters for Mineralised Zone 2
Orientation Range 1 (m) Range 2 (m) Range 3 (m)
Major Semi-Major Minor C0 C1 C2 C3 Zone Semi- Semi- Semi- (%) (%) Major Minor (%) Major Minor (%) Major Minor Dip Azi Dip Azi Dip Azi Major Major Major (o) (o) (o) (o) (o) (o) 201 20 45 0 135 70 45 0.25 0.6 50 50 20 0.15 90 75 60 ------202 35 60 0 150 55 60 0.25 0.6 50 50 20 0.15 90 75 60 ------203 25 70 0 160 65 70 0.25 0.6 50 50 20 0.15 90 75 60 ------204 25 70 0 160 65 70 0.25 0.6 50 50 20 0.15 90 75 60 ------205 25 80 0 170 65 80 0.25 0.6 50 50 20 0.15 90 75 60 ------206 25 60 0 150 65 60 0.25 0.6 50 50 20 0.15 90 75 60 ------207 30 90 0 180 60 90 0.25 0.6 50 50 20 0.15 90 75 60 ------208 25 90 0 180 65 90 0.25 0.6 50 50 20 0.15 90 75 60 ------209 30 70 0 160 60 70 0.25 0.6 50 50 20 0.15 90 75 60 ------210 40 80 0 170 50 80 0.25 0.6 50 50 20 0.15 90 75 60 ------211 25 90 0 180 65 90 0.25 0.6 50 50 20 0.15 90 75 60 ------212 25 95 0 185 65 95 0.25 0.6 50 50 20 0.15 90 75 60 ------213 25 90 0 180 65 90 0.25 0.6 50 50 20 0.15 90 75 60 ------214 10 35 0 125 80 35 0.25 0.6 50 50 20 0.15 90 75 60 ------215 25 90 0 180 65 90 0.25 0.6 50 50 20 0.15 90 75 60 ------216 20 45 0 135 70 45 0.25 0.6 50 50 20 0.15 90 75 60 ------217 30 65 0 155 60 65 0.25 0.6 50 50 20 0.15 90 75 60 ------218 35 75 0 165 55 75 0.25 0.6 50 50 20 0.15 90 75 60 ------219 25 60 0 150 65 60 0.25 0.6 50 50 20 0.15 90 75 60 ------220 35 80 0 170 55 80 0.25 0.6 50 50 20 0.15 90 75 60 ------221 30 90 0 180 60 90 0.25 0.6 50 50 20 0.15 90 75 60 ------222 30 65 0 155 60 65 0.25 0.6 50 50 20 0.15 90 75 60 ------223 10 340 0 430 80 340 0.25 0.6 50 50 20 0.15 90 75 60 ------225 20 265 0 355 70 265 0.25 0.6 50 50 20 0.15 90 75 60 ------226 25 260 0 350 65 260 0.25 0.6 50 50 20 0.15 90 75 60 ------227 30 305 0 395 60 305 0.25 0.6 50 50 20 0.15 90 75 60 ------229 20 315 0 405 70 315 0.25 0.6 50 50 20 0.15 90 75 60 ------
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Table 17.2.3_4 Husab Uranium Project (Zone 3 and 4) Variogram Parameters for Mineralised Zones 3 and 4
Orientation Range 1 (m) Range 2 (m) Range 3 (m)
Major Semi-Major Minor C0 C1 C2 C3 Zone Semi- Semi- Semi- (%) (%) Major Minor (%) Major Minor (%) Major Minor Dip Azi Dip Azi Dip Azi Major Major Major (o) (o) (o) (o) (o) (o) 351 20 120 0 210 70 120 0.25 0.6 200 100 5 0.15 250 175 30 ------352 25 95 0 185 65 95 0.25 0.6 200 100 5 0.15 250 175 30 ------353 25 85 0 175 65 85 0.25 0.6 200 100 5 0.15 250 175 30 ------354 50 105 0 195 40 105 0.25 0.6 200 100 5 0.15 250 175 30 ------355 30 105 0 195 60 105 0.25 0.6 200 100 5 0.15 250 175 30 ------356 30 105 0 195 60 105 0.25 0.6 200 100 5 0.15 250 175 30 ------357 30 75 0 165 60 75 0.25 0.6 200 100 5 0.15 250 175 30 ------440 15 250 0 340 75 250 0.25 0.6 200 100 5 0.15 250 175 30 ------
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Figure 17.2.3_2 Correlogram for Husab Uranium Project Zone 2 Combined Lodes
upper left (Major) 0º 015º, upper right (Semi-Major) 15º 105º, lower left (Minor) -75º 105º
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17.2.4 Block Model
A block model was created in the National grid (UTM WGS84 33south) using Datamine mining software with a parent cell size of 25m (Easting) by 25m (Northing) by 15m (RL) which was sub- blocked down to 3.125m (Easting) by 3.125m (Northing) by 1.875m (RL). No rotation was applied to the block model. The block model parameters are summarised below in Table 17.2.4_1. The variables coded into the block model are shown below in Table 17.2.4_2.
Table 17.2.4_1 Husab Uranium Project (Zones 2 - 4) Block Model Parameters
Easting (X) Northing (Y) RL (Z) Min. Coordinates 7,499,500 503,500 0 Extent 7,508,000 508,000 600 Block size (m) 25 25 15 Sub Block size (m) 3.125 3.125 1.875 Rotation 0° 0° 0° Total Blocks – Mineralisation Total Blocks – Final Model 8,043,689
Table 17.2.4_2 Husab Uranium Project (Zone 2 - 4) Block Model Variables
Attribute Name Type Decimals Background Description RESCAT Integer - 4 Classification (1=meas; 2=ind; 3=inf, 4=unclass) RCLASS Character - UNC Classification (MEAS; IND; INF, UNC) DENSITY Real 2 2.65 Estimated Density from top-cut composites Weathering Code 1=soil/alluvium, 2=Laterite, 3=oxidised, 4=partially oxidised, GROUND Integer - 6 5=Transitional, 6=fresh, 99=dumps, 199=open pit, 0=air OXIDN Character - PMRY Weathering (OXID, TRN1, TRN2, PMRY, DUMP, AIR) 4 domains: Zone 1- Sub-Zones 1-39 & 666; Zone 2 – Sub-zones 1-29; ZONE Integer - 0 Zone 3 – Sub-Zones 351-357; and Zone 4 – Sub-zones 440 Rock types (CHUS, OVB, KARB, ROSG, KHAN, CALC, GEOL Character - ROCK ALSK, GRN) Is Mined Flag: 0=not mined, 1=mined open pit, MINED Integer - 0 2=mined UG Development, 3=mined UG stopes, 4=dumps U3O8c Real - 5 Estimated Uranium from top-cut composites Sulphide Presence Flag; SULPHIDE Integer - 0 1= yes, 0 = no using 0.5 SU_PROPN SU_PROPN Real - 0 Estimated Sulphide presence
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17.2.5 Grade Estimation
OK Estimate
U3O8 grade was estimated into to the block model using OK for top-cut U 3O8 related variables. Sample neighbourhood testing was adopted from the Zone 1 data.
As the Extract drilling had been completed on a regular grid pattern, drillhole data clustering was not significant and similar sample selection criteria were used for the mineralised zones. A staged sample search strategy for Zones 2, 3, and 4 was used based upon neighbourhood testing and is summarised in Tables 17.2.5_1 and 17.2.5_2.
The variogram parameters used for the estimation were based upon the variography discussed in Section 17.2.3 and are summarised in Table 17.2.5_1. Domain control was used for the OK estimate using whole block discretisation of 5(X) by 5 (Y) by 3(Z). Any sub-blocks within the 3D limit of each whole block were assigned the whole block OK estimate.
Validation
Validation routines were undertaken in a similar manner to those of Zone 1 (Section 17.1.8). Comparison of average informing composite grade and average block model grade, by easting, northing and elevation, for Zones 2, 3 and 4, are shown in Figures 17.2.5_1 to 17.2.5_3.
Overall, there was a good comparison between the informing composite data input into the model and the resulting block grades. Areas of significant variance within the block model were also dealt with in the block model by limiting the amount of Inferred classification.
Density
The bulk density values used for the resource model were based upon the data previously analysed in this Section (see Table 17.2.3_2). All mineralised material inherited the bulk density from the host lithology.
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Figure 17.2.5_1 Comparative Plots of Informing Composites and Block Model Grade Husab Uranium Project – Zone 2
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Figure 17.2.5_2 Comparative Plots of Informing Composites and Block Model Grade Husab Uranium Project – Zone 3
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Figure 17.2.5_3 Comparative Plots of Informing Composites and Block Model Grade Husab Uranium Project – Zone 4
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Table 17.2.5_1 Husab Uranium Project – Zone 2 Sample Search Parameters – Ordinary Kriging
Pass No. Major Semi-Major Minor Search Radii Number of Composites Datset Used Dip Azim Dip Azim Dip Azim Minimum (Sub-ZONE) 1 2 3 Major Semi- Major Minor Maximum (o) (o) (o) (o) (o) (o) (run 1 / 2 / 3) 201 89.6% 10.4% 0.0% 20 45 0 135 70 45 120 120 40 14 / 8 / 4 24 202 45.6% 54.4% 0.0% 35 60 0 150 55 60 120 120 40 14 / 8 / 4 24 203 91.8% 8.2% 0.0% 25 70 0 160 65 70 120 120 40 14 / 8 / 4 24 204 97.2% 2.8% 0.0% 25 70 0 160 65 70 120 120 40 14 / 8 / 4 24 205 77.6% 22.4% 0.0% 25 80 0 170 65 80 120 120 40 14 / 8 / 4 24 206 97.3% 2.7% 0.0% 25 60 0 150 65 60 120 120 40 14 / 8 / 4 24 207 78.9% 21.1% 0.0% 30 90 0 180 60 90 120 120 40 14 / 8 / 4 24 208 0.0% 100.0% 0.0% 25 90 0 180 65 90 120 120 40 14 / 8 / 4 24 209 51.2% 45.7% 3.0% 30 70 0 160 60 70 120 120 40 14 / 8 / 4 24 210 92.4% 7.6% 0.0% 40 80 0 170 50 80 120 120 40 14 / 8 / 4 24 211 96.4% 3.6% 0.0% 25 90 0 180 65 90 120 120 40 14 / 8 / 4 24 212 94.3% 5.7% 0.0% 25 95 0 185 65 95 120 120 40 14 / 8 / 4 24 213 80.2% 19.8% 0.0% 25 90 0 180 65 90 120 120 40 14 / 8 / 4 24 214 86.2% 13.8% 0.0% 10 35 0 125 80 35 120 120 40 14 / 8 / 4 24 215 37.8% 62.2% 0.0% 25 90 0 180 65 90 120 120 40 14 / 8 / 4 24 216 96.3% 3.7% 0.0% 20 45 0 135 70 45 120 120 40 14 / 8 / 4 24 217 77.3% 22.7% 0.0% 30 65 0 155 60 65 120 120 40 14 / 8 / 4 24 218 71.2% 28.8% 0.0% 35 75 0 165 55 75 120 120 40 14 / 8 / 4 24 219 0.9% 94.9% 4.2% 25 60 0 150 65 60 120 120 40 14 / 8 / 4 24 220 98.5% 1.5% 0.0% 35 80 0 170 55 80 120 120 40 14 / 8 / 4 24 221 68.7% 31.3% 0.0% 30 90 0 180 60 90 120 120 40 14 / 8 / 4 24 222 69.2% 30.8% 0.0% 30 65 0 155 60 65 120 120 40 14 / 8 / 4 24 223 60.0% 40.0% 0.0% 10 340 0 430 80 340 120 120 40 14 / 8 / 4 24 225 54.7% 45.3% 0.0% 20 265 0 355 70 265 120 120 40 14 / 8 / 4 24 226 63.2% 36.8% 0.0% 25 260 0 350 65 260 120 120 40 14 / 8 / 4 24 227 92.5% 7.5% 0.0% 30 305 0 395 60 305 120 120 40 14 / 8 / 4 24 229 89.6% 10.4% 0.0% 20 315 0 405 70 315 120 120 40 14 / 8 / 4 24
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Table 17.2.5_2 Husab Uranium Project – Zones 3 & 4 Sample Search Parameters – Ordinary Kriging
Pass No. Major Semi-Major Minor Search Radii Number of Composites Datset Used Dip Azim Dip Azim Dip Azim Minimum (Sub-ZONE) 1 2 3 Major Semi- Major Minor Maximum (o) (o) (o) (o) (o) (o) (run 1 / 2 / 3) 351 0.0% 92.7% 7.3% 20 120 0 210 70 120 120 120 40 14 / 8 / 4 24 352 0.0% 89.6% 10.4% 25 95 0 185 65 95 120 120 40 14 / 8 / 4 24 353 7.5% 90.3% 2.2% 25 85 0 175 65 85 120 120 40 14 / 8 / 4 24 354 0.0% 93.8% 6.2% 50 105 0 195 40 105 120 120 40 14 / 8 / 4 24 355 1.1% 95.8% 3.1% 30 105 0 195 60 105 120 120 40 14 / 8 / 4 24 356 0.0% 86.2% 13.8% 30 105 0 195 60 105 120 120 40 14 / 8 / 4 24 357 82.9% 17.1% 0.0% 30 75 0 165 60 75 120 120 40 14 / 8 / 4 24 440 20.9% 75.9% 3.3% 15 250 0 340 75 250 120 120 40 14 / 8 / 4 24
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Resource Reporting and Classification
The resource estimate for the Husab Uranium Project deposit – Zones 2 to 4 - has been categorised in accordance with the criteria laid out in the NI43-101 and the JORC Code (2004). Indicated and Inferred Mineral Resources were defined during the validation of the grade estimates, with detailed consideration of the CNI43-101 categorisation guidelines.
The classification of the Indicated and Inferred Mineral Resources was based on the confidence level of the key criteria that were considered during resource classification as presented in Table 17.2.5_3
Table 17.2.5_3 Husab Uranium Project (Zones 2 to 4) Confidence Levels of Key Categorisation Criteria
Items Discussion Confidence Drilling Techniques RC/Diamond - industry standard approach. High Logging Standard nomenclature applied with recording and apparent high quality. High Drill Sample Recovery Recorded as good High Sub-sampling Techniques Industry standard for both RC and diamond drilling High and Sample Preparation Quality of Assay Data Good internal laboratory and external quality control data available for the majority Moderate to of the chemical assaying. Factored radiometric data is considered to be globally High equivalent to chemical assaying, but can show local differences. Verification of Sampling QAQC analysis is within industry acceptable standards. High and Assaying Location of Sampling Most drillhole collars surveyed by DGPS surveyed and most drillholes have been High Points downhole surveyed. Data Density and Nominal 50m by 50m to 100m by 200m drillhole collar spacing. Moderate to Distribution High Audits or Reviews Coffey Mining has reviewed the site drilling and sampling procedures. High Database Integrity No material errors identified. High Geological Interpretation Infill drilling is likely to change the mineralisation shapes and understanding of Low to structural and grade continuity Moderate Estimation and Modelling Estimates based on detailed statistical and geostatistical analysis. Moderate Techniques Cutoff Grades Range of cutoff grades reported. High Mining Factors or Whole block estimates for all mineralised regions completed. The effect of N/A Assumptions emulating smaller mining blocks has not been investigated.
Blocks were classified as Indicated Mineral Resources based upon regions which had well established geological continuity and a nominal data spacing of 50m by 50m to 50m by 100m. Blocks not classified as Indicated Mineral Resources and which had a reasonable geological continuity and a data spacing of 100m by 100m to 100m by 200m were classified as Inferred Mineral Resources.
No mining has been undertaken at Husab Uranium Project to date.
The resource estimate for Zones 2 to 4 of the Husab Uranium Project, reported above selected cutoffs to a depth of 500m, is summarised below. The preferred cutoff for reporting
the resources is 100ppm U 3O8.
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Table 17.2.5_4 Husab Uranium Project – Zones 2 to 4 August 6 2010 Resource Estimate by Zone
Reported at various cutoffs, Preferred cutoff : 100ppm U 3O8 Ordinary Kriged Estimate based upon 3m Top-cut U 3O8 Composites Parent Cell Dimensions of 25mNS by 25mEW by 15mRL
Lower Cutoff Tonnage Grade Contained U O Contained U O 3 8 3 8 (ppm U 3O8) (Mt) (ppm U 3O8) (MKg) (MLb) Zone 2 100 118.8 520 61.8 136.9 200 110.0 550 60.5 133.7 Indicated 300 87.7 630 55.3 121.2 400 66.3 720 47.8 104.8 100 26.8 520 13.9 30.5 200 24.5 550 13.5 29.7 Inferred 300 19.3 630 12.1 26.8 400 14.1 740 10.4 22.9 Zone 3 100 43.0 250 10.7 24.0 200 24.3 330 8.0 17.5 Inferred 300 11.8 410 4.9 10.7 400 4.3 530 2.3 5.0 Zone 4 100 14.4 570 8.2 17.9 200 13.0 610 7.9 17.5 Inferred 300 11.4 660 7.5 16.6 400 9.0 740 6.7 14.7 Note: Figures have been rounded.
Table 17.2.5_5 Husab Uranium Project - August 6 2010 Resource Estimate - All Zones
Reported at various cutoffs, Preferred cutoff : 100ppm U 3O8 Ordinary Kriged Estimate based upon 3m Top-cut U 3O8 Composites Parent Cell Dimensions of 25mNS by 25mEW by 15mRL
Lower Cutoff Tonnage Grade Contained U 3O8 Contained U 3O (ppm U 3O8) (Mt) (ppm U 3O8) (MKg) (MLb) Indicated 100 241.0 480 117 257.0 200 214.5 520 112 247.6 300 165.7 610 101 220.6 400 122.5 700 85 187.4 Inferred 100 125.5 400 50 110.3 200 96.1 470 45 99.9 300 67.2 570 38 84.0 400 44.4 680 30 66.7 Note: Figures have been rounded.
Coffey Mining is unaware of any mining, metallurgical, infrastructure or other relevant factors which may materially affect the resources for Zones 1 to 4. The availability of suitable water and power supplies may be key factors in any future mining studies.
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17.2.6 Comments and Recommendations
The Husab Uranium Project area contains significant uranium mineralisation associated with uraniferous leucocratic granites (alaskites) within the highly prospective Central Zone of the Damara Orogeny. The mineralised alaskites tend to occur along or proximal to the unconformity contact between the Khan Formation and Rössing Formation.
The August 2010 resource (Table 1.4_1; and Tables 17.1.9_2 and 17.2.5_4) represents a significant increase in Indicated Mineral Resources relative to the previous July 2009 Resources for Zones 1 and 2, and incorporates maiden resources for Zones 3 and 4.
Coffey Mining has reviewed the drilling, sampling, assaying and field procedures used by Extract and consider them to be of high quality.
Further bulk density information is required in the mineralised portions of the deposit and of the overburden. Further infill and extensional drilling is required to raise the level of confidence of the Indicated and Inferred Mineral Resources.
Due to the increasingly complex relationship of mineralisation and alaskite lithologies, it is recommended that in future an MIK approach be considered to generate a recoverable SMU model.
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18 MINERAL RESERVE ESTIMATES AND MINING METHODS
18.1 Mining operations
18.1.1 Mining Methodology
The Husab Project will be developed as two large open pits centred around the Zone 1 and Zone 2 orebodies. The mining process will follow a conventional open pit approach entailing planning, drilling, blasting, loading and hauling from the open pits. Ore will be transported either directly to the primary crusher or placed in stockpiles for later crusher feed as required. Waste material will either be deposited wi th the Mine Residue Facility (MRF) or utilised for site earthworks if it is of a suitable quality. The mine will operate 361 days per annum (allowing for 2 public holidays and 2 lost days for weather delays) on a 24 hour basis with 3 rotating shifts of 8 hours each.
Due to the lower cost of power relative to fuel price it has been determined that considerable benefit can be generated by adopting a trolley -assist mine haulage system. Trolley -assist is a truck based haulage system whereby diesel -electric tr ucks are provided with AC electrical power directly from an overhead line in the same manner as a city tram. An indicative layout is shown in Figure 18.1.1_1.
Figure 18.1.1_1 Trolley Assist Layout
The benefits of trolley-assist include increased truck speed up ramp and therefore improved fleet productivity and decreased fuel consumption due to the trucks being effectively idle when “on trolley”. With the increased productivity is a decrease in the required fleet size and therefore associated capital expenditure. However, there are additional costs associated with the trolley infrastructure and required truck modifications. There is a minimal amount of additional stripping needed as a result of a w ider ramp to allow for the trolley infrastructure and straightening of the pit ramps to aid the truck operators to maintain the pantograph of the truck in effective contact with the overhead trolley line.
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Initial assessment of the trolley assist-system indicated that a considerable saving in cost was achievable over the life of the operation. A final sensitivity analysis, carried out using the DFS cost model, generated a Life of Mine (LOM) operating cost saving of ~9% or $0.17/tonne. The LOM capital costs increased by ~7% due to the significant trolley infrastructure required, resulting in a total cost saving of ~6% or $230M over LOM. This equates to a saving of 5% in Net Present Cost over LOM.
The benefit of trolley-assist increases with higher fuel prices. As such trolley-assist provides a degree of risk mitigation against adverse fluctuations in this fundamental cost parameter. Conversely, trolley-assist is sensitive to changes in power price. Sensitivity analysis indicates that a ~150% increase in power cost would be required for the trolley-assist cost of mining to converge on the conventional haulage cost. Conversely fuel cost would need to decrease by ~80% to equate to the base case trolley-assist cost. It is highly unlikely that fuel prices will decrease to this level. Power prices may increase, but it is reasonable to assume fuel will also increase, maintaining the relative benefit of trolley-assist.
A MRF has been developed to cater for storage of both mine waste rock and process plant tailings. This disposal / storage option is essentially a facility whereby tailings are transported onto the dumping area by truck along with the mine waste and inter-mingled at the dump face.
18.2 Open Pit Optimisation
The objective of open pit optimisation is the determination of a generalised open pit shape (shell) that provides the highest value for a deposit. It is from analysis of all the shells generated in the optimisation process that a single shape can be selected as the guide for a practical ultimate pit design.
The final pit design defines the Ore Reserve and subsequently LOM schedules and associated cashflows can be developed. Hence, the pit optimisation process is the critical first step in the development of any open cut mineral extraction project as it not only assesses mining and processing parameters, but other global project variables such as marketing and financial assumptions.
In addition to defining the ultimate size of the open pit, the pit optimisation process also provides an indication of potential areas for interim mining stages. These intermediate mining stages allow the pit to be developed in a practical and incremental manner, while at the same time targeting high grade ore.
The pit optimisation process used the latest available information to ensure it makes use of all accumulated knowledge at the time. The parameters include and are not limited to:
The JORC classified Resource completed by Coffey Mining Pty in August 2010 (ASX release 10 August, 2010). This will be referred to as the Resource Model.
Updated geotechnical parameters supplied by Golder Associates in mid 2010.
Updated mining operating and capital costs, mining parameters and bench height recommendations derived from earlier studies carried out during 2010.
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Updated process recovery, project capital costs, processing costs, selling costs, mining and processing production rates inclusive of respective ramp-ups provided by AMEC during 3 rd quarter 2010.
18.2.1 Resource Model
The starting point for any optimisation is the resource model. The model utilised for the DFS was generated by Coffey Mining in July 2010 utilising the latest drilling information available. It was provided in Datamine format as one large model combining the four areas of uranium mineralisation currently classified as Mineral Resources, “Zones 1, 2, 3 and 4”. However, for the purposes of the DFS only Zones 1 and 2 were considered.
The Mineral Resource model was then imported into MineSight, the mine planning software utilised by ORElogy. The sub-blocked resource model was regularised as part of this importation process. However, the detail provided by the sub-blocks has been maintained through the use of “block percent” parameters that preserve the block proportions for the sub- blocks and their associated materials and resource classes.
The result of the importation into MineSight was reconciled back to the original Datamine resource model.
It is important to select a block size that takes into account the machinery size required to achieve productivity targets and their respective selective mining capabilities. Equipment size was considered in the bench height analysis (refer to Section 18.2.4), resulting in a minimum bench height of 7.5m for selective mining. As the block model has to reflect the smallest bench height to be used, a regular block size (x, y, z) of 12.5m x 12.5m x 7.5m was adopted within the MineSight model.
18.2.2 Density
All the blocks within the model that contained grade also had a modelled density. However, some further refinement of the waste densities was carried out as part of building the MineSight mining model.
The original Mineral Resource model was flagged with two material codes identifying calcrete (i.e. barren overburden material) from fresh rock. These two material types were then utilised to flag the global densities of the deposit.
A third material type was then flagged in the MineSight model. This material was referred to as “free-dig” material, denoting weathered overburden material that could be freely excavated with a shovel without the need for drilling and blasting. It was assumed free-dig constituted the first 20m from topography as recommended by an independent blasting consultant, and was coded into the model accordingly. The densities assigned to the material types described above are detailed in Table 18.2.2_1 below.
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Table 18.2.2_1 Husab Uranium Project Material Density
Material t/m³ Free-dig 1.80 Calcrete 2.20 Fresh 2.65
18.2.3 Ore Loss and Dilution
For the purposes of pit optimisation and the subsequent reporting of mineable reserves an allowance must be made for ore loss and dilution. It was determined that the most appropriate approach was to apply the parameters on a block by block basis within the resource model. The resulting regularised mining model has ore percentages and ore grades that have ore loss and dilution incorporated.
The process used to determine ore loss and dilution is as follows:
The dip of the ore body was estimated for every block throughout the model using the three dimensional ore body outlines developed for resource estimation.
A custom software script performed the following calculations:
If the ore percentage in the block was less than 5%, it was assumed the ore could not be practically isolated and mined and therefore ore loss was set to 100% (i.e. the block ore percent is set to 0, effectively reclassifying the block as waste).
If the ore percent was greater than 95%, the block was considered within the ore body and could be mined without ore loss or dilution.
For the ore percentages lying between these two bounds, block specific dilution and ore loss were calculated on the basis of the geometry of the orebody boundary and the geometry of the advancing mining face within that block. This assumed that the majority of mining would take place across the strike of the orebody and with the dip (i.e. the orebody dip azimuth and mining face azimuth are in the same direction) as shown in Figure 18.2.3_1.
For orebody dips of less than 10° the dilution and oreloss from the above approach becomes prohibitive. In this case it is assumed a combination of variable depth drill and blast with selective dozing will result in effectively a dilution skin of ~1m (i.e. ~13% of bench height).
The resulting ore loss and dilution estimates were coded within the block model, and a revised ore percentage and grade were calculated.
As the dilution and ore loss vary from block to block the values for any given optimisation shell or pit design will vary depending on the blocks within the shell or design. As an indication, the global ore loss and dilution for the entire resource are provided in Table 18.2.3_1.
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Figure 18.2.3_1 Ore Loss Dilution Geometry (Orebody Dip ≥ 10°)
Table 18.2.3_1 Husab Uranium Project Global Ore Loss and Dilution by Zone
Zone Ore Loss (%) Dilution (%) 1 3.5 3.8 2 4.9 7.1 Average 4.2 5.4
18.2.4 Bench Height Selection
A bench height study was undertaken to determine the most appropriate bench height for the project. The selection process balanced the benefits of an increased bench height against the impact of ore losses and dilution.
Higher benches have the following operational benefits:
A lower bench turnover rate - Bench turnover measures of the number of benches that are commenced and completed during a given year. Mining of a bench involves a number of separate activities (for example: drop cutting, drilling, blasting, grade control, ore mark-out, load and haul, working area maintenance, final wall pre-split and battering).
Larger equipment size – A higher bench height means that larger equipment can be utilised. Larger equipment provides the benefits of:
Higher productivity (t/h) and therefore lower unit operating cost (US$/t)
Reduced fleet size simplifying the operation and leading to a reduction in:
Overall personnel Road maintenance equipment Operational safety incidents. Productivity also generally increases due to less congestion.
These benefits will ensure the required material movement of 15Mt/a mill feed is maintained.
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However, increased bench height also results in less ore selectivity along the ore and waste boundary. This results in:
Higher ore loss where ore is lost to waste; and
Higher mining dilution where ore is diluted with waste, which has the effect of increasing ore tonnes but lowering mill feed grade.
The bench height study assessed the trade-off between these various parameters by applying the ore loss and dilution values calculated for various bench heights within WHITTLE and assessing the relative variations in ore inventory and Net Present Value. As a result of this study a maximum bench height of 7.5m has been selected in those areas where selective mining is required to minimise ore loss and dilution. In broad continuous areas of either ore or waste, a 15m bench height is planned.
18.2.5 Geotechnical Parameters
Golder Associates SA completed a detailed geotechnical analysis of the Swakop Uranium Husab Project. A brief synopsis of their findings is provided below:
Three major geotechnical domains were identified, being free-dig, calcrete and fresh rock.
Kinematic analysis was undertaken where three modes of failure were examined for each of the sectors, and a slope configuration calculated based upon the selected bench height.
Inter-ramp stability was also assessed using probabilistic techniques.
In general, the geotechnical investigations demonstrated that the rock mass conditions are good and will allow for fairly steep pit slopes. On the smaller bench scale, there is potential to develop wedge or planar failures in areas due to the intersection of joints and batters. However, the calculated factors of safety have highlighted that these should not present a significant risk. The risk associated with these types of failures can by mitigated by maintaining good blasting practices and batter slopes.
With the exception of the free-dig domain, three different slope profiles were developed based on the level of probability of failure, and hence risk. These profiles were termed Conservative, Intermediate and Aggressive. The Intermediate parameters have been selected for optimisation and design purposes and they are outlined in Table 18.2.5_1.
To calculate the overall slopes for application in the pit optimisations software (Whittle 4X), allowances were made for the following:
Ramp width inclusive of trolley-assist infrastructure where applicable
Approximate depth of pit
Number of ramps
Number of catch benches in fresh rock (refer to Table 18.2.5_1).
Table 18.2.5_2 details the calculation of total ramp width (i.e. crest to toe).
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Table 18.2.5_1 Husab Uranium Project Geotechnical Slope Design Parameters
Domain Detail Units Value Bench Height (m) 15 Batter Angle (deg) 45 Free-Dig Berm Width (m) 6.5 Inter-ramp angle (deg) 35 Bench Height (m) 15 Batter Angle (m) 70 Calcrete Berm Width (m) 12 Overall slope inclusive of Catch Berms (deg) 41 Bench Height (m) 30 Batter Angle (deg) 80 Berm Width (m) 15 Fresh Inter-ramp Angle (deg) 56 Catch berm width (m) 25 Number of benches between catch berms (#) 3 Overall slope inclusive of catch berm (deg) 54.5
Table 18.2.5_2 Husab Uranium Project Ramp Width Calculation
Truck Width Number of Trolley-Assist Ramp Type Total (m) Truck Widths Infrastructure (m) Trolley 8.5 4 4 38 Non-Trolley 8.5 4 0 34 1-Way 8.5 2 0 17
The calculation of the overall slopes for Whittle is detailed in Table 18.2.5_3 below. As the free-dig and calcrete depths are relatively small, no allowance was made for ramps in this material as any ramp would only impact over a limited length of wall. The depth of material types and ramp layouts were based on previously developed pit designs.
Table 18.2.5_3 Husab Uranium Project Overall Slope Calculations
Inter- Number of Ramps Number of Slope Wall Overall Wall Material Ramp Catch Area Trolley Non-Trolley 1-Way Height Slope Angle Berms 1 35 - - - - 20 35 2 35 - - - - 20 35 Free Dig 3 35 - - - - 20 35 4 35 - - - - 20 35 1 41 - - - - 30 41 2 41 - - - - 30 41 Calcrete 3 41 - - - - 40 41 4 41 - - - - 50 41 1 56 2 - - 1 320 45 2 56 2 - - 1 320 45 Fresh 3 56 1 2 1 2 400 43 4 56 2 - - 2 300 42
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Figure 18.2.5_1 shows the location of the slope areas referred to in Table 18.2.5_3 above. The areas were defined utilising the pit designs shown in Figure 18.2.5_1, which were the most recent design versions available at the time the optimisation work commenced. The designs provided the best approximation of a practical final ramp layout as the shape and size of the pits were not expected to change fundamentally. Zone 1 pit has three slope areas due to its relatively complex geometry and ramp layout. Zone 2 pit is a single slope area as its shape is relatively simple and effectively has one ramp along each wall down to a very flat pit base.
Figure 18.2.5_1 Whittle Slope Areas
Slope Area 1
Slope Area 2
Slope Area 3
Slope Area 4
18.2.6 Mining Costs
Mining costs for the purpose of optimisation were based on a prior mining study leading to the DFS (Pre-DFS Report – July 2010). The study was based on a conventional hydraulic shovel and rigid dump truck mining methodology, but with the addition of trolley-assist haulage on main in-pit and dump ramps. Costs were developed from first principles and incorporated equipment, personnel and consumables costs for all aspects of the mining operation. The costs were generated by pit stage, bench level and material type. These costs were then applied to a LOM schedule that included construction of an associated waste dump. By this means, the waste haulage cost from any given bench could be more accurately estimated by accounting for the advance of the waste dump over time.
From this array of costs a weighted average mining cost for ore and for waste, by zone and by bench, was generated. These costs were then approximated by either a linear or polynomial function. A script was generated to calculate a mining cost for every block in the model. This model was then exported to Whittle.
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The Whittle optimisation algorithm works on the basis that all material, as a minimum requirement, will be mined to the surface and deposited on the waste dump. Therefore this “waste mining cost” is applied as the mining cost to all material. Whittle then determines whether a block of material should be processed on the basis of the block generating more revenue than any additional costs incurred. The additional costs can be:
Processing Cost – The cost of sending the block through the plant
Selling Cost – This may be any combination of:
A processing related cost, depending on the type of extraction process being used
A cost attracted by the product such as shipping or marketing
A royalty and /or tax on the final product
Incremental Ore Mining Cost – The cost difference between mining materials as waste to the waste dump versus ore to the crusher. This cost may consist of:
Grade control costs
Fixed administrative and overhead costs
Stockpile re-handle costs
The cost differential of a longer or shorter haul. This differential may be positive or negative depending on the location of the different destinations.
The waste mining cost is applied as the Base Mining Cost across all blocks. The additional Incremental Ore Mining Cost is included with the Processing Cost so as to only be applied to the ore.
A summary of the global weighted costs, as used in the optimisation, is provided in Table 18.2.6_1 below. Note that final economic evaluation of the project has been performed based on updated cost estimates, as described in the feasibility study.
Table 18.2.6_1 Husab Uranium Project Global Average Unit Mining Costs (US$/t)
Waste Zone Ore Total Free-Dig Calcrete Fresh at 7.5m Fresh at 15m Average Zone 1 $ 2.81 $ 0.82 $ 1.07 $1.91 $ 1.50 $ 1.44 $ 1.62 Zone 2 $ 2.83 $ 0.75 $ 1.02 $1.90 $ 1.46 $ 1.33 $ 1.52 Total $ 2.82 $ 0.79 $ 1.03 $ 1.91 $ 1.48 $ 1.39 $ 1.57
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18.2.7 Processing Costs, Mill Throughput and Timing
Mill throughput is provided in terms of a percentage of the plant design capacity of 15Mt/a as shown in Figure 18.2.7_1.
Figure 18.2.7_1 Process Throughput Ramp-Up
Based on the curve and a plant nameplate capacity of 15Mtpa, annual throughputs were calculated and used in this optimisation.
The ramp-up in plant throughput produces an associated variation in processing cost. This is because the plant operating costs consist of a component that varies with throughput tonnes and a fixed component. When the plant is operating at its low initial throughput, the annual fixed costs are spread over a lower annual tonnage, thereby generating a higher unit cost. As the mill throughput rate increases, the unit cost for the fixed component decreases accordingly.
The process costs and throughput rates as utilised in Whittle optimisation are provided in Table 18.2.7_1. Note that economic evaluation of the project has been performed based on updated cost estimates, as described in the feasibility study.
Table 18.2.7_1 Husab Uranium Project Process Throughput and Costs
Mill Throughput Processing Cost Year (Mt/a) (US$ / ROM tonne) Year 1 9.5 US$ 10.71 Year 2 14.6 US$ 9.96 Year 3 15.0 US$ 9.92
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18.2.8 Processing Recovery and Recovery Ramp-Up
A significant amount of metallurgical testwork had been undertaken to estimate the process plant performance.
The uranium recovery was quantified on the basis of an average tailings grade by zone. In addition there was also a fixed loss through the process plant. These are detailed in Table 18.2.8_1.
Table 18.2.8_1 Husab Uranium Project Process Losses
Zone Grade (ppm) Zone 1 55 Tailings Residual Grade Zone 2 44 Zone 1 Plant Loss 9 Zone 2
Utilising a fixed residual grade as the basis for recovery means that recovery is not a fixed percentage of the head grade, but in fact varies with the head grade, consequently the recovered grade has been calculated on a block by block basis in the model, resulting in a final grade that accounts for mining ore loss, mining dilution and process losses. This is the uranium grade utilised in the optimisation process.
However, there is also a ramp-up profile associated with the process recovery over the initial three years of operation. Therefore, Whittle utilised the diluted and recovered grade in the block model as detailed above and applied the process recovery ramp-up to 100% as detailed in Table 18.2.8_2.
Table 18.2.8_2 Husab Uranium Project Process Recovery Ramp-Up
Year Effective Process Recovery (%) Year 1 70% Year 2 90% Year 3 100%
18.2.9 Price, Selling Cost and Royalties
The following parameters were used:
A base price of US$65.00/lb of recovered U 3O8 is used.
A selling cost of US$2.50/lb that accounts for transportation and marketing costs.
A Namibian government royalty of 3%.
Consequently the net price for the purposes of open pit optimisation is US$60.55/lb.
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18.2.10 Capital Expenditure and Discount Rate
The capital expenditure requirements for the project are a combination of mining equipment, process plant and supporting infrastructure. It includes, but is not limited to, the following components:
Process plant and infrastructure including process and administration buildings.
Tailings facility (disposal pad and stormwater management).
Production, ancillary and support mining equipment.
Mining workshops, associated storage, support and administration facilities.
Process maintenance and administration facilities.
Power and water reticulation.
Access to site (road and railway).
EPCM and owner costs.
Process plant and associated infrastructure costs were provided by AMEC Minproc in August 2010. The costs of the mining complex were also generated by AMEC Minproc as part of a previous study carried out early in 2010. The costs associated with the mining equipment were generated by ORElogy. This was on the basis of budgetary costs received by OEM’s during the 3 rd quarter 2010 and included all production, support and ancillary equipment.
The scheduling of capital within WHITTLE is detailed in Table 18.2.10_1 below and was based on:
Processing plant and infrastructure – A plant start date of January 2014, with the capital spread over the previous two years on a 3/8 and 5/8 split respectively.
Mining equipment and infrastructure – The most up-to-date mining schedule and associated capital cost schedule developed in Q2 2010. It was based on a July 2012 start date for mining (i.e. 18 months of pre-strip prior to plant commissioning). The capital for mining infrastructure is applied in 2012 (i.e. Year 1).
The capital cost estimate presented here is used for the purposes of open pit optimisation. An updated capital cost estimate, as used in economic evaluation of the project for the purposes of the feasibility study, is shown in Section 23.
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Table 18.2.10_1 Husab Uranium Project Capital Costs (US$ M)
Mining Equipment and Process Plant and Year Mining Complex Infrastructure 2012 US$188.50 US$357.40 2013 US$90.30 US$595.70 2014 US$22.30 2015 US$10.20 2016 US$9.70 2017 US$27.70 2018 US$7.50 2019 US$58.50 2020 US$8.10 2021 US$17.90 2022 US$39.00 2023 US$122.80 2024 US$7.60 2025 US$14.40 2026 US$0.30 2027 US$0.00 Total US$625.70 US$953.20 For all optimisations carried out for this study a discount rate of 10% was applied.
18.2.11 Whittle Optimisation Process
The process that Whittle undertakes is to vary the base input price both up and down to produce a series of shells. Each shell produced generates the maximum cashflow for the input parameters and its associated factored price. The lower factored price will produce smaller shells, the higher price larger shells resulting in a set on “nested” shells with the lower valued shell lying inside the higher valued shells. These nested shells are useful for a number of reasons:
The smaller shells give a guide to where initial mining should occur as the smaller shells will drive towards the areas of highest value in the ore body.
The larger shells provide an indication of how much additional mineralisation may become economic, or alternatively what current ore may become unviable should parameters change in the future.
Allows Whittle to develop a schedule for mining the deposit over time and therefore a discounted cashflow to be generated.
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Along with the undiscounted cashflow of each shell, Whittle generates the following two standard discounted cashflows:
Worst case discounted cashflow assumes that, for any given shell, extraction is undertaken sequentially from the top to bottom of the shell level by level. This means overburden is removed in advance and there are no interim shells or stages to access higher value ore earlier. This is clearly not a preferred pit extraction scenario from a value perspective.
Best case discounted cashflow assumes that for any given shell, extraction is undertaken sequentially from the smallest shell generated by the optimisation out to the largest shell selected. This provides the best value for that largest shell. However, this approach is constrained by practical considerations such as:
The distance between successive shells potentially being too narrow to mine.
The shells being too tight and small to allow for practical mining.
These two discounted cashflows provide the extremities of the possible value that a shell is able to generate. A shell is selected as the basis for subsequent pit designs that is cognisant of a number of potentially conflicting practical and strategic objectives such as:
The capability to generate a practical schedule in line with Whittle outputs.
Acceptable utilisation of the resource.
Mine life.
18.2.12 Whittle Optimisation Results and Shell Selection
Optimisations were carried out on the basis of a process throughput rate of 15 Mtpa utilising Indicated Mineral Resources pertaining to Zone 1 and Zone 2. Additional analysis was undertaken on the impact of including Inferred Mineral Resources pertaining to Zone 1 and 2, and also to all four zones within the Mineral Resource model. Note that these assessments that contain Inferred Mineral Resources are preliminary in nature. They are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as Mineral Reserves. There is no certainty that this preliminary assessment would be realised.
However, the optimisations incorporating Indicated material only were used as the basis for selecting the Base Case shell on which subsequent pit designs and ore reserves were calculated.
Figure 18.2.12_1 displays the ore tonnes, waste tonnes and best case discounted cashflow for the Base Case optimisation. It is from this optimisation, derived upon the Indicated Mineral Resource only, that a suitable shell for the final pit design is selected. This is in line with both JORC Code and NI43-101 guidelines for generating ore reserves.
The waste tonnes increase constantly with increasing shell size. Project value is relatively flat, with less than 1% variation in best case discounted cashflow from Shell 25 to Shell 36.
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Figure 18.2.12_1 Base Optimisation Results for 15Mtpa Throughput Rates Ore Tonnes, Waste Tonnes and Best Case Discounted Cash Flow vs. WHITTLE Shell DFS Diluted Indicated Model - Base Case 1 800 $2 100 $2 000 Maximum Best Case Discounted $1 900 Cash Flow = $1990,4M Ore 1 600 $1 800 Tonnes = 202,7Mt $1 700 $1 600 $1 500 1 400 $1 400 $1 300 $1 200 1 200 $1 100 $1 000 $900 1 000 $800 $700 $600
Mtonnes 800 $500 $400 $300 $200 600 $100 Tonnes Ore $0 Total Tonnes -$100 400 Maximum Best Case Discounted Cash M$ Flow -$200 Best Case Discounted Cash Flow -$300 Worst Case Shell -$400 200 -$500 Average Case Shell -$600 -$700 0 -$800 1 3 5 7 9 11 13 15 17 19 21 23 25 27 29 31 33 35 37 39 41 43 45 Whittle Shell Number
It has been determined that the maximum best case discounted cashflow shell will be used as the basis for pit design for this study. Selection of the best case shell is acceptable on the following basis:
Figure 18.2.12_1 indicates there is little variation in cashflow as the shells approach the maximum best case. This indicates there is reduced risk to project value with a selected higher shell while simultaneously maximising the resource and associated mine life.
The sensitivity results indicate that the optimisation reaches an economic limit for much of the current Indicated resource. Any small beneficial variation in parameters will result in a very small incremental increase in shell size.
Previous optimisation and scheduling studies have indicated that generating a practical design and schedule results in cashflows that approaches the Best Case Whittle results. This is a result of:
The relatively robust nature of the deposit
The fact that Whittle is mining both zones at the same time when applying the Worst Case scheduling approaches as it treats them as one large pit. Consequently the Worst Case schedule generates significant upfront waste mining by mining Zone 2 in advance of what is required. A practical mining schedule can take advantage of the strip ratio benefits of the Zone 1 pit against the higher grade of the Zone 2 pit to improve project value.
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Therefore the shell selected as the basis for design and production scheduling is shell 36 derived from the Indicated Mineral Resource Only (Base Case) optimisation. A summary of Shell 36 is provided in Table 18.2.12_1 below.
Table 18.2.12_1 Husab Uranium Project Summary of Selected Shell 36
Material Whittle Cashflows Mill Worst Ore Waste Total Rec. U O Undisc. Best Case Strip Life 3 8 Case Ratio U O kTonnes 3 8 kTonnes Tonnes Mlb US$M Total (ppm) 202,674 500 1,383,322 1,585,996 192.1 US$5,350 US$1,208 US$1,990 6.83 13.9
It should be emphasised that the cashflow results generated by the optimisation work are only used for a comparative evaluation of various options and to assess the sensitivity of the project to variations in key parameters.
18.2.13 Sensitivity
A sensitivity analysis was carried out on the Base Case scenario. The objective of the sensitivity analysis is to assess the effect on the Base Case optimisation to changes in the key parameters denoted below.
Mining Cost (± 15%)
Processing / G&A Cost (± 15%)
Commodity Price (± 15%)
Process Recovery (± 10%)
Capital (± 15%)
Discount Rate (± 20%)
Wall Slopes (± 15%)
Inclusion of Inferred material.
Figure 18.2.13_1 indicates the percentage variation from the Base Case in terms of the ore tonnes contained within the Best Case optimal shell.
Figure 18.2.13_2 indicates the percentage variation from the Base Case in terms of the discounted cashflow for the Best Case optimal shell.
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Figure 18.2.13_1 % Variation from Base Case - Ore Tonnes
Figure 18.2.13_1 % Variation from Base Case - Best Case Discounted Cashflow
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The following Table 18.2.13_1 provides a guide to the ranking of sensitivities.
Table 18.2.13_1 Husab Uranium Project Sensitivity Ranking for a 15% Variation in Key Parameter
% Variation Description Ore Tonnes Best Case NPV Insensitive <1% <3% Slightly Sensitive 1% -5% 3% -10% Sensitive 5% -10% 10% -25% Highly Sensitive ≥10% ≥25%
The ore tonne variation indicates changes to the size and shape of the pit shell, and is therefore a reflection of the robustness of the ore body at a fundamental level. As variations in NPV are generally higher than ore tonnes, a greater variation in NPV is accepted before the project is considered “sensitive” and this is reflected in the rankings above.
On the basis of these rankings, Table 18.2.13_2 summarises the relative sensitivities of the optimisation. The sensitivities shown are for the Indicated ore only option as this optimisation is what the subsequent pit designs are based upon.
Table 18.2.13_2 Husab Uranium Project Sensitivity Summary
Parameter Ore Tonnes NPV Mining Cost Slightly Sensitive Sensitive Processing / G&A Cost Slightly Sensitive Slightly Sensitive Process Recovery Slightly Sensitive Highly Sensitive Price Slightly Sensitive Highly Sensitive Capital Insensitive Slightly Sensitive Discount Rate Insensitive Sensitive Wall Slope Slightly Sensitive Slightly Sensitive
18.2.14 Ore Tonnage Sensitivity
The sensitivity of ore tonnes ranges from insensitive to slightly sensitive against the defined parameters. A 15% variation in key parameters does not generate an ore tonnage variation greater than 5%. This demonstrates the comparative robustness of the shell, and therefore any associated pit design, to variations in these key project parameters. Consequently the mine life will not be sensitive to any reasonable changes in the project fundamentals.
The primary cause for this lack of sensitivity is the fact that there is very little marginal material within the deposit.
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The shells are reaching a natural economic limit in regard to costs. Any variations in cost result in a minimal impact on ore tonnes and overall size of pit. Similarly, if walls are steepened this does not release significantly more ore. The optimisation simply drives slightly deeper to gain better grade and uses the resulting reduced strip ratio to generate a higher value.
A global average economic cutoff grade of 140ppm has been calculated on the basis detailed in Table 18.2.13_3 below. This is expressed in terms of the model grade inclusive of ore loss and dilution (i.e. plant head grade).
Table 18.2.13_3 Husab Uranium Project Economic Cutoff Grade
Value Item Unit Zone 1 Zone 2 Processing Cost US$/t US$9.92 Incremental Ore Cost US$/t US$0.90 Total Cost of Ore US$/t US$10.82 Price US$/lb. US$65.00 Royalty % 3% Selling Cost US$/lb. US$2.50 Net Price US$/ppm 0.133 Residual Grade ppm 55.0 44.0 Plant Losses ppm 9.0 9.0 Total Losses ppm 64.0 53.0 145 134 Cutoff Grade (COG) ppm 140
Figure 18.2.13_3 is a conventional grade-tonnage curve for the Zone 1 / Zone 2 resource. If this figure is assessed on the basis of Table 18.2.13_3 above, it indicates that:
Only ~4% of the ore body lies beneath the economic cutoff grade
As an indication of the more marginal ore lying closer to the economic cutoff grade, only ~10% of the deposit lies between 140ppm and 200ppm.
This highlights that there is a limited amount of “marginal” ore and even less sub- economic mineralisation. Hence the optimisation tends to generate a fairly consistent shell that converts approximately 85% of the resource to an ore inventory within a shell, regardless of variations to the input parameters.
18.2.15 NPV Sensitivity on Best Case Shell
The project is highly sensitive to changes in uranium price and recovery, as these directly affect the bottom line. For a variation of 15% in either price or recovery, the project NPV varies by some 40%.
The variable recovery generated by a fixed tails residue grade results in lower grade ore being more adversely affected (on a proportional basis), than higher grade ore. Therefore head grade is also a key project parameter in the production scheduling process.
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Figure 18.2.13_3 Grade Tonnage Curve for Resource
18.2.16 Shell Geometry and the Effect of Inferred Mineralisation
Inclusion of Zone 1 and Zone 2 Inferred mineralisation adds considerably to the ore tonnes (approximately 24%). However, the strip ratio only increases by 2%, indicating that a considerable proportion of the Inferred material already lies within the optimal shell on the “Indicated Only” optimisation.
The inclusion of Inferred mineralisation in Zone 1 and Zone 2 pits increases NPV by 13%. This is despite an ore inventory increase of ~24%. This is because the additional ore generated extends the mine life beyond 15 years and as such is more heavily discounted. In addition the ore is subject to a higher incremental strip. Note that these assessments that contain Inferred Mineral Resources are preliminary in nature. They are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as Mineral Reserves. There is no certainty that this preliminary assessment would be realised.
To provide a graphical depiction between the Indicated Only and Indicated + Inferred optimisations, the following figures (Figure 18.2.16_1 to Figure 18.2.16_7) provide a range of cross-sections through the Zone 1 and Zone 2 shells.
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Figure 18.2.16_1 Zone 1 Cross-Section 7507000 N
Figure 18.2.16_2 Zone 1 Cross-Section 7506500
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Figure 18.2.16_3 Zone 1 Cross-Section 7506000
Figure 18.2.16_4 Zone 1 Cross-Section 7505300
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Figure 18.2.16_5 Zone 1 Cross-Section 7504000
Figure 18.2.16_6 Zone 1 Cross-Section 7503500
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Figure 18.2.16_7 Zone 1 Cross-Section 7503000
The above figures illustrate that the additional mineralisation captured by the Indicated and Inferred Mineral Resource shell includes:
Inferred mineralisation already within the Indicated Only optimal shell. Conversion of this material into Indicated classification would immediately increases the Shell 36 ore inventory without any increase in shell size. This additional material amounts to
approximately 16.4Mt at 440ppm U 3O8, or an additional 8% of ore. This is effectively transferring material previously classified as waste into ore which results in the strip ratio of Shell 36 decreasing by approximately 9%.
Inferred mineralisation outside of the Indicated Only shell. This equates to approximately
19.8Mt at 444ppm U3O8. This material is spread across the two zones at depth, with an easily identifiable portion to the north of Zone 1.
Indicated mineralisation outside of the Indicated Only shell. This mineralisation is determined to be sub-economic for the Indicated only optimisation, but when included with the Inferred mineralisation is then captured by the Best Case shell. Consequently
this mineralisation is of lower grade, being approximately 13.0Mt at 363ppm U 3O8.
These results clearly indicate the potential to increase the Whittle shell size, and therefore the associated reserve, by conversion of current Inferred resource to an Indicated category. This would increase the life of the mine in excess of 3 years, generating a total life (excluding pre- strip) of some 17 years.
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18.3 Pit Designs
The Whittle output identifies those areas that the optimisation process considers to be high value, which may be as a result of:
Maximising grade to the mill in the early years and / or
Deferring waste stripping as far as possible into the future.
The highest value areas initially developed by Whittle are on the north-western corner of Zone 1 pit (low strip) and the central section of Zone 2 pit (high grade). Therefore starter pits have been designed in these two areas and their impact on project economics assessed through comparative scheduling and life of mine (LOM) cashflow.
Interim and final pit designs need to ensure that the design parameters specified in Table 18.2.5_1and Table 18.2.5_2 are adhered to. In addition the mining access must be coherent and acceptable minimum mining widths be maintained. The following design considerations were followed:
The open pits are designed for the implementation of trolley-assist. This requires that ramps are kept as straight as possible for as long as possible, with an inside radius of curvature no less than 200m.
Dual access is established along the final limits of both pits though not in all associated stages. This is primarily to allow for emergency access from the pit in the event of wall failure, but also serves to reduce hauling distances from the respective pits to crusher, stockpiles or the waste rock dump.
The north western corner of Zone 1 has been trimmed back in order to remain inside the current lease boundary. The stand-off is approximately 15m, sufficient for a safety bund and light vehicle access.
The layouts of the Zone 1 and Zone 2 stages are shown in Figure 18.3_1 and Figure 18.3_2 respectively.
There is minimal waste stripping required at the northern end of Zone 1 pit. The depth to the top of fresh rock is on average 30m of free dig and calcrete materials. There is no ore within the sand and conglomerate layers.
A significant amount of waste stripping is required for the entire Zone 2 pit. On average 50m to 60m of free dig and calcrete materials are required to be removed to access the ore body. Similarly to Zone 1 pit, there is no ore within the free dig and calcrete layers.
Stage 3
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Figure 18.3_1 Zone 1 Stage Layouts
Stage 2
Stage 1
Stage 3
Stage 4
Figure 18.3_1 Zone 2 Stage Layouts
Stage 3
Stage 4
Stage 1
Stage 2
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Figure 18.3_3 below displays the final mine site layout including ultimate pits, MRF, stockpiles and ROM Pad.
Figure 18.3_3 Mine Site Layout
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18.3.1 Mineable Ore Reserve Estimates
This reserve has been generated in accordance with the guidelines developed by the Joint Ore Reserve Committee (JORC 2004). They have subsequently been reconciled to the standards prescribed by NI 43-101.
This provides the minimum standards for reserve reporting. The resource estimate includes indicated and inferred mineralisation. Only indicated mineralisation has been incorporated in the conversion from a resource to an ore reserve.
The Probable Mineral Reserves are based on Indicated Mineral Resources only and as such are available to be converted to Probable status.
The Probable Mineral Reserve estimate is presented in Table 18.3.1_1 and defines the final probable ore reserve for the project at 205.0Mt of ore at a diluted grade of 497ppm that results
in 224.8Mlb of U 308 at a strip ratio of 7.3:1.
Table 18.3.1_1 Husab Uranium Project Husab Probable Ore Reserve by Stage Designs
Ore 1 Waste Total Grade 2 U O Zone Stage Strip Ratio 3 8 (Mt) (Mt) (Mt) (ppm) (Mlb) 1 8.1 38.5 46.5 4.8 416 7.4 2 24.8 123.6 148.4 5.0 446 24.4 1 3 33.3 248.6 281.9 7.5 447 32.8 4 30.9 309.5 340.4 10.0 552 37.6 Total 97.1 720.1 817.2 7.4 477 102.2 1 11.8 90.3 102.1 7.7 715 18.6 2 45.7 250.5 296.2 5.5 438 44.1 2 3 43.4 342.5 385.9 7.9 546 52.2 4 7.0 96.1 103.1 13.8 495 7.6 Total 107.8 779.5 887.3 7.2 515 122.5 Grand Total 205.0 1,499.6 1,704.5 7.3 497 224.8 1. Based on a variable economic cutoff grade calculated block by block inclusive of variable process plant recovery. 2. Diluted head grade presented to mill.
18.3.2 Mining Fleet
The primary loading fleet comprises three large electric-powered rope shovels to mine the bulk waste on 15m benches whilst a fleet of smaller diesel hydraulic shovels address all ore loading requirements on 7.5m benches in order to minimise ore loss and dilution. The smaller shovels are able to manage both a 7.5m and 15m bench height. In–pit blending will minimise the extent of re-handling of ore from stockpile to crusher to cater for short-term grade variations over life of mine.
Large electric-powered drill rigs service the large electric loading units whilst the smaller and more manoeuvrable diesel rigs would similarly service the smaller diesel-powered loading units.
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Ore and waste are transported by a fleet of 39 diesel electric drive haul trucks in the +300 tonne class. Trolley-assisted hauling has been included in the base case, and will be implemented on most up-ramp sections of the open pits and ramps accessing the MRF.
The remainder of the mining production fleet consists of support equipment that includes graders, track and wheel dozers, front-end loaders, rock breakers and utility excavators.
Specific mining activities are planned to be outsourced. These include the repair and maintenance of the mobile mining fleet, blasting operations, tyre and haul road management as well as drilling operations relating to grade control and resource definition.
18.3.3 Compliance with Whittle
The pit designs developed upon the selected Whittle Shell 36 generally show a close correlation. The eastern pit limit of Zone 1 pit has been straightened to allow for the trolley ramp and the southern pit limit of Zone 2 pit has also been simplified due to the complex ramp layout required to maintain access. These modifications have increased the waste tonnes by 8.7% with an associate increase in ore tonnes by 0.8% that results in a net increase in strip ratio from 6.7 to 7.3. This is summarised by pit in Table 18.3.3_1.
Table 18.3.3_1 Husab Uranium Project Correlation between Pit Designs and Original Optimal Shell 36
Material Ore Description Zone Waste Total Strip U O Cont. U O (Mt) 3 8 3 8 (Mt) (Mt) Ratio (ppm) (Mlb) WHITTLE Base 1 95.3 482 101.3 647.4 742.7 6.8 1 Case Shell 36 2 109.2 517 124.4 732.5 841.7 6.7 Pit Optimisation Total 204.5 501 225.7 1,379.90 1,584.40 6.7 1 97.1 477 102.2 720.1 817.2 7.4 Mine Design 2 107.8 515 122.5 779.5 887.3 7.2 Pit Design Total 205.0 497 224.8 1,499.6 1,704.5 7.3 % Variation from Shell 0.2% -0.8% -0.4% 8.7% 7.6% 8.4% 1. Note MineSight was utilised to report the separate Whittle inventories from Zone 1 and Zone 2. There are minor reporting differences between MineSight and Whittle systems (<1%)
The generation of pit slope parameters for optimisation is somewhat iterative, as a number of assumptions need to be made in advance in regard to pit geometry and ramp layouts. The resulting design maybe considerably different in this respect if the pit geometry changes significantly or a revised access layout is adopted. Therefore to obtain a more definitive comparison between the final pit design and an optimal shell developed on the same basis, an additional optimisation run was completed with wall slopes based on the final design slopes. These results are outlined in Table 18.3.3_2 below and as expected show a closer reconciliation between the optimal shell generated and the final designs, with no significant differential on ore tonnage and grade.
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Table 18.3.3_2 Husab Uranium Project Correlation between Pit Designs and Revised Optimal Shell 36
Material Ore Description Zone Waste Total Strip U O Cont. U O (Mt) 3 8 3 8 (Mt) (Mt) Ratio (ppm) (Mlb) WHITTLE Base 1 95.3 482 101.3 666.6 761.9 7.0 1 Case Shell 36 2 108.8 517 124 782.1 890.9 7.2 Pit Optimisation Total 204.1 501 225.3 1,448.70 1,652.80 7.1 1 97.1 477 102.2 720.1 817.2 7.4 Mine Design 2 107.8 515 122.5 779.5 887.3 7.2 Pit Design Total 205.0 497 224.8 1,499.6 1,704.5 7.3 % Variation from Shell 0.4% -0.8% -0.2% 3.5% 3.1% 3.1% 1. Note MineSight was utilised to report the separate Whittle inventories from Zone 1 and Zone 2. There are minor reporting differences between MineSight and Whittle systems (<1%)
18.4 Reconciliation to Final DFS Parameters
Subsequent to the development of the pit design and associated reserves detailed above, a comprehensive open pit scheduling programme was undertaken to assess a number of different mine development strategies.
Scheduling was undertaken utilising evORElution, a highly sophisticated block by block scheduling tool developed in-house by ORElogy. This tool is characterised by its use of generic algorithms to generate a range of feasible schedules over a number of generations. Only the best schedules are selected as the basis for the subsequent generations, allowing the process to propagate towards an optimal or best fit solution to achieve the desired objectives.
evORElution develops schedules based on the following:
A set of user defined objectives such as material movement, value maximisation, etc.
On a block-by-block as opposed to a bench-by-bench scheduling approach. Consequently the software does not need to utilise bench averaging. Instead the true grades and true strip ratios are reported in any reporting period.
Produces a variety of solutions from a single simulation run. The solutions generated by evORElution are all valid but vary based on different trade-offs between various objectives. This gives the user:
A choice of solutions to examine.
A degree of sensitivity of each of the trade-offs.
An indication of potential paths for improvements in further scheduling efforts.
Any schedule generated by evORElution adheres to a number of imposed rules. Some of these rules are implied as, for example, a block can’t be mined if the block above is still in place. Other rules are user-imposed such as minimum mining width, frequency of excavator moves etc.
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The scheduling study evaluated over 50 schedules, each taking a differing approach to the goal of providing the highest project net present value within acceptable practical constraints. Increasing value is generally achieved by either bringing forward revenue or deferring costs. The schedules utilised combinations of the following strategies:
Deferring waste movement to reduce up-front mining costs.
Targeting areas of high grade to increase up-front revenue.
Apply an elevated cutoff grade (COG) to ensure maximum grade through the plant at any given time. Material between the elevated COG and the economic COG (refer to Table 18.2.13_3) is stockpiled and feed to the mill at the end of the mine life.
However, a further practical constraint that was applied as part of the scheduling process was the implementation of a mill head grade limit of 600ppm. The results of the scheduling exercise indicated that:
Utilising a “high strip/high grade” approach, which effectively focused all mining within Zone 2 initially, consistently generated the lowest project value.
The value of an elevated COG strategy increased very little above 250ppm, as the benefit was negated by the application of the 600ppm mill through put limitation.
The schedule finally selected as the DFS Base Case was referred to as the Zone 2 Balanced Hybrid option. “Z2 Balanced” referred to the pit release and material movement strategy. It indicated a balanced approach to LOM material movement and strip ratio with an emphasis on Zone 2 as an ore source wherever possible. “Hybrid” referred to the cutoff grade strategy, which consisted of a 250ppm elevated COG for all of Zone 1 and Stage 2 of Zone 2.
Once the Base Case had been selected a set of 3D surfaces were exported from evORElution for each of the schedule periods. These surfaces are reminiscent of a Whittle shell, and indicate the exact location of the mining face at the end of every period. From these shells detailed quarterly and annual mining face position plans were generated in 3D utilising the stage designs as the final guide. This was in order to confirm the practicality of the evORElution schedule, which in turn validated the interim and final pit designs developed.
The final DFS costs and operating parameters were all considerably reviewed and refined subsequent to the optimisation and design work being completed. Therefore it is valuable to reconcile the final parameters developed against those used in the optimisation process to determine if there is any significant divergence that may fundamentally alter the validity of the reserve. Table 18.4_1 details the variation in key parameters from the original optimisation to the final DFS value. The parameters detailed are the same as those assessed in the optimisation sensitivity analysis (refer to Section 18.2.13). This sensitivity analysis indicated that the WHITTLE shell is insensitive to variations in capital.
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Table 18.4_1 Husab Uranium Project Variation in Key Parameters Optimisation to Final DFS
% Var. from Effect of Parameter Unit Optimisation DFS Optimisation Shell Size Mining Cost $/ total tonne $1.58 $1.66 5% - Processing / G&A Cost $ / ore tonne $9.96 $14.07 41% - Net Commodity Price 1 $ / lb $60.55 $61.55 2% + Wall Slopes % N/A 2 N/A 2 3% 2 + 1. including transport costs, marketing costs, royalties etc. 2. wall slope changes vary by area and rock horizon and therefore cannot be expressed as a single variation. The 3% average variation is an estimate only and indicates a general steepening of the slopes.
An optimisation run was completed with the final DFS parameters applied. This resulted in a reduction of the ore inventory of the Best Case shell of ~5% to 192.5Mt.
However, it should be emphasised that considerable work is currently being undertaken as part of the company initiated Mine Optimisation and Resource Extension program (MORE) to investigate opportunities to continue to add additional project value. These initiatives include:
A resource update is planned for Q2, 2011 incorporating drilling completed until the end of January, 2011. The main focus has been infill drilling of Zone 1 and Zone 2 to define Measured Resources and increase the quantum of Indicated Mineral Resources, both by upgrading the classification of Inferred Mineral Resources within the current mine plan (and therefore not included within the reserve estimate) and by definition of additional resources.
Finer grind process: Potential to result in reduced leach residue grade which would result in increased recovery and simplified solid liquid separation circuit.
Elevated temperature acid leach: Potential to result in reduced leach residue grade which would result in increased recovery.
Direct IX or SX to replace Eluex: Potential to lower plant capital expenditure.
On-site acid production: Procurement of sulphuric acid represents a significant component of the Project’s operating cost. The possible upside from on-site production of acid, including the potential to generate electricity and provide heat for an elevated temperature leach process, is being assessed.
Work on these value adding areas is planned to continue through the project development phase and, when appropriate, feature in an updated ore reserve and associated mine plan.
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19 ENERGY, WATER AND PROCESS MATERIALS
19.1 Water Supply
The supply of water to the Husab mine has been identified as a very crucial aspect for both construction and permanent operations. Existing fresh water resources in the area do not have adequate capacity to supply the projected regional demand and desalination of seawater is considered to be the only viable solution for permanent water supply.
Water demand is expected to be 1.2Mm³ p.a. during the development phase and 6.5Mm³ p.a. at steady state. During development, water will be supplied through a temporary water supply pipeline to be constructed from the NamWater reservoir near the Rössing Mine. During operation, water is expected to be sourced from the proposed desalination plant at Mile 6, to be constructed and operated by third parties.
The company is a member of the Erongo Mining Water Users Group (EMWUG) which is working with the National Desalination Task Force (NDTF) to investigate and implement a strategy to deliver water to the project in line with the envisaged development timetable. After recommendations were tabled to the Namibian Cabinet in February 2011, a Public-Private- Partnership (PPP) has been approved for the structuring of a new desalination plant. However, to date there is no commitment from any party to build a desalination plant, and the company continues to assess potential temporary or fall back solutions in line with the project’s development timetable.
Figure 19.1_1 Water Supply to Husab Mine
Pipeline Legends:
Permanent Supply for Existing Pipelines Temporary Construction
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19.2 Power Supply
The electrical power utility provider in Namibia, NamPower, commands sufficient generated and/or imported energy sources to provide both the temporary and permanent power needs for the Husab mine. Capital, to be supplied by Swakop Uranium, is however required to fund the extensions to the transmission network that runs parallel to the B2 highway between Windhoek and Swakopmund. Referred to as the “western corridor” this transmission network extension requires inter alia a dual 220kV transmission line branch, and a new 220kV NamPower sub-station at the Husab mine to enable the mine to be included in the 220kV supply ring. The solutions for both temporary and permanent power supply have been developed around this arterial supply route and fits well with the timeline of the Husab mine’s development and production schedule, as follows:
Temporary power will be supplied through a 220/66kV temporary switchyard situated near the Arandis airport and transmitted across the Khan valley to the sub-station at Husab mine. Temporary power will be transmitted at 66kV for the bulk of the period leading up to availability of permanent power.
Permanent power will be supplied at 220kV from a new NamPower Main Transmission Substation situated on the west side of the Khan river. Permanent power is scheduled to be available immediately prior to process plant product commissioning.
At the Husab 220/66/11kV customer supplied sub-station permanent power made available from NamPower at 220kV will step down to 66kV and 11kV and be reticulated at 66kV in an overhead line ring main around the mining pits where skid mounted, 66/33kV and 33/6.6kV step down transformers will be used to supply electrical power to the haul truck trolley assist and the mining fleet of electrically powered shovels and drills.
An 11kV sub-station will distribute electrical power to plant, admin, crushing and mining complexes.
The electrical power supply configuration, cost and timeline have been agreed with NamPower during the DFS phase. Implementation of the NamPower temporary and permanent supply schemes is awaiting project approval.
Discussions with NamPower have taken place and are ongoing.
Figure 19.2_1 shows the proposed modified transmission network for the 220kV supply to Husab mine.
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Figure 19.2_1 Modified 220kV Supply
19.3 Project Supply Chain And Logistics
Process reagents are expected to be sourced from around the world The bulk of the logistic movements, including distribution of end product, are expected to be routed through the Port of Walvis Bay. Regional supplies, particularly from South Africa are expected to be directed through Windhoek on road carted consignments.
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Reagents suppled to the process plant of the Husab mine are listed in Table 19.3_1.
Table 19.3_1 Husab Uranium Project Reagent Supply Quantities
Commodity Units Annual Consumption Delivery to Site Grinding Media t 6,000 Pallets in container Sulphuric Acid t 396,930 Road tanker Lime t 14,080 Road tanker Pyrolusite t 43,846 Bulk Bags in container Flocculant t 2,318 Bulk Bags in container Coagulant t 214 1m³ IBC U IX Resin m³ 353 Bulk Bags in container Extractant t 5.61 Drum (on pallet in container) Diluent t 405 Road tanker Modifier m³ 5.45 Drum (on pallet in container) Sodium Carbonate t 3,126 Bulk tanker Sodium Hydroxide t 18,250 Bulk tanker Hydrogen Peroxide t 1,380 25.5t Isotainer Drums each 28,432 Truck Pallets each 7,108 Truck Water treatment Chemicals t 41.4 Pallets in container
For the purposes of the DFS, the supply chain and logistics for reagents is assumed to be the responsibility of the Vendors concerned. Quoted prices represent a delivered to site figure.
The bulk of logistics movements to site are made up of:
Sulphuric acid supply to site – 1200 tons per day
Fuel - over 100,000 litres per day
Personnel movements - reported in Section 20.4 of this report.
The majority of the reagents can be supplied to site in 20 ft containers which will be handled and carted with standard equipment using ridge stackers and flat bed trucks. This is aligned with the storage philosophy on site where the hard stand area provides stacking and storage space for containers.
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20 INFRASTRUCTURE
20.1 Access Roads
A preferred access route was determined on the basis of the following criteria:
Distance to travel – setting an objective of having the shortest possible routes between Swakopmund and Arandis to Husab
Cost of private road – setting the objective of making maximum use of existing public roads in order to limit the cost of a private road
Environmental – being sensitive to parameters that influence the environmental rating of the selection e.g. footprint through sensitive area; entry point into NNNP; minimising dust, and vehicle emissions.
A number of trade-off studies were carried out in order to select a preferred access route to Husab. One such study compared the cost of transporting goods to the mine site by rail with the cost of transport by road. The marginally lower operating costs of rail transport, compared with road transport, were shown to be insufficient to outweigh the capital investment required. Consequently the project has focussed on construction of a black top asphalt road to site.
The planned access road runs from an existing railway siding on the B2 highway, across the Khan valley to join the Old German Railway (OGR) route where it crosses the Khan River. The new private road leaves the river bed at Paddaklip, and follows the same OGR route up the southern traverse to the proposed gatehouse of Husab mine, which is located at the old Welwitsch station site. The new private access road is approximately 22km long and represents a total travelling distance of 65km between Swakopmund and Husab mine. This will be the shortest travel distance from Swakopmund of all existing uranium mine operations in the Erongo region and can be used as an incentive when recruiting permanent staff. Figure 20.1_1 shows the alternative access routes considered.
20.2 Port Facilities
The supply of reagents from international suppliers for the operational phase will be predominantly by ship to Walvis Bay harbour, and then by road to the Husab site . Reagent suppliers will remain responsible for the supply of reagents to the mine site, and will thus make use of their existing storage facilities.
The reagent supplies will be supplied in 20ft containers to standardise the logistical movement and material handling facilities as much as possible. However, reagents such as sulphuric acid, caustic lye, sodium carbonate, lime and diluents will be delivered by bulk road tanker. Pyrolusite from Morocco will be bulk bagged at source and shipped by bulk carrier to the port of Walvis Bay. The pyrolusite will be transported to site by flat-bed truck. Smaller quantities of reagents will be supplied containerised bulk bags, drums or IBCs. Hydrogen peroxide will be supplied in 25.5t isotainers
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Figure 20.1_1 Alternative Routes Considered
Option No. Route Name Existing Infrastructure en route
1 OGR Existing tar road (B2) from Swakop to Goanikontes junction 2 OGR (Alt) Existing tar road (B2) from Swakop to new junction 3 Aurecon (West) Existing tar road (B2) from Swakop to Goanikontes junction 4 Aurecon (East) Existing tar road (B2) from Swakop to Goanikontes junction 5 Khan Mine Existing tar road (B2) from Swakop to Rössing airstrip junction Existing tar road (B2) from Swakop to Valencia junction. Section of new 6 Valencia gravel road to Valencia mine(base quality?) Existing main gravel road (C28) from Swakop to Welwitchia plains junction 7 Welwitchia (incl. 17km of secondary tar section). Existing park gravel road from C28 junction to point close to large Welwitchia
20.3 Housing
Swakop Uranium will adopt a policy of encouraging private ownership of housing by its employees.
The urban and township development potential of Swakopmund has been investigated as part of the DFS. An early survey of this situation confirmed the assumptions on which the housing policy was developed.
20.4 Personnel Transport
The operational phase of Swakop Uranium’s Husab mine will embrace a policy of transporting employees to work. An estimated 1200 employees and contractors will be engaged in operations and will require transport from Swakopmund, Walvis Bay and Arandis to the Husab mine.
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Bus routes in the towns will be established for pick-up and drop-off of employees. It is envisaged that the majority of shift workers will be resident in Arandis, while the majority of daytime workers will reside in Swakopmund and Walvis Bay.
Employee logistic movements are expected to require a fleet of 10 35 seat buses each travelling approximately 700km/day on asphalt roads, with a further two buses available on stand by.
It is envisaged that the mine and processing plant will operate 24 hours a day, employing three shifts of eight hours each. In addition a daytime shift will be worked on a five day week basis starting at 08H00, and finishing at 16H00.
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21 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT
21.1 Location
The Husab Project lies some 60km east of the coastal town of Swakopmund, approximately 5km south of the Rössing Uranium Mine in the north of the Namib Naukluft National Park (NNNP) and in an area of high and unique bio-diversity. Apart from mining, tourism is an important industry for the area. The NNNP, the harsh and stark beauty of the Swakop and Khan Rivers, the Big Welwitschia and Welwitschia fields feature on the tourist itinerary. The Husab Project is situated southwest of the Khan River and immediately north of the Welwitschia fields. The closest towns to the site are Arandis 18km to the north and Swakopmund to the west.
21.2 Legislation Governing Uranium Mining In Namibia
The Republic of Namibia has five tiers of law and a number of policies relevant to Uranium mining. Key policies currently in force include The Minerals Policy of Namibia (2002), Namibia’s Environmental Assessment Policy for Sustainable Development and Environmental Conservation (1995) and the Atomic Energy and Radiation Protection Act (2005).
The applicable laws and policies for the Husab Project are provided below.
Namibia’s Environmental Assessment Sustainable Development and Environmental Conservation Policy of 1995 requires that all listed policies, programmes and projects be subject to an EIA;
Environmental Management Act, No 7 of 2007. The regulations are still to be promulgated to give this Act effect. Until this occurs, all environmental impact assessments are guided by Namibia's Environmental Assessment Policy for Sustainable Development and Environmental Conservation of 1995;
The Environmental Investment Fund of Namibia (not yet enforced) makes provision for fines for environmental offenders;
The Minerals (Prospecting & Mining) Act, No 33 of 1992 regulates reconnaissance, prospecting and mining of minerals;
The Water Act, No. 54 of 1956, inherited from South Africa, provides for the control, conservation and use of water; being surface water, sea water and ground water;
Forest Act, No 12 of 2001 makes provision for the protection of various species of plants;
Nature Conservation Ordinance, No 4 of 1975 establishes a guiding framework for habitat and species conservation, including wildlife management and utilisation in National Parks;
Hazardous Substances Ordinance 14 of 1974. This ordinance provides for the safe handling, storage and disposal of hazardous substances;
The National Heritage Act, No 27 of 2004 provides for the protection and conservation of places and objects of heritage significance and the registration of such places and objects;
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Atomic Energy and Radiation Protection Act (5 of 2005) provides the mechanism for obtaining authorization related to activities involving possible exposure to ionizing radiation;
Namib Naukluft National Park management and tourism development plan;
The Namibia Water Corporation Act, 12 of 1997 enables the supply of bulk water so long as the required quantity and quality of water is available;
Policy for Prospecting and Mining in Protected Areas and National Monuments, 1999, aims to promote sustainable development by guiding prospecting and mining in protected areas and national monuments.
Swakop Uranium has worked within the relevant laws and legislations and the relevant applications for permits have been, and will be, applied for as need arises.
For the Husab Mine and its infrastructure development, future legislation developments that may be important are:
The regulations for the Environmental Management Act may require provision of, and guarantees for, funds for the closure and rehabilitation of the mine and infrastructure;
In addition, environmental officers of the Ministry of Environment and Tourism (MET) will have the right to inspect a site and issues fines if they deem the project to be non- compliant in an area;
Parks and Wildlife Management Bill (in prep) which aims to provide a legal framework for maintenance of ecosystems, essential ecological processes and the biological diversity of Namibia. The Bill allows MET and Ministry Mines and Energy (MME) to agree to withdraw certain areas within parks from mining. Apart from these “no go” areas, mining within parks would only be permitted with written authorization from the Minister of MET;
The NNNP management and tourism development plan states that no development should result in the decline of more than 10% in the population of a species of special interest.
Two other requirements under which the Husab Project will have to operate are the Namib Naukluft National Park rules and the Strategic Environmental Assessment (SEA) for the central Namib uranium rush and its subsequent Strategic Environmental Management Plan (SEMP) developed to limit the cumulative impact of the anticipated development of several uranium mines and associated industries.
21.3 The Impact Assessment Process
Metago Environmental Engineers (Pty) Ltd (Metago) was the independent firm of consultants appointed by Swakop Uranium to undertake the EIA and related processes for both the Husab Mine EIA and the associated Linear Infrastructure EIA. Several local independent scientists and specialists were also sub-contracted by Metago to undertake specific studies for the EIA assessments.
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The final EIAs and the EMP were reviewed by an independent body, the Southern African Institute of Environmental Assessors (SAIEA), prior to their submission to the MET.
Full consultation took place throughout the EIA process. Government, the regional authorities, and local industries as well as the public have been consulted, and given opportunities to provide input to the project.
The EIAs and EMP have been prepared to meet the requirements of the Equator Principles.
The EIA and EMP for the Husab Project, processing plant, and mine residue facility were submitted in November 2010 to the MET. The Environmental Clearance Certificate was received on 25 January 2011.
The EIA for the Linear Infrastructure (roads, power, communications and water) will be submitted to the authorities in early June 2011 together with the updated EMP. It is anticipated that the Environmental Clearance Certificate will be secured before the end of July 2011.
Swakop Uranium has to implement and manage the identified commitments and procedures contained in the EMP document.
21.4 An Overview of The Project Area
The area in which the Husab Mine and its related infrastructure are to be developed is a virtually pristine Namibian ecosystem that is mostly untouched by any anthropogenic developments except for the old German railway, and some tourism and exploration activities within the NNNP, and tourism, power and water lines, quarrying and mining to the north of the Khan River.
The greater project area is important for the central Namib biodiversity because it is situated in a triangular area boarded by two significant rivers (the Khan River to the northwest and the Swakop River to the south) that only flow periodically as a result of the development of dams upstream. The rivers form a linear oasis that cross declining fog and increasing rain isohyets as one progresses upstream. It is believed that these rivers and associated valleys allow water and nutrients to reach into the desert from the wetter hinterland and for fog to reach further into the desert from the coast than would be expected.
Given that water and nutrients are key ecological drivers in the desert, this is an important and unique aspect that has resulted in the varied habitats, floral and faunal life in the area. The scarcity of water, an annual average rainfall of less than 50mm per annum, and evaporation rate of over 3,000mm per annum, the prevailing coastal winds and hot desiccating winter “berg” winds, and relatively warm daytime temperatures has resulted in specially adapted fauna and flora resulting in many endemics with specific roles in the eco-system. The contrasting habitats associated with the plains, the river valleys and the transitional zones in- between provides a range for many taxa of conservation importance some of which are endangered, data deficient, vulnerable, near threatened and/or protected.
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21.5 Major Issues of Concern That Were Identified and Recommended Mitigation Measures
The findings of the Husab Mine and Linear Infrastructure Impact assessments have highlighted several areas of concern regarding the biodiversity and socio-economics of the project area. These are outlined below, together with the recommended mitigation measures that could ameliorate the negative impact:
Ridges and rocky outcrops, particularly limestone outcrops, harbour a number of plant endemics such as protected Lithops sp and Aloe sp. Infrastructure will avoid these areas as far as is possible and affected plants will be rescued;
The soils and desert pavement are sensitive and disturbance thereof can lead to rapid erosion by wind and consequent loss of vegetation, nutrients and the seed bank. The area of disturbance will be kept as small as is possible and rehabilitation will take place following disturbance;
The prediction from the air dispersion modelling is that particulate matter less than ten micron in size (PM10) is a material concern. In this regard, the incremental impacts are generally of low significance, but because the existing baseline PM10 concentrations for the region already exceed the evaluation criteria used, the cumulative impacts have a high significance at the closest third party receptor points (Arandis, Rössing Uranium Mine and the big Welwitschia tourist site) in both the unmitigated and mitigated scenario.
The Husab Sand Lizard, Pedioplanis husabensis, is thought to be in the centre of its restricted range in the Husab area. Studies have been conducted on its local range and habitat, and further scientific work is to be done to determine its red data status;
One of the largest fields of the iconic and protected Welwitschia mirabilis occur on EPL3138. New deposits are being discovered under these fields. A working group of scientists of various disciplines has been formed to undertake specific work on the plants to determine their recruitment, longevity and population dynamics;
As the project progresses, it is becoming apparent that near surface water flows may be a significant contributory factor providing water to the Welwitschias. The open pits and mine residue facility may impact the hydrology of this system. Investigations in this regard are being developed and implemented;
The ephemeral channels that flow predominantly south-westwards across the site bring fresh water to the Welwitschia field and harbour a number of the larger plants found in the area. Infrastructure is being kept off them as far as is possible, and the near surface flow in the channels is being investigated;
There are several springs in the tributaries to the Khan River. Springs can be small ecosystems in themselves and are an important water source for the large animals in the region (Oryx, Zebra, Kudu etc);
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The movement of large animals and birds between water sources and grazing areas will be hampered or impeded by the construction of the mine and its related infrastructure, and by the power, water and road routes through narrow valleys;
The Hartmann’s Mountain Zebra is classified as vulnerable and may be impacted upon as a result of large and linear infrastructure blocking pathways to springs and the Khan River. The Husab Project will assist with studies to determine the impact and recommended mitigation measures;
Several protected Acacia eriolobas (Camel thorn) will be affected by the permanent access road to site. Permits for their destruction will be obtained;
River crossings through both the Kahn and Swakop River may impede the free movement of animals along the rivers, and the power line will be a hazard to birds. Mitigation measures will be implemented;
The impacts on the tourism are not as great as originally anticipated although the noise levels in the area will increase, lighting will be visible from great distances and there will be increased traffic along major routes. The sense of “wilderness” will be lost from certain areas. Alternate tourism spots will need to be identified and other sites upgraded, in conjunction with the NNNP;
The embankments of an old narrow gauge German rail line run across a part of the Husab Mine site and are protected. Remnants of the embankment will be affected by the permanent access road and the mine. The site of the Welwitschia station at the top of valley leading from the Khan River has virtually disappeared, but it has been suggested that it be resurrected as a tourism site;
Negative social impacts of the proposed Husab Mine include, amongst others, increased pressure on local amenities such as schools, medical facilities and the potential increase in the spread of pandemic diseases. The project will, as part of responsible corporate citizenship, contribute to the relevant areas needing support.
The Husab Mine will bring a number of economic benefits to the Erongo region and to Namibia in general. Suffice to say that an estimated 8% of the national GDP can be generated by this one mine alone.
21.6 Conclusion
In summary, there are a number of predicted negative impacts of the Husab Mine and related infrastructure on the bio-physical and socio-economic environments. However, the financial benefits of the mine to the local, regional and Namibian economy will be great. There will be people who oppose the project on the grounds of potential negative environmental and social impacts, while others would support the project on the grounds of potential positive economic impacts. Swakop Uranium must meet the commitments contained in the environmental management plans to reduce the overall impact on the environment.
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22 MARKET STUDIES AND CONTRACTS
22.1 Demand for Uranium
The principal use of uranium is as a fuel for nuclear reactors. In 2009, total installed nuclear
capacity was 373 GWe and demand for uranium was estimated to be around 174Mlbs U3O8.
Installed capacity is forecast to increase significantly as a result of new build programmes around the world, as several developing economies seek to increase the share of nuclear power in their energy balance to the levels of more developed countries.The main drivers of nuclear power are the increase in global energy demand, reduced tolerance for greenhouse gas emissions and improvements in nuclear power technology and economics. Prior to the incident at the Fukushima Daiichi nuclear station in Japan on March 11 , 2011, nuclear capacity in the upper scenario was forecast to more than double by 2030 (WNA, 2009).
Variable costs, including fuel, represent a relatively small part of the overall cost of generation from nuclear power, and nuclear power stations tend to provide baseload generation with little variation in power output. Consequently, demand for uranium is relatively inelastic and is expected to keep pace with the expansion in nuclear capacity. Prior to the incident at Fukushima, reactor requirements were forecast to increase to 277Mlbs per year in 2020 and 364Mlbs per year by 2030.
Immediately after the Fukushima incident several countries, including China, called for a review of safety standards of existing and planned nuclear plants which could have an impact on the future development of nuclear programmes.
22.2 Marketing of Uranium
Mining companies generally produce and sell uranium as uranium oxide (either U 3O8 or UO 4), which is typically delivered to a conversion facility for conversion to uranium hexafluoride
(UF 6) prior to enrichment to increase the proportion of fissile U 235 before fabrication for use as nuclear fuel.
Most uranium oxide is traded through bilateral agreements between suppliers (principally mining companies) and users (principally nuclear power generators). There is no terminal market for uranium oxide; spot and term prices are published by a small number of independent companies based on information gained from market participants. The spot market price is the most widely quoted, although in recent years spot market transactions have accounted for only approximately 20% of the total market. Conversion, enrichment and fabrication services are typically purchased separately by the end user.
The uranium oxide market is expected to tighten in the medium term owing to the growth in global nuclear capacity, the decline in production from existing mines, the reduction in available secondary supplies and delays to the development of new supply as a result of a more difficult financing environment.
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Analysts’ medium to long term forecasts for uranium oxide prices range between $70 and
$85/lb. Extract has for evaluation purposes assumed a flat real price of $65/lb U 3O8 for the life of mine.
Extract has engaged with potential customers to assess demand for production from the Husab Uranium Project, and has identified several possible strategic contracting opportunities. Extract is confident that it will become an attractive supplier to end-users, as a result of the Husab Uranium Project’s ability to offer geographic diversification and long term security of supply. As at the date of this report, the Project had not entered into any commitments for the sale of uranium.
22.3 Taxation
An overview of the fiscal system in Namibia, outlining the principal taxes and duties expected to be payable by the project, is provided below. Taxation of the parent company, and/or individual investors is not considered in this overview.
The rate of corporate income tax payable by mining companies is 37.5%, payable on taxable profits. Capital allowances on machinery, equipment and vehicles, may be taken on a straight line basis over three years. Exploration and development costs may be deducted in the year that mining commences.
A royalty of 3.0% of gross sales is expected to be applicable.
Value Added tax (VAT) may be chargeable on sales and paid on purchases within Namibia. Where applicable, the VAT rate is 15%, although certain items are zero rated for VAT. It is expected that uranium produced by the Husab Uranium Project will be exported, and will therefore not be subject to VAT within Namibia. Namibia is a member of the Southern Africa Customs Union (SACU) and the Southern African Development Community (SADC). No duties are levied on intra SACU trade. Customs duties paid according to the Common Customs Tariff of SACU on imports outside SACU. Import VAT is payable on importation of goods into Namibia at various rates dependent on the item and its source. No VAT is chargeable on services imported. VAT paid is expected to be claimed back from the Namibian VAT authorities as input tax through monthly VAT returns.
Withholding taxes are payable on some payments to institutions in certain jurisdictions. Dividend payments are subject to a withholding tax of 10%, with lower rates applied to payments to certain treaty countries. No withholding tax is payable on interest payments made to companies overseas. Withholding taxes on services may be payable.
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23 CAPITAL AND OPERATING COSTS
23.1 Cost Estimates
The capital and operating cost estimates are presented in US dollars assuming an exchange rate of US$1 = ZAR7.5. Estimates are presented in real terms with a base date of January 2011 and are considered to have an accuracy of ±10%.
23.1.1 Capital Cost Estimate
Based on the currently envisaged mine fleet, base case process design and the currently planned strategy for infrastructure provision, capital costs are estimated to be approximately US$1,480 million (see Table 23.1.1_1).
Table 23.1.1_1 Husab Uranium Project Estimated Capital Cost
Item US$ Million Initial Mine Fleet & Infrastructure 407 Processing plant 529 Waste & Tailings Facility 71 Infrastructure & Temporary facilities 210 Indirect costs (EPCM, Owners costs, other) 158 Contingency 105 1,480
Initial mine fleet and infrastructure costs include the initial mine and ancillary fleet and equipment, initial infrastructure for trolley assist, mining complex and on-site roads and provision of initial spares. Processing plant costs include the cost of construction of the processing plant and on-site infrastructure. Infrastructure costs include the Project’s expected contribution towards the cost of a shared water pipeline, and the full cost of a smaller dedicated pipeline to site from this shared pipeline. It is anticipated that the costs of construction of a new desalination facility will be borne by a third party. Infrastructure costs include allowance for provision of temporary and permanent electrical supply. Temporary facilities include the cost of a construction camp to be built in close proximity to the Husab site.
Inclusive of pre-strip and other capital and operating costs totalling US$179 million incurred prior to commissioning and during ramp-up, the project cost is estimated at US$1,659 million. The Project Cost excludes allowance for finished goods inventory in transit and held at conversion facilities, debtor payment terms, creditor payment terms, escalation, and financing costs (including fees and interest during construction).
Over the project life, ongoing capital expenditure, including sustaining capital and progressive installation of the trolley assist infrastructure, is estimated to total approximately US$415 million. End-of-life environmental rehabilitation costs estimated at $32 million are assumed to be largely offset by equipment salvage value of $31 million.
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23.1.2 Operating Cost Estimate
The steady state cost of production is estimated at US$28.5/lb, excluding royalties, marketing and transport, and excluding cost escalation (see Table 23.1.2). Operating costs including royalties, marketing and transport are estimated at US$32.0/lb.
Table 23.1.2_1 Husab Uranium Project Estimated Operating Costs
Capital US$ / lb Mining 13.9 Processing (1) 13.4 G&A 1.2 Cost of Production 28.5 Royalty 2.0 Transport & Marketing 1.5 Operating Cost 32.0 Notes (1) a) The process plant operating costs are estimated at US$12.40/lb U3O8 equivalent. An additional cost of US$1.00 has been added to include the operating cost of the mine residue facility (MRF). b) The manpower costs only include process plant operating personnel. All other labour categories are included in the general and administrative (G & A) costs. c) Waste disposal is included in the G & A costs. d) Fuel costs used were the prevailing prices at the time the estimate was produced. e) Cost for mill liners, crusher liner, filter cloth, lubricants, etc. were included with the vendor quotes for the individual packages and these costs have been capitalised for the first two years of operation. f) Reagent consumables including water – the operating cost model takes in account these costs on a year by year basis.
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24 ECONOMIC ANALYSIS
24.1 Taxation
An overview of the fiscal system in Namibia, outlining the principal taxes and duties expected to be payable by the project, is provided below. Taxation of the parent company, and/or individual investors is not considered in this overview.
The rate of corporate income tax payable by mining companies is 37.5%, payable on taxable profits. Capital allowances on machinery, equipment and vehicles, may be taken on a straight line basis over three years. Exploration and development costs may be deducted in the year that mining commences.
A royalty of 3.0% of revenue is expected to be applicable.
Value Added tax (VAT) may be chargeable on sales and paid on purchases within Namibia. Where applicable, the VAT rate is 15%, although certain items are zero rated for VAT. It is expected that uranium produced by the Husab Uranium Project will be exported, and will therefore not be subject to VAT within Namibia. Namibia is a member of the Southern Africa Customs Union (SACU) and the Southern African Development Community (SADC). No duties are levied on intra SACU trade. Customs duties are paid according to the Common Customs Tariff of SACU on imports outside SACU. Import VAT is payable on importation of goods into Namibia at various rates dependent on the item and its source. No VAT is chargeable on services imported. VAT paid is expected to be claimed back from the Namibian VAT authorities as input tax through monthly VAT returns.
Withholding taxes are payable on some payments to institutions in certain jurisdictions. Dividend payments are subject to a withholding tax of 10%, with lower rates applied to payments to certain treaty countries. No withholding tax is payable on interest payments made to companies overseas. Withholding taxes on services may be payable.
24.2 Economic Analysis – Results
Based on the Probable Mineral Reserves defined to date, the project is forecast to process
205Mt ore and to produce 198Mlbs U3O8 equivalent over the 16 year life of mine (from start of mining to end of processing).
A cashflow forecast is shown in Table 24.2_1, assuming a uranium price of US$65/lb. The Project annual cashflow model indicates a cash outflow of US$1,628 million prior to the start of commissioning. Additional cash outflow is expected in the initial months of the first year of operation, giving rise to an estimated total project cost of US$1,659 million. The net present value of forecast post tax cashflows, discounted at 8% per year, is $822 million based on existing reserves.
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The payback period, assuming a uranium price of US$65/lb, is estimated to be 6.8 years from project start, or 4.1 years from commissioning. The value of the project, as calculated using a discounted future cashflow analysis based on proven and probable reserves defined to date, is shown in Table 24.2_2 below for various market price and discount rate assumptions.
Estimates exclude allowance for finished goods inventory in transit and held at conversion facilities, debtor payment terms, creditor payment terms, escalation, and financing costs (including fees and interest during construction).
24.3 Economic Analysis - Sensitivity
The sensitivity of the valuation to certain parameters, based on existing reserves, is shown in Table 24.3_1 below.
Extract Resources has commenced a programme to investigate opportunities to add significant additional value through the MORE programme. Opportunities to extend the economic life of the project through definition of additional reserves are described below. In addition, several potential process enhancements are being investigated:
Finer grind process: Potential to result in reduced leach residue grade which would result in increased recovery and simplified solid liquid separation circuit.
Elevated temperature acid leach: Potential to result in reduced leach residue grade which would result in increased recovery.
Direct IX or SX to replace Eluex: Potential to lower plant capital expenditure.
On-site acid production: Procurement of sulphuric acid represents a significant component of the Project’s operating cost. The possible upside from on-site production of acid, including the potential to generate electricity and provide heat for an elevated temperature leach process, is being assessed.
Work on these value adding areas is planned to continue and to feature, as appropriate, in the final mine plan to be initiated during the project development phase.
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Table 24.2_1 Husab Uranium Project Forecast Cashflows (Reserves only)
. Total / Year Unit Average -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 Total Mined Mt 1,705 - 9.0 78.8 93.1 123.4 140.6 158.4 168.1 162.0 141.3 126.4 129.7 128.2 133.5 74.7 25.1 12.2 - Waste Mined Mt 1,500 - 9.0 75.9 82.5 107.3 123.1 138.5 147.9 144.8 124.2 109.1 115.7 113.1 118.3 63.2 18.7 8.4 - Ore Mined Mt 205 - - 3.0 10.6 16.1 17.5 19.9 20.2 17.2 17.1 17.3 14.0 15.2 15.2 11.6 6.4 3.8 - Grade Mined ppm 497 - - 324 564 549 422 433 439 482 452 462 568 590 462 567 677 660 - Net Stockpile Movement Mt - - 3.0 0.6 1.5 2.6 4.9 5.2 2.2 2.1 2.3 (1.0) 0.2 0.2 (3.4) (8.6) (11.2) (0.4) Ore Processed Mt 205 - - - 10.0 14.6 15.0 15.0 15.0 15.0 15.0 15.0 15.0 15.0 15.0 15.0 15.0 15.0 0.4 Grade Processed ppm 497 - - - 559 578 503 505 508 530 485 473 550 556 506 521 404 316 187 Recovery % 88.0% - - - 84.9% 89.8% 87.7% 89.1% 88.4% 88.9% 87.8% 88.1% 90.0% 89.9% 88.3% 88.2% 85.0% 81.0% 79.9%
Uranium Production Mlbs U 3O8 equiv 198 - - - 10.5 16.7 14.6 14.9 14.9 15.6 14.1 13.8 16.4 16.5 14.8 15.2 11.4 8.5 0.1 Uranium Price $/lb 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 65.0 Revenue $m 12,852 - - - 668 1,084 947 966 966 1,013 916 895 1,065 1,075 961 987 738 551 21 - Royalty $m (386) - - - (20) (33) (28) (29) (29) (30) (27) (27) (32) (32) (29) (30) (22) (17) (1) - Transport & Marketing $m (304) - - - (16) (26) (22) (23) (23) (24) (22) (21) (25) (25) (23) (23) (17) (13) (1) Net Revenue $m 12,162 - - - 632 1,026 896 915 914 959 867 847 1,008 1,017 909 934 698 521 20 - Mine Operating Costs $m (2,711) - - - (132) (198) (214) (225) (245) (235) (238) (212) (230) (224) (228) (164) (83) (73) (9) - Plant Operating Costs $m (2,654) - - - (133) (185) (193) (197) (194) (196) (195) (192) (197) (199) (195) (195) (191) (187) (5) - G&A Operating Costs $m (215) - - - (16) (15) (15) (15) (15) (15) (15) (15) (15) (15) (15) (15) (15) (15) (5) Total Operating Costs $m (5,579) - - - (280) (398) (422) (437) (454) (446) (448) (419) (442) (438) (438) (374) (289) (276) (19) EBITDA $m 6,583 - - - 352 628 474 478 460 512 419 429 565 579 471 560 409 245 1 Tax Payments $m (1,721) ------(46) (161) (182) (147) (149) (203) (210) (174) (207) (151) (91) (0) Initial Capex $m (1,480) (155) (637) (689) ------Capitalised Pre-Production Open $m (148) (3) (32) (113) ------Sustaining / Enhancement Capex $m (429) - - - (107) (96) (41) (29) (26) (30) (23) (41) (9) (10) (5) (8) (2) (0) (0) Cashflow $m 2,804 (157) (669) (802) 244 533 433 403 274 300 248 238 353 360 292 344 255 154 1
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Table 24.2_2 Husab Uranium Project Project NPV (Reserves only)
Discount Rate (post-tax, real) NPV (US$m) 6% 8% 10% 60 800 522 302 65 1,153 822 560 U3O8 Price (US$/lb) 70 1,502 1,119 814 75 1,850 1,414 1,067
Table 24.3_1 Husab Uranium Project Valuation Sensitivity
Change In Parameter Sensitivity Project Cost Op. Cost NPV Payback Period ($m) ($/lb) (8%) (years) + $5 / lb (1.4) +0.2 +297 (0.7) Uranium Price - $5 / lb +3.8 (0.1) (300) +0.8 + 10ppm (0.7) (0.6) +82 (0.1) Grade - 10ppm +0.7 +0.6 (82) +0.2 + 10% +20.1 +2.9 (189) +0.5 Operating Cost - 10% (18.4) (2.9) +188 (0.3)
24.4 Mine Life Extension
Extract Resources believes that there is significant potential to extend the current 16 year life of the mine through definition of additional reserves.
A resource update is planned for Q2, 2011 incorporating drilling completed until the end of January, 2011. The main focus has been infill drilling of Zone 1 and Zone 2 to define Measured Mineral Resources and increase the quantum of Indicated Mineral Resources, both by upgrading the classification of Inferred Mineral Resources within the current mine plan (and therefore not included within the reserve estimate) and by definition of additional resources.
Since the August, 2010 resource update Extract has completed an additional 375 resource definition drillholes at Zones 1 and 2 which have added 118,392 metres of drilling to the resource database. The programme is continuing with up to 6 rigs on site targeting a further 100,000 metres drilling over the next 12 months.
Additionally, open pit optimisation of the updated resource is expected to result in definition of additional reserves and a lower strip ratio, following detailed geotechnical review which indicated potential for an increase in slope angles.
Extract Resources intends to continue its exploration programme in Zones 3, 4 and 5, Middle Dome, Salem, Ida Dome, and Pizzaro areas, which are not included in the DFS.
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25 INTERPRETATION AND CONCLUSIONS
The Husab Uranium Project area contains one of the largest uranium deposits in the world with significant uranium mineralisation associated with uraniferous leucocratic granites (alaskites) within the highly prospective Central Zone of the Damara Orogeny. The mineralised alaskites tend to occur along or proximal to the unconformity contact between the Khan Formation and Rössing Formation.
The August 2010 resource (Table 1.4_1; and 17.1.9_2 & 17.2.5_4) represents a significant increase in Indicated Mineral Resources relative to the previous July 2009 Mineral Resources for Zones 1 and 2, and now incorporates maiden resources for Zones 3 and 4.
Potential remains to expand the Mineral Resource inventory at the Husab Uranium Project through extension of known deposits such as Zones 1, 2, 3 and 4 and definition of resources at prospects such as Zone 5, Middle Dome and Salem.
Coffey Mining has reviewed the drilling, sampling, assaying and field procedures used by Extract and consider them to be of high quality.
Further bulk density information is required in the mineralised portions of the deposit and of the overburden. Additional infill and extensional drilling is required to raise the level of confidence of the Indicated and Inferred Resources.
Extract has defined a base case mine plan and process plant design, including plans for delivery of the infrastructure necessary to support the project. The DFS has demonstrated the technical and economic viability of developing Husab into one of the largest uranium mines in the world.
The DFS is based on:
Indicated Mineral Resources defined at Zones 1 and 2;
Open pit mining by truck and shovel from two separate pits to maintain a sustained rate of 15Mt pa over the life of mine with an average strip ratio of 7:1 (waste:ore);
A waste and plant tailings storage facility (the mine residue facility);
Ore crushing and overland conveying to a new processing facility employing milling, leaching, ion exchange, solvent extraction and precipitation plant and equipment to produce approximately 15 million lbs pa of U3O8 equivalent; and
Provision of temporary and permanent power and water supplies, access roads, temporary and permanent buildings and structures necessary to support the Project.
Husab Uranium Project, Namibia – MINEWPER00713AE Page: 187 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
Capital costs for the Project are estimated at US$1,480 million, including initial mine fleet, process plant and supporting infrastructure. Inclusive of pre-strip and other pre-production operating costs of US$179 million, the Project Cost is estimated at US$1,659 million. This estimate excludes allowance for finished goods inventory in transit and held at conversion facilities, debtor payment terms, creditor payment terms, escalation, and financing costs (including fees and interest during construction).
Production costs are estimated at US$28.5/lb, excluding royalties, marketing and transport and cost escalation. Operating costs including royalties, marketing and transport are estimated at US$32.0/lb.
Husab Uranium Project, Namibia – MINEWPER00713AE Page: 188 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
26 RECOMMENDATIONS
The DFS defines a base case mine plan and process plant design, including plans for delivery of the infrastructure necessary to support the project. Several opportunities to add further value have been identified, including the proposed update of the resource model, mine plan optimization, and processing enhancements. Extract Resources has commenced a MORE programme to investigate these opportunities.
A resource update is planned for Q2, 2011 incorporating drilling completed until the end of January, 2011. The main focus has been infill drilling of Zone 1 and Zone 2 to define Measured Mineral Resources and increase the quantum of Indicated Mineral Resources, both by upgrading the classification of Inferred Mineral Resources within the current mine plan (and therefore not included within the reserve estimate) and by definition of additional resources.
Since the August, 2010 resource update Extract has completed an additional 375 resource definition drillholes at Zones 1 and 2 which have added 118,392 metres of drilling to the resource database. The programme is continuing with up to 6 rigs on site targeting completion of approximately 100,000 metres of drilling over the next 12 months.
Closer spaced drilling of the Husab deposit has revealed an increasingly complex relationship of mineralisation and alaskite lithologies. It is recommended that in future a probabilistic, Multiple Indicator Kriging (MIK) resource model be considered, to generate a recoverable SMU model over Zones 1 and 2.
Additionally, open pit optimisation of the updated resource (Q2, 2011) is expected to result in definition of additional reserves and a lower strip ratio, following detailed geotechnical review which indicated potential for an increase in slope angles.
Once the Q2, 2011 Mineral Resource update has been completed Extract should continue with further infill drilling at Zones 1 and 2 with the aim of defining additional Measured Resources. The expectation would be to complete another resource update (H1 2012) for Zones 1 and 2 to take through to the commencement of mining. The 2012 updated resource should be subject to detailed optimisation and mine planning such that an updated mine plan and Mineral Reserve is expected to be completed during the second half of 2012.
Extract also intends to continue its exploration programme in Zones 3, 4 and 5, Middle Dome, Salem, Ida Dome, and Pizzaro areas, which are not included in the DFS. Definition of additional reserves would be expected to add additional value and mine life to the project.
Husab Uranium Project, Namibia – MINEWPER00713AE Page: 189 43-101 Technical Report – 20th May 2011 Coffey Mining Pty Ltd
Several potential process enhancements are being investigated and include the following:
Finer grind process: Potential to result in reduced leach residue grade which would result in increased recovery and simplified solid liquid separation circuit.
Elevated temperature acid leach: Potential to result in reduced leach residue grade which would result in increased recovery.
Direct IX or SX to replace Eluex: Potential to lower plant capital expenditure.
On-site acid production: Procurement of sulphuric acid represents a significant component of the Project’s operating cost. The possible upside from on-site production of acid, including the potential to generate electricity and provide heat for an elevated temperature leach process, is being assessed.
Work on these value adding areas is planned to continue and to feature, as appropriate, in the final mine plan to be initiated during the project development phase. Table 27_1 summarises an estimated budget for the above recommendations.
Table 27_1 Husab Uranium Project Estimated Budget for Recommendations for 2011/2012
Item/Activity Cost (US$) Resource Estimation and Mining Studies $1.5M Additional Resource Drilling – Zones 1 & 2 $10.0M Additional - Husab Uranium Project exploration $5.0M Engineering/Pre-development $17.0M Total $33.5M
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27 REFERENCES
Anglo American Prospecting Services Namibia (Pty) Limited. 1982. Prospecting Grant M46/3/444 Husab, South West Africa/Namibia. Annual Prospecting Report for the year ended 31st December 1981.
Anglo American Prospecting Services (Pty) Ltd - Swakop Exploration (Pty) Limited. 1980. Husab Uranium Venture. Prospecting Grant M46/3/444 South West Africa/Namibia. Annual Prospecting Report for the year ended 31st December 1979.
Bevam, J. 2011. Report on Taxation and Economic Analysis for the Husab Uranium Project. Internal Company Report. Extract Resources.
Bothe, H.W., 1980. Prospecting Grant M46/3/487 Welwitschia Uranium Joint Venture. Renewal Report August 1980. Anglo American Prospecting Company, 4pp.
Culpan, N., 2008. Comparison of downhole gamma spectrometer logging with chemical assays. Internal Company Memorandum, Extract Resource Limited. 5 December 2008.
Culpan, N., 2009. Rössing South Densities. Internal Company Memorandum. 5 January 2009.
Courtellier, G.R., 1975. Report on geological, geochemical and geophysical investigations on Prospecting Grant known as “Eintracht” M46/3/606. Aquitaine SWA unpublished report, 8pp.
Davies, S. 2011. Report on uranium Markets and Contracts. Internal Company Report. Extract Resources.
Dorrington, R. 2008. Agreed Restructure – Extract Resources Limited and Kalahari Minerals plc. ASX Media Release. Dated 5 September 2008. Obtained from http://www.asx.com.au/asx/research/CompanyInfoSearchResults.jsp?searchBy=asxCo de&allinfo=on&asxCode=EXT#headlines
Freemantle, G. June 2009. Preliminary Report to Extract Resources (Swakop Uranium) of mineralogical results from QEMSCAN analyses of selected surface and core samples of the Rössing South Deposits. Unpublished Technical Report.
Freemantle, G. September 2010. Petrographic, geochemical and mineralogical assessments of Rössing South alaskitic and metasedimentary rocks from NQ core samples. Unpublished Technical Report.
Freyer, EE, Badenhorst, FP and Krupp, KP. 1991. Renewal Report, Grant 1771, North Namib-Naukluft Park. Anglo American Prospecting Services unpublished report.
Hill, MH. October 2010. Rössing South Metallurgical Testwork Report – Final. Unpublished Technical Report.
Inwood, N. 2009, National Instrument 43-101 Technical Report, Rössing South – January 2009 Resource Update. Unpublished technical report
Inwood, N. 2009, National Instrument 43-101 Technical Report, Rössing South – January 2009 Resource Update. Unpublished technical report.
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Jayasekera, S., Turner, J., van der Meulen, D. October 2010. Rössing South Uranium Project Integrated Pilot Plant Testwork Final Report. SGS Lakefield Oretest. Unpublished Technical Report.
Lambert,I., McKay, A., and Miezitis, Y. 1996 Australia's uranium resources: trends, global comparisons and new developments, Bureau of Resource Sciences, Canberra, with their later paper: Australia's Uranium Resources and Production in a World Context, ANA Conference October 2001.
Lobo-Guerrero, A., Granitoid-related iron-oxide-copper-gold Mineralisation, greater Lufilian Arc, Zambia and Namibia. In, Muhling, R., Goldfarb, N., Vielreicher, F., Bierlein, F., Stumpfl, E., Groves, D.I., Kenworthy, S. (Eds) Extended Abstracts, SEG 2004, Predictive Mineral Discovery Under Cover.
Mokwena, A., Sadiki, M., Kalala, J.T. November 2010. Pilot Semiautogenous Milling Testwork on Rössing South run of mine sample. Mintek. Unpublished Technical Report.
Morel, V., 2007. Husab JV and Uis Uranium Projects, Central Western Namibia. Technical Report by RSG Global Consulting for Extract Resources Limited.
Ministry of Mines and Energy Namibia. Legislation on minerals (Prospecting and Mining) Act. 1992 (Art No 33 of 1992).
Mitiri, J. June 2009. UltraSort testwork report. Unpublished technical report.
Penkethman, A. & Spivey, M. 2008. Information Document – Namibia Uranium Project. May 2008.
Prince, K.E. and Kelly, I.J. May 2009. Technical Memorandum: AM/TM2009_14_05, Rössing South Ore Mineralogy. Unpublished Technical Report. ANSTO Minerals.
Prince, K.E. and Kelly, I.J. June 2009. Technical Memorandum: AM/TM2009_12_06, Rössing South Mineralogy <38µm alaskite ore and leach residue. Unpublished Technical Report. ANSTO Minerals.
Prince, K.E. and Kelly, I.J. August 2009. Technical Memorandum: AM/TM2009_20_08, Rössing South + 53 m and – 53 m Fractions. Unpublished Technical Report. ANSTO Minerals.
Prince, K.E. and Kelly, I.J. November 2009. Technical Memorandum: AM/TM2009_23_11, Mineralogy of select Extract <10µm leach residues. Unpublished Technical Report. ANSTO Minerals.
Prince, K.E. and Kelly, I.J. November 2009. Technical Memorandum: AM/TM2009_12_11, Rössing South ore variability study. Unpublished Technical Report. ANSTO Minerals.
Rössing Uranium Ltd, 1982. Interim Report on Exclusive Prospecting Grant M46/3/1229- Farms Hatsamas and Coas, Windhoek District. Unpublished company report, 14pp.
Speiser, A. 2005. Environmental Assessment and Management Plan (Phase 1 and 2) for West Africa Gold Exploration (Namibia) Pty Ltd – Exploration Activities on EPL 3138.
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Spivey, M. 2006. Information Document – Namibia Uranium Project. December 2006.
Spivey, M. 2006. West Africa Gold Exploration (Namibia) (Pty) Ltd. Report on Exploration EPL3138 “Husab” for the period ending 19.04.2007. Volume 1, Text and Plans. Report No. TR3138(1) Submitted to the Ministry of Mines and Energy, December 2006.
Spivey, M and Penkethman, A. 2009. Discovery of the granite (alaskite) hosted Rössing South uranium deposit in Namibia; an exploration perspective. Conference Presentation Abstract, Roundup 2009.
Stobart, B. 2011. Environmental Studies, Permitting and Social or Community Impact. Internal Company Report. Metago Environmental Engineers.
Swakop Exploration (Pty) Limited. 1974. Prospecting Grant M46/3/444 over area known as Husab Area, Registration “G” South West Africa. Annual Prospecting Report for 1973.
Swakop Exploration (Pty) Limited. 1975. Prospecting Grant M46/3/444 over area known as Husab Area, Registration “G” South West Africa. Annual Prospectus Report for 1974.
Swakop Exploration (Pty) Limited. 1976. Prospecting Grant M46/3/444 over area known as Husab Area, Registration “G” South West Africa. Annual Prospectus Report for 1975.
Swakop Exploration (Pty) Limited. 1978. Husab Uranium Joint Venture. Prospecting Grant M46/3/444 South West Africa. Annual Prospecting Report for 1977.
Swakop Exploration (Pty) Limited. 1979. Husab Uranium Joint Venture. Prospecting Grant M46/3/444 South West Africa. Annual Prospecting Report for 1978.
Swakop Exploration (Pty) Limited. 1981. Husab Uranium Joint Venture. Prospecting Grant M46/3/444 South West Africa/Namibia. Annual Prospecting Report for the year ended 31st December 1980.
Schneider, GIC and Seeger, KG 1993. The mineral resources of Namibia. Geological Survey of Namibia, Ministry of Mines and Energy, Republic of Namibia.
Townend, R. August 2008. Our reference 22352. SEM examination of three crushed rock samples, after screening at 106 microns, for uranium minerals. Unpublished technical report.
Townend, R. November 2009. Our reference 22600. Preparation of one polished thin section of one drill core and examination (optical/SEM) for uranium minerals. Concentration of heavy minerals from part crushed drill core in TBE liquid and XRD identification (RDD 23 362m). Unpublished technical report.
Townend, R. November 2009. Our reference 22611. SEM examination of polished sections of four Feed and four Leach residue samples for uranium minerals. Unpublished technical report.
Townend, R. June 2010. Our reference 22737. SEM/BSE/EDS examination of 10 uranium ore tails with reference to uranium minerals. Unpublished technical report.
The CIA World Fact Book, website www.cia.gov. 2010.
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Wilson, P., 2008. Rössing South Assay Quality Control Report, December 2009. Internal Company Report. Extract Resources Limited. December 2008.
World Nuclear Association, 2009. The Global Nuclear Fuel Market, Supply and Demand 2009-2030.
www.bannermanresources.com. 2011 .
www.paladinresources.com. 2011 .
www.Rössing.com 2010 .
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28 DATE AND SIGNATURE PAGE
The “Qualified Persons” (within the meaning of NI 43-101) for the purposes of this report are shown below. The effective date of this report is 20 May 2011.
[Signed] Neil Inwood MSc, MAusIMM Principal Resource Geologist Coffey Mining Pty Ltd. Signed on the 20 May 2011
[Signed] Steve Le Brun M AusIMM Principle Consultant Coffey Mining Pty Ltd. Signed on the 20 May 2011
[Signed] Steve Craig AusIMM Managing Director ORElogy Signed on the 20 May 2011
[Signed] Ross Cheyne M AusIMM Director ORElogy Signed on the 20 May 2011
[Signed] Mike Valenta M SAIMM Managing Director Metallicon Signed on the 20 May 2011
[Signed] Hugh Browner F SAIMM Engineering Manager AMEC Minproc Signed on the 20 May 2011
[Signed] Steve Amos F SAIMM Technical Manager AMEC Minproc Signed on the 20 May 2011
Husab Uranium Project, Namibia – MINEWPER00713AE Page: 195 43-101 Technical Report – 20th May 2011
Appendix 1
QAQC Report
Rössing South Assay Quality Control Report July 2010
P. D. Wilson
Rössing South Assay QC Report July 2010
1. Introduction
1.1. Data Reviewed
The quality control data reviewed in this report comprises all field and laboratory QC results for the Rössing South project received in the period 15/7/2009 to 7/7/2010.For QC results in the period prior to 15/7/2009 refer to “Rossing South December 2008 Assay QC Report” and “Rossing South July 2009 Assay QC Report” by the same author. The report includes data from the Zone 1 and Zone 2 prospects. All primary analytical work was carried out by Genalysis Laboratory Services Pty Ltd. Sample preparation was undertaken at their facility in Johannesburg and the pulps were analysed in Perth. A single analytical method was used by Genalysis: AT/MS (four-acid digest in Teflon tube, ICP-MS finish).
Ultra Trace Geoanalytical Laboratories was selected as an umpire laboratory using Ultra Trace code ICP304 (peroxide fusion, ICP-MS finish).
Table 1.1 compares the numbers of each QC control type inserted by Extract Resources with the total number of sample results received during the period of this report.
Table 1.1 Summary of QC Sampling
QC Control Count % Standards 4870 2.5 Blanks 5206 2.7 Duplicates 4656 2.4 Umpire Repeats 910 0.5 Total QC 15642 8 All samples 196084 100
1.2. Field Procedures
Reverse Circulation
Quality control samples in the form of certified reference standards, field duplicates and blank samples were inserted into the sample stream as part of the normal sample sequence. Most recently the three QC sample types were inserted in the ratio of 1:33, giving an overall ratio of close to 1:10.
Sample numbers ending in “00”, “30” and “70” were allocated to standards. Sample numbers ending in “36”, “66” and “76” were allocated to duplicates of the previous samples. Sample numbers ending in “10”, “40” and “80” were allocated to blanks. In addition extra blanks, up to a maximum of 5 per hole, were targeted on high grade intervals as indicated by a hand-held spectrometer. These were placed 2 samples after the high grade sample because the laboratory uses two mills and processes alternate samples through each. Contamination from a high grade sample should be apparent in the next sample through the mill, which would be the second sample after the high grade one.
20 July 2009 1 Rössing South Assay QC Report July 2010
Diamond Drilling
QC procedures for diamond drilling were similar to RC with the exceptions that no field duplicates were inserted and standards were inserted in the ratio of 1:40, giving an overall ratio of 1:36.
1.3. Charting
The plots used in this report were produced using QAQCReporter version 1.10.3. Standard control charts plot an “Instance” number along the X-axis. On charts of this type each result is separated by an equal distance on the X-axis and the instance number represents the order of results sorted by Report Date, Batch No and SampleID . While there is no guarantee that the laboratory has processed all samples in batch and sample number order on a given day, the X-axis is an approximation of the order of processing through the laboratory.
2. Field Standards and Blanks
Extract Resources used 14 African Mineral Standards (AMIS) of various uranium concentrations in the Rössing South drilling program. The standards used and their expected values by ICP analysis are listed in Table 2.1. The minimum and maximum expected values are two standard deviations from the mean, established by round-robin analysis among multiple laboratories. Certificates detailing the certification procedure are available from http://www.amis.co.za/
Table 2.1. AMIS standards used by Extract Resources
U ppm by ICP Standard Expected Min Max AMIS0046 93 84.5 101.5 AMIS0054 1410 1314 1506 AMIS0076 1502 1336 1668 AMIS0085 263 242 284 AMIS0086 127 117.5 136.5 AMIS0090 890 831 949 AMIS0097 527 497 557 AMIS0098 819 723 915 AMIS0100 1474 1398 1550 AMIS0113 456 411 501 AMIS0131 294 268 320 AMIS0133* 3323 3157 3489 AMIS0154 682 621 743 AMIS0156 195 161 229 * The expected value for AMIS0133 has been determined by XRF only
The results for the standards analysis are shown in Table 2.2. Individual control charts are included in Appendix 1. Outliers, many of which when investigated, are the result of misidentification of standards in the field submissions have been excluded from the summary statistics in Table 2.2 and the plots in Appendix 1.
20 July 2009 2 Rössing South Assay QC Report July 2010
Table 2.2. Statistics for Standards Submitted by Extract Resources, analysed by ICP
Genalysis Perth Standard AMIS0046 AMIS0054 AMIS0076 AMIS0085 AMIS0086 AMIS0090 AMIS0097 Expected Value (EV) 93 1410 1502 263 127 890 527 U/U3O8??ppm Expected Min 84.5 1314 1336 242 117.5 831 497 Expected Max 101.5 1506 1668 284 136.5 949 557 Count 384 237 145 497 477 48 506 Minimum 83.9 1287.1 1388.2 204.7 96.4 801.0 460.8 Maximum 122.6 1545.3 1641.6 305.4 146.0 951.3 592.7 Mean 99.1 1426.5 1528.8 264.3 129.6 888.9 527.5 Std Deviation 3.083 44.639 45.736 9.863 4.997 29.851 18.702 % in Tolerance 75.5 95.4 100 95.4 90.4 91.7 89.5 % Bias 6.6 1.2 1.8 0.5 2.1 -0.1 0.1
Genalysis Perth Standard AMIS0098 AMIS0100 AMIS0113 AMIS0131 AMIS0133 AMIS0154 AMIS0156 Expected Value (EV) 819 1474 456 294 3323 682 195 Expected Min 723 1398 411 268 3157 621 161 Expected Max 915 1550 501 320 3489 743 229 Count 644 408 358 501 609 24 32 Minimum 739.6 1278.0 396.2 262.6 2905.3 686.2 183.7 Maximum 916.2 1647.8 520.9 367.3 3692.2 747.3 239.1 Mean 836.4 1471.3 446.6 296.3 3230.3 723.4 210.1 Std Deviation 26.199 44.969 16.712 9.548 119.561 14.879 8.084 % in Tolerance 99.8 91.4 98 97.2 72.1 83.3 96.9 % Bias 2.1 -0.2 -2.1 0.8 -2.8 6.1 7.7
The mean values for each standard show a bias from the expected value ranging from -2.8% to +7.7%.
• The bias on AMIS0156 at 7.7% is at the high end but is based on 32 values.
• AMIS0154 is similar and shows a bias of 6.1% based on only 24 values. • AMIS0046 shows a high positive bias of 6.6% from a count of 384 values. Because the expected value of 93 ppm is at the low end of the grade range this represents a departure from the expected value of 6 ppm.
• The remaining standards all report bias within +/- 3%.
20 July 2009 3 Rössing South Assay QC Report July 2010
Blanks
Table 2.4 lists the statistics for blanks inserted by Extract Resources. Some variability is evident in the control chart for field blanks. Where these have been investigated they invariably are associated with high grade sections of the drill holes that contain them, suggesting that the cause is low level contamination between samples. Performance of the blanks was queried with the lab in May 2009, prior to the period covered by this report. Performance subsequently improved and has continued to improve following these investigations.
More than 99% of blanks reported within acceptable limits.
Table 2.4. Statistics for Extract Resources Field Blanks
Field Blanks ICP Expected Value (EV) 2.5 Expected Min 0 Expected Max 10 Count 5206 Outliers 2 Minimum BDL Maximum 71 Mean 2.08 Std Deviation 11.7 % in Tolerance 99
3. Laboratory Standards and Blanks
Genalysis inserted a range of 6 standards and 2 blanks into the sample stream to monitor internal QC and provided Extract with the expected values and upper and lower limits for these standards. Genalysis used some of the same AMIS standards as Extract and these have been separately identified in the database by the use of a “_LAB” suffix in the standard ID. The expected minimum and maximum values listed in Table 3.1 are those provided by Genalysis and are not the same as the tolerances given in the certificates for the AMIS standards.
The statistics for each of the lab’s internal standards are listed in Table 3.1 and individual control charts are included in Appendix 2. As expected, the internal QC of the laboratory is very good with nearly all results within the limits accepted by the lab, and few outliers. Bias is slightly positive across the range of ore grade standards. OREAS 45P shows the highest bias of 6.5 percent, but this standard is effectively a blank and can be disregarded.
20 July 2009 4 Rössing South Assay QC Report July 2010
Table 3.1. Statistics for Standards Inserted by Genalysis (Note “_LAB” extension removed from AMIS standard IDs to save space in this table)
Genalysis Perth Standard CUP-1 DH-1a OREAS 45P UTS-1 AMIS0004 AMIS0086 Expected Value (EV) 1280 2600 2.4 49 88 127 Expected Min 1088 2200 2.04 41.65 74.8 117.5 Expected Max 1472 3000 2.76 56.35 101.2 136.5 Count 1349 1324 1319 1289 1325 367 Minimum 1154 2268 0.02 33.8 63.5 110.9 Maximum 1499.7 2875.5 29.6 104 103.2 143.7 Mean 1299.8 2571.7 2.5 49.3 88.6 128.8 Std Deviation 43.28 90.7 1.1 7.4 3.585 4.629 % in Tolerance 99 100 84 79 99 95 % Bias 1.42 -1.09 6.522 0.516 0.715 1.4
The statistics for acid and control blanks are also very good. Table 3.2 gives a summary of the statistics. Table 3.2. Blanks Inserted by Genalysis
Standard Acid Blank Control Blank Expected Value (EV) 0 0 Expected Min -0.5 -0.5 Expected Max 0.5 0.5 Count 720 2323 Minimum 0 0 Maximum 5 1.1 Mean 0.03 0.04 Std Deviation 0.176 0.069 % in Tolerance 99.9 99.9
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4. Duplicates
4.1. Field Duplicates
Extract Resources inserted field duplicates in a ratio of 3 per 100 samples in their RC drilling program. In November 2009, part way through the period of this report, the method by which duplicates were collected was modified on the recommendation of Neil Inwood of Coffey Mining. The two methods are coded as RC_SPLIT (old method) and RC_DUP (new method). In the original procedure the duplicate was collected as a split of the subsample that was to be dispatched to the lab and so was a duplicate of the lab sample only. In the new procedure the entire sampling process is replicated in order to produce the duplicate.
A total of 2267 pairs of results were produced by the old method and 2389 by the new. The results for the two methods are plotted separately.
Scatter plots show good precision for the data set with most results falling within the 20% precision limit. A number of the outliers shown are strongly suspected to be the result of recording errors in the field where standards and blanks have been placed in positions where duplicates should have been collected and which have been recorded as duplicates. Figure 4.1.1 is a scatter plot showing the full range of grade values. Figure 4.1.2 and 4.1.3 show more detail of the data in the grade range 0 – 1200 ppm U and 0 – 500 ppm U respectively.
20 July 2009 6 Rössing South Assay QC Report July 2010
Figure 4.1.1a Scatter Plot Field Duplicates by Old Method - All data
Figure 4.1.1b Scatter Plot Field Duplicates by New Method - All data
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Figure 4.1.2a Scatter Plot Field Duplicates by Old Method - 0 to 1200 ppm
Figure 4.1.2b Scatter Plot Field Duplicates by New Method - 0 to 1200 ppm
20 July 2009 8 Rössing South Assay QC Report July 2010
Figure 4.1.3a Scatter Plot Field Duplicates by Old Method - 0 to 500 ppm
Figure 4.1.3b Scatter Plot Field Duplicates by New Method - 0 to 500 ppm
20 July 2009 9 Rössing South Assay QC Report July 2010
Quantile-Quantile plots of the same data show a small positive bias in the range 300 – 900 ppm for results produced by the old procedure, but in the plots for the new procedure this bias is absent or slightly negative. However, this appears to be a result of low numbers of duplicate pairs in this range and the presence of a few outliers shifting the means, rather than any discernible difference caused by the change in sampling method.
Figure 4.1.3a Quantile-Quantile Plot Field Duplicates by Old Method - All data
Figure 4.1.3b Quantile-Quantile Plot Field Duplicates by New Method - All data
20 July 2009 10 Rössing South Assay QC Report July 2010
Figure 4.1.4a Quantile-Quantile Plot Field Duplicates by Old Method - 0 to 1200 ppm U
Figure 4.1.4b Quantile-Quantile Plot Field Duplicates by New Method - 0 to 1200 ppm U
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Figure 4.1.5a Quantile-Quantile Plot Field Duplicates by Old Method - 0 to 500 ppm U
Figure 4.1.5b Quantile-Quantile Plot Field Duplicates by New Method - 0 to 500 ppm U
20 July 2009 12 Rössing South Assay QC Report July 2010
A measure of the precision of duplicate pairs is the Mean Absolute Paired Difference (%MAPD). This value is often referred to as %HARD (Half Absolute Relative Difference) or %RPHD (Relative Paired Half Difference).
For a duplicate pair A and B %MAPD is defined as:
%MAPD = Absolute of [(A-B)/Average(A:B)*100*0.5]
Figure 4.1.5 is a plot of %MAPD against percentile rank for field duplicates. Grades below 10 ppm U have been excluded from the plot because of the high relative differences exhibited by duplicate pairs of very low grades. The data were plotted for three grade ranges: all data, 0 – 1000 ppm U and 0 – 500 ppm U. In each case 95% of the data fell within the 20% precision limits for field duplicates, regardless of the method. Only the plot showing all data is included here.
Figure 4.1.5a %MAPD Plot Field duplicates by Old Method
Figure 4.1.5b %MAPD Plot Field duplicates by New Method
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4.2. Lab Pulp Checks
Genalysis performed a total of 2946 pulp checks as part of their internal QC and the results show very good precision. A tighter precision limit of 10% is used for pulp checks because they are not subject to the same variability as field duplicates through the sampling and preparation process. The %MAPD plot (Figure 4.2.5) shows that 97% of the checks fall within this limit. Note that, as for the field duplicates, only results for grades greater than 10 ppm have been included in the plot.
Figure 4.2.1 Scatter Plot Lab Pulp Checks- All data
Figure 4.2.2 Scatter Plot Lab Pulp Checks- 0 to 1200 ppm
20 July 2009 14 Rössing South Assay QC Report July 2010
Figure 4.2.2 Scatter Plot Lab Pulp Checks- 0 to 500 ppm
Figure 4.2.3 Quantile-Quantile Plot Lab Pulp Checks- All data
Figure 4.2.4 Quantile-Quantile Plot Lab Pulp Checks- 0 to 1200 ppm
20 July 2009 15 Rössing South Assay QC Report July 2010
Figure 4.2.5 Quantile-Quantile Plot Lab Pulp Checks- 0 to 1200 ppm
Figure 4.2.5 %MAPD Plot Lab Pulp Checks
20 July 2009 16 Rössing South Assay QC Report July 2010
5. Umpire Laboratory Results
A total of 910 pulps recovered from Genalysis were sent to Ultra Trace for re-analysis. Samples were manually selected from the Zone 1 and Zone 2 prospects at Rössing South. The number of samples represented approximately 10% of the +75ppm U3O8 samples analysed in the period 15/7/09 to 7/7/10. Samples were selected to cover the depth, strike and cross-strike extents of the Zone 1 and 2 mineralisation and the selection of samples was drawn from each of the section lines across the prospect in groups of roughly 10 consecutive samples.
Ultra Trace used an hydroxide fusion digest followed by ICP-MS and the original analyses by Genalysis were done with a 4-acid digest ICP-MS.
A comparison of the umpire repeat results for drill samples is shown by the series of plots below. As with previous umpire results, reported in “Rossing South August 2009 Umpire Results” and “Rossing South July 2009 Assay QC Report” by the same author, the Ultratrace fusion method has a higher yield of U compared to the 4-acid digest used at Genalysis. It is particularly evident between 700 and 1300 ppm U. On average the umpire results are 3.75% higher than the original results. The %MAPD plot shows that 88% of the data falls within the 10% precision limits for laboratory duplicates.
Figure 5.2 Scatter Plot Umpire Repeats, All Data
20 July 2009 17 Rössing South Assay QC Report July 2010
Figure 5.3 Scatter Plot Umpire Repeats, 0 – 5000 ppm U
Figure 5.3 Scatter Plot Umpire Repeats, 0 – 1500 ppm U
20 July 2009 18 Rössing South Assay QC Report July 2010
Figure 5.4 Q-Q Plot Umpire Repeats, All Data
Figure 5.5 Q-Q Plot Umpire Repeats, 0 – 5000 ppm U
20 July 2009 19 Rössing South Assay QC Report July 2010
Figure 5.6 Q-Q Plot Umpire Repeats, 0 – 1500 ppm U
Figure 5.7 %MAPD Plot Umpire Repeats, 10 – 20000 ppm U
20 July 2009 20 Rössing South Assay QC Report July 2010
6. Data Quality Summary
The review of the data from field and laboratory quality control samples shows that:
• The majority of the certified standards inserted by Extract Resources reported within 3% of the expected values for these standards no consistent positive or negative bias was demonstrated
• 95% of field duplicates inserted by Extract Resource report within the expected 20% tolerance limits
• Blanks inserted by Extract Resources show some contamination from high-grade material, but performance of these blanks improved over the period of the report
• Umpire laboratory repeats are consistent and show a small bias of about 3.75%, which may be attributable to the 4 acid digest being used at Genalysis leaving some uranium behind in insoluble phases e.g. Zircon.
• Laboratory QC data show very high levels of internal accuracy and precision
Peter Wilson Database Manager Extract Resources 21 July 2010
20 July 2009 21 Rössing South Assay QC Report July 2010
Appendix 1.
Control Charts for Extract Resources Standards and Blanks
20 July 2009 22 Rössing South Assay QC Report July 2010
AMIS0046
Standard AMIS0046 Expected Value (EV) 93 Expected Min 84.5 Expected Max 101.5 Count 384 Minimum 79.8 Maximum 98.2 Mean 99.1 Std Deviation 3.083 % in Tolerance 75.5 % Bias 6.6
20 July 2009 23 Rössing South Assay QC Report July 2010
AMIS0054
Standard AMIS0054 Expected Value (EV) 1410 Expected Min 1314 Expected Max 1506 Count 237 Minimum 837.0 Maximum 932.5 Mean 1426.5 Std Deviation 44.639 % in Tolerance 95.4 % Bias 1.2
20 July 2009 24 Rössing South Assay QC Report July 2010
AMIS0076
Standard AMIS0076 Expected Value (EV) 1502 Expected Min 1336 Expected Max 1668 Count 145 Minimum 79.0 Maximum 102.0 Mean 1528.8 Std Deviation 45.736 % in Tolerance 100 % Bias 1.8
20 July 2009 25 Rössing South Assay QC Report July 2010
AMIS0085
Standard AMIS0085 Expected Value (EV) 263 Expected Min 242 Expected Max 284 Count 497 Minimum 271.0 Maximum 272.0 Mean 264.3 Std Deviation 9.863 % in Tolerance 95.4 % Bias 0.5
20 July 2009 26 Rössing South Assay QC Report July 2010
AMIS0086
Standard AMIS0086 Expected Value (EV) 127 Expected Min 117.5 Expected Max 136.5 Count 477 Minimum 115.3 Maximum 139.8 Mean 129.6 Std Deviation 4.997 % in Tolerance 90.4 % Bias 2.1
20 July 2009 27 Rössing South Assay QC Report July 2010
AMIS0090
Standard AMIS0090 Expected Value (EV) 890 Expected Min 831 Expected Max 949 Count 48 Minimum 276.0 Maximum 276.0 Mean 888.9 Std Deviation 29.851 % in Tolerance 91.7 % Bias -0.1
20 July 2009 28 Rössing South Assay QC Report July 2010
AMIS0097
Standard AMIS0097 Expected Value (EV) 527 Expected Min 497 Expected Max 557 Count 506 Minimum 818.0 Maximum 938.0 Mean 527.5 Std Deviation 18.702 % in Tolerance 89.5 % Bias 0.1
20 July 2009 29 Rössing South Assay QC Report July 2010
AMIS0098
Standard AMIS0098 Expected Value (EV) 819 Expected Min 723 Expected Max 915 Count 644 Minimum 1182.8 Maximum 1414.5 Mean 836.4 Std Deviation 26.199 % in Tolerance 99.8 % Bias 2.1
20 July 2009 30 Rössing South Assay QC Report July 2010
AMIS0100
Standard AMIS0100 Expected Value (EV) 1474 Expected Min 1398 Expected Max 1550 Count 408 Minimum 2376.1 Maximum 2758.8 Mean 1471.3 Std Deviation 44.969 % in Tolerance 91.4 % Bias -0.2
20 July 2009 31 Rössing South Assay QC Report July 2010
AMIS0113
Standard AMIS0113 Expected Value (EV) 456 Expected Min 411 Expected Max 501 Count 358 Minimum 2.0 Maximum 7.8 Mean 446.6 Std Deviation 16.712 % in Tolerance 98 % Bias -2.1
20 July 2009 32 Rössing South Assay QC Report July 2010
AMIS0131
Standard AMIS0131 Expected Value (EV) 294 Expected Min 268 Expected Max 320 Count 501 Minimum 13.0 Maximum 26.0 Mean 296.3 Std Deviation 9.548 % in Tolerance 97.2 % Bias 0.8
20 July 2009 33 Rössing South Assay QC Report July 2010
AMIS0133
Standard AMIS0133 Expected Value (EV) 3323 Expected Min 3157 Expected Max 3489 Count 609 Minimum 21.0 Maximum 61.0 Mean 3230.3 Std Deviation 119.561 % in Tolerance 72.1 % Bias -2.8
20 July 2009 34 Rössing South Assay QC Report July 2010
AMIS0154
Standard AMIS0154 Expected Value (EV) 682 Expected Min 621 Expected Max 743 Count 24 Minimum 70.0 Maximum 101.0 Mean 723.4 Std Deviation 14.879 % in Tolerance 83.3 % Bias 6.1
20 July 2009 35 Rössing South Assay QC Report July 2010
AMIS0156
Standard AMIS0156 Expected Value (EV) 195 Expected Min 161 Expected Max 229 Count 32 Minimum 98.8 Maximum 122.2 Mean 210.1 Std Deviation 8.084 % in Tolerance 96.9 % Bias 7.7
20 July 2009 36 Rössing South Assay QC Report July 2010
Field Blank
Field Blanks ICP Expected Value (EV) 2.5 Expected Min 0 Expected Max 10 Count 5206 Outliers 2 Minimum BDL Maximum 71 Mean 2.08 Std Deviation 11.7 % in Tolerance 99
20 July 2009 37 Rössing South Assay QC Report July 2010
Appendix 2.
Control Charts for Genalysis Standards and Blanks
20 July 2009 38 Rössing South Assay QC Report July 2010
AMIS0004
Standard AMIS0004 Expected Value (EV) 88 Expected Min 74.8 Expected Max 101.2 Count 1325 Minimum 63.5 Maximum 103.2 Mean 88.6 Std Deviation 3.585 % in Tolerance 99 % Bias 0.715
20 July 2009 39 Rössing South Assay QC Report July 2010
AMIS0086
Standard AMIS0086 Expected Value (EV) 127 Expected Min 117.5 Expected Max 136.5 Count 367 Minimum 110.9 Maximum 143.7 Mean 128.8 Std Deviation 4.629 % in Tolerance 95 % Bias 1.4
20 July 2009 40 Rössing South Assay QC Report July 2010
CUP-1
Standard CUP-1 Expected Value (EV) 1280 Expected Min 1088 Expected Max 1472 Count 209 Minimum 1182.8 Maximum 1414.5 Mean 1282.9 Std Deviation 38.25 % in Tolerance 100 % Bias 0.23
20 July 2009 41 Rössing South Assay QC Report July 2010
DH-1a
Standard DH-1a Expected Value (EV) 2600 Expected Min 2200 Expected Max 3000 Count 1324 Minimum 2268 Maximum 2875.5 Mean 2571.7 Std Deviation 90.7 % in Tolerance 100 % Bias -1.09
20 July 2009 42 Rössing South Assay QC Report July 2010
OREAS 45P
Standard OREAS 45P Expected Value (EV) 2.4 Expected Min 2.04 Expected Max 2.76 Count 1319 Minimum 0.02 Maximum 29.6 Mean 2.5 Std Deviation 1.1 % in Tolerance 84 % Bias 6.522
20 July 2009 43 Rössing South Assay QC Report July 2010
UTS-1
Standard UTS-1 Expected Value (EV) 49 Expected Min 41.65 Expected Max 56.35 Count 1289 Minimum 33.8 Maximum 104 Mean 49.3 Std Deviation 7.4 % in Tolerance 79 % Bias 0.516
20 July 2009 44 Rössing South Assay QC Report July 2010
GENALYSIS BLANKS
Standard Acid Blank Control Blank Expected Value (EV) 0 0 Expected Min -0.5 -0.5 Expected Max 0.5 0.5 Count 720 2323 Minimum 0 0 Maximum 5 1.1 Mean 0.03 0.04 Std Deviation 0.176 0.069 % in Tolerance 99.9 99.9
20 July 2009 45
Appendix 2
Certificates of Qualified Persons
Coffey Mining Pty. Ltd.
Certificate of Qualified Person
As an author of the report entitled “National Instrument 43-101 Technical Report Husab Uranium Project - May 2011 Project Update” (the “Report”) dated May 20, 2011, prepared on behalf Extract Resources Limited (the “Company”), do hereby certify that:
1. My name is Neil Andrew Inwood and I am a Principal Consultant - Resources with the firm of Coffey Mining Pty. Ltd. of 1162 Hay Street, West Perth, WA, 6005, Australia.
2. I am a practising geologist and a member of the AusIMM (210871).
3. I am a graduate of Curtin University of Technology in Western Australia with a BSc in Geology in 1993 and a PGradDip in Hydro-Geology in 1994. In 2007 I graduated from the University of Western Australia with an MSc in Geology and from Edith Cowan University with a Post Graduate Certificate in Geostatistics.
4. I have practiced my profession continuously since 1994.
5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
6. I visited the Husab property and surrounding areas on several occasions in 2008, 2009 and 2010. I have performed consulting services and reviewed files and data supplied by Extract resources from 2008 to May 2011.
7. I contributed to and am responsible for all sections of this report apart from Sections 5.3, 16, 18, 19, 20, 23, and 24 and the associated text in the summary, conclusions and recommendations.
8. As of the date of this certificate, to the best of my knowledge, information and belief, the study contains all scientific and technical information that is required to be disclosed to make the Study not misleading.
9. I am independent of Extract Resources pursuant to section 1.4 of the Instrument.
10. I have read the National Instrument and Form 43-101F1 (the “Form”) and the Report has been prepared in compliance with the Instrument and the Form.
11. I do not have nor do I expect to receive a direct or indirect interest in the Husab Project property of Extrac Resources, and I do not beneficially own, directly or indirectly, any securities of Extract Resources or any associate or affiliate of such company.
12. As of the date of this certificate, to the best of my knowledge, information and belief, the Report contains all scientific and technical information that is required to be disclosed to make the Report not misleading.
Dated at Perth, Western Australia, on 20 May 2011.
[Signed] Neil Inwood BSc (Geology) Principal Consultant - Resources MSc (Geology) Post Grad Cert Geostatistics
Appendix 2 – Certificates of Qualified Persons page: 1
Coffey Mining Pty. Ltd.
Certificate of Qualified Person
As an author of the report entitled “National Instrument 43-101 Technical Report Husab Uranium Project - May 2011 Project Update” (the “Report”) dated May 20, 2011, prepared on behalf Extract Resources Limited (the “Company”), do hereby certify that:
1. My name is Steven Le Brun and I am a Principal Consultant - Resources with the firm of Coffey Mining Pty. Ltd. of 1162 Hay Street, West Perth, WA, 6005, Australia.
2. I am a practising geologist and a member of the AusIMM (202832) and of MICA
3. 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.
4. I have practiced my profession continuously since 1987.
5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).
6. I have not visited the Husab property.
7. I contributed to and am responsible for portions of Section 17 and the associated text in the summary, conclusions and recommendations.
8. As of the date of this certificate, to the best of my knowledge, information and belief, the study contains all scientific and technical information that is required to be disclosed to make the Study not misleading.
9. I am independent of Extract Resources pursuant to section 1.4 of the Instrument.
10. I have read the National Instrument and Form 43-101F1 (the “Form”) and the Report has been prepared in compliance with the Instrument and the Form.
11. I do not have nor do I expect to receive a direct or indirect interest in the Husab Project property of Extrac Resources, and I do not beneficially own, directly or indirectly, any securities of Extract Resources or any associate or affiliate of such company.
12. As of the date of this certificate, to the best of my knowledge, information and belief, the Report contains all scientific and technical information that is required to be disclosed to make the Report not misleading.
Dated at Perth, Western Australia, on 20 May 2011.
[Signed] Steve Le Brun BSc (Hons) (Geology Sciences) Principal Consultant - Resources MSc (Mineral Exploration and Mining Geology)
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[signed]
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[signed]
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[signed]
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[signed]
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CERTIFICATE OF QUALIFIED PERSON
I, Stephen Amos, as a reviewer of the report entitled “National Instrument 43-101 Technical Report Husab Uranium Project – May 2011 Project Update” (the “ Report ”) dated May, 2011, prepared on behalf Extract Resources Limited (the “ Company ”), do hereby certify that:
1. My My name is Stephen Amos and I am employed by AMEC Minproc in the capacity of Technical Manager. The office AMEC Minproc is located at Highbury House, Hampton Office Park, 20 Georgian Crescent, Bryanston, South Africa;
2. I am a graduate of the University of the Witwatersrand with a BSc Honours degree in Applied Chemistry and an MSc in Metallurgy. I am a Metallurgist, with 21 years experience in process engineering, design, management, commissioning and R&D. I am a Fellow of South African Institute of Mining and Metallurgy (SAIMM), a professional society as defined by NI 43-101. I have practiced my profession as a metallurgist continually since 1990, about 21 years. I have worked for Anglo Platinum (1990 - 2005) and AMEC Minproc (previously GRD Minproc) (2006 to present);
3. I have read the definition of "qualified person" set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“ NI 43-101 ”) and certify that by reason of my education, affiliation with a professional association as defined in NI 43-101, and past relevant work experience, I fulfill the requirements of a "qualified person" for the purposes of NI 43-101;
4. I have not inspected the Husab property;
5. I have reviewed section 23 relating to the processing costs of the Report;
6. I am independent from the Company as described in section 1.4 of NI 43-101;
7. AMEC Minproc has prepared the Definitive Feasibility Study for the Husab project. I have been involved in the management of the process engineering design of the plant;
8. I have read the NI 43-101 and the Report has been prepared in compliance with NI 43-101; and
9. As of the date of this certificate, to the best of my knowledge, information and belief, the Report contains all scientific and technical information that is required to be disclosed to make the Report not misleading.
Dated this 20th day of May, 2011
[signed] Stephen Amos , FSAIMM Technical Manager AMEC Minproc
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