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NI 43-101 TECHNICAL REPORT MERLIN PROJECT PRE

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NI 43-101 TECHNICAL REPORT MERLIN PROJECT PRE Powered By Docstoc
					AMC Consultants Pty Ltd
ABN 58 008 129 164



Level 19, 114 William Street
MELBOURNE VIC 3000

T   +61 3 8601 3300
F   +61 3 8601 3399
E   am cmelb@amcconsultants.com .au




                                 NI 43-101 TECHNICAL REPORT
                     MERLIN PROJECT PRE-FEASIBILITY STUDY
                                NW QUEENSLAND, AUSTRALIA


 Prepared by AMC Consultants Pty Ltd in accordance with the requirements of
 National Instrument 43-101, “Standards of Disclosure for Mineral Project”, of
                   the Canadian Securities Administrators



Qualified Persons:             Peter McCarthy, Director, Principal Engineer, AMC
                               Consultants Pty Ltd, FAusIMM (CP).

                               John Horton, Principal Geologist, Golder Associates Pty Ltd,
                               FAusIMM (CP), MAIG

                               Tom Hunter, Associate Director, Jacobs E&C Australia Pty
                               Ltd, FAusIMM

Submitted to:                  IVANHOE AUSTRALIA LIMITED
                               Level 13, 484 St Kilda Road
                               Melbourne, VIC 3004
                               Australia




                                             AMC 111041
                                  Effective Date: 10 October 2011



                                                                                UNIT ED
  ADELAIDE            BRISBANE          MELBOURNE            PERTH                             VANCOUVER
                                                                              KINGDOM
+61 8 8201 1800     +61 7 3839 0099    +61 3 8601 3300   +61 8 6330 1100                      +1 604 669 0044
                                                                           +44 1628 778 256

                                          www.amcconsultants.ca
IVANHOE AUSTRALIA LIMITED
NI 43-101 Technical Report Merlin Pr oject Prefeasibility Study




                                                   CONTENTS

1     SUMMARY ...............................................................................................................1
2     INTRODUCTION......................................................................................................6
      2.1 Terms of Reference and Purpose of this Report ...........................................6
      2.2 Qualifications of Consultants ..........................................................................7
      2.3 Site Visits ........................................................................................................7
3     RELIANCE ON OTHER EXPERTS .........................................................................8
4     PROPERTY DESCRIPTION AND LOCATION .......................................................9
      4.1 Property Location............................................................................................9
      4.2 Land Tenure....................................................................................................9
5     ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE
      AND PHYSIOGRAPHY ..........................................................................................14
      5.1 Accessibility ..................................................................................................14
      5.2 Vegetation and Climate ................................................................................14
      5.3 Water Supply ................................................................................................14
      5.4 Infrastructure .................................................................................................15
      5.5 Mining Surface Infrastruture .........................................................................15
6     HISTORY................................................................................................................16
7     GEOLOGICAL SETTING AND MINERALISATION ..............................................20
      7.1 Regional Geology .........................................................................................21
      7.2 Local and Property Geology .........................................................................26
      7.3 Mineralisation................................................................................................29
            7.3.1    Copper Mineralisation.....................................................................30
            7.3.2    Molybdenum-Rhenium Mineralisation ............................................31
8     DEPOSIT TYPES...................................................................................................34
9     EXPLORATION ......................................................................................................36
10    DRILLING ...............................................................................................................40
      10.1 Drilling Methods ............................................................................................40
      10.2 Collar Surveys...............................................................................................42
      10.3 Down Hole Surveys ......................................................................................42
      10.4 Recoveries and Rock Quality .......................................................................43
11    SAMPLE PREPARATION, ANALYSIS AND SECURITY .....................................44
      11.1 Methods44
      11.2 Procedures....................................................................................................44
            11.2.1 Sample Dispatches .........................................................................44
            11.2.2 Sampling of Diamond Drill Core .....................................................45
            11.2.3 Sampling of Reverse Circulation Cuttings .....................................46
      11.3 Bulk Densities ...............................................................................................46
      11.4 Magnetic Susceptibility .................................................................................48
      11.5 Logging 48
      11.6 Database Management ................................................................................49
      11.7 Adequacy of Sampling..................................................................................50
      11.8 Quality Assurance and Quality Control Procedures ....................................50
      11.9 Laboratories ..................................................................................................52
      11.10 Sample Preparation ......................................................................................52
      11.11 Analyses .......................................................................................................53
      11.12 Monitoring .....................................................................................................55
            11.12.1 Standard Reference Materials........................................................56
            11.12.2 Field Blanks ....................................................................................59
      11.13 Duplicates .....................................................................................................60
      11.14 Checking Programs ......................................................................................76



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            11.14.1 Programs ........................................................................................76
            11.14.2 Set MD1 - Digestion Checks ..........................................................77
            11.14.3 Set MD2 - Metallurgical Checks .....................................................77
            11.14.4 Set MD3 - Discovery Checks .........................................................77
            11.14.5 Set MD4 - Method Validation Checks ............................................78
            11.14.6 Set MD5 - Routine Checks .............................................................80
            11.14.7 Set MD6 - Routine Checks .............................................................81
            11.14.8 Set MD7 - Routine Checks .............................................................81
            11.14.9 Set MD8 - Routine Checks .............................................................82
            11.14.10 Set MD9 - High Grade Ratio Checks .........................................83
            11.14.11 Twinning ......................................................................................84
            11.14.12 Security and Chain of Custody ...................................................87
      11.15 Adequacy of Sample Preparation, Analytical, and Security
            Procedures....................................................................................................87
12    DATA VERIFICATION ...........................................................................................89
13    MINERAL PROCESSING AND METALLURGICAL TESTING .............................93
      13.1 Metallurgical Testwork ..................................................................................93
      13.2 Summary of Testwork...................................................................................93
14    MINERAL RESOURCE ESTIMATES ....................................................................95
      14.1 Resource Domains .......................................................................................95
      14.2 Grade Domains.............................................................................................95
      14.3 Mo Domaining (MODOM) .............................................................................97
      14.4 Cu Domaining (CUDOM) ............................................................................100
      14.5 Polymetallic Domaining (P_DOM)..............................................................101
      14.6 Geology Domains (ROCK) .........................................................................102
      14.7 Weathering Domains (MINL) ......................................................................103
      14.8 Metallurgical Domains (METDOM).............................................................103
      14.9 Domain Definitions......................................................................................105
      14.10 Combined Domains (DOM) ........................................................................105
      14.11 Domain Boundary .......................................................................................108
      14.12 Data Preparation.........................................................................................108
            14.12.1 Database Preparation...................................................................108
            14.12.2 Recovery .......................................................................................111
            14.12.3 Dry Bulk Density ...........................................................................112
            14.12.4 Default Grades ..............................................................................113
            14.12.5 Domain Flagging...........................................................................113
            14.12.6 Compositing ..................................................................................114
            14.12.7 Declustering ..................................................................................116
            14.12.8 Top-Cutting ...................................................................................117
      14.13 Data Analysis ..............................................................................................119
            14.13.1 Density (SG) Data.........................................................................119
            14.13.2 Statistics........................................................................................119
            14.13.3 Variography...................................................................................120
            14.13.4 QKNA ............................................................................................124
            14.13.5 Geological Block Model ................................................................125
            14.13.6 Grade Estimation Parameters ......................................................130
            14.13.7 Estimation Results ........................................................................132
      14.14 Model Validation .........................................................................................136
      14.15 Mineral Resource Classification .................................................................139
      14.16 Mineral Resource Statement ......................................................................140
15    MINERAL RESERVE ESTIMATES .....................................................................145
16    MINING METHODS .............................................................................................147



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      16.1 Geotechnical ...............................................................................................147
      16.2 Stoping Methods .........................................................................................149
      16.3 Mine Design ................................................................................................151
      16.4 Mining Schedule .........................................................................................152
17    RECOVERY METHODS ......................................................................................155
      17.1 Process Plant Design Criteria ....................................................................155
           17.1.1 General .........................................................................................155
           17.1.2 Ore Blending .................................................................................155
           17.1.3 Crushing........................................................................................155
           17.1.4 Grinding.........................................................................................156
           17.1.5 Flotation ........................................................................................156
           17.1.6 Services & Tailings .......................................................................156
           17.1.7 Concentrate Handling ...................................................................156
           17.1.8 Roaster & Off Gas Handling .........................................................156
           17.1.9 Rhenium Recovery .......................................................................157
           17.1.10 Calcine Leach & Molybdenum Recovery .....................................157
      17.2 Process Plant Design .................................................................................157
18    PROJECT INFRASTRUCTURE ..........................................................................164
      18.1 Paste Plant..................................................................................................165
      18.2 Infrastructure ...............................................................................................165
           18.2.1 Electric Power...............................................................................165
           18.2.2 Other Infrastructure.......................................................................166
      18.3 Tailings and Effluent Management.............................................................166
      18.4 Personnel ....................................................................................................168
19    MARKET STUDIES AND CONTRACTS .............................................................170
      19.1 Marketing Studies .......................................................................................170
      19.2 Molybdenum ...............................................................................................170
      19.3 Rhenium......................................................................................................173
      19.4 Contracts.....................................................................................................175
           19.4.1 Mining............................................................................................175
           19.4.2 Haul Road Access to Osborne .....................................................175
           19.4.3 Power Station................................................................................176
           19.4.4 Process Plants ..............................................................................176
           19.4.5 Feasibility Study............................................................................177
           19.4.6 On-going Engineering Work .........................................................177
20    ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR
      COMMUNITY IMPACT ........................................................................................178
      20.1 The Project Operation: ...............................................................................178
      20.2 Waste Rock Characterisation .....................................................................178
      20.3 Groundwater ...............................................................................................179
      20.4 Surface Water .............................................................................................179
      20.5 Conclusion ..................................................................................................179
      20.6 Permits 179
      20.7 Social/Community Relations Requirement ................................................180
21    CAPITAL AND OPERATING COSTS..................................................................181
      21.1 Capital Costs...............................................................................................181
           21.1.1 Summary Costs ............................................................................181
           21.1.2 Direct Costs ..................................................................................181
           21.1.3 Indirect Costs ................................................................................182
           21.1.4 Contingency ..................................................................................183
           21.1.5 Exclusions .....................................................................................183
           21.1.6 Total Project Capital Cost .............................................................184



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            21.1.7 Cost to First Production ................................................................184
      21.2 Operating Costs ..........................................................................................184
            21.2.1 Introduction ...................................................................................184
            21.2.2 Employee Costs............................................................................185
            21.2.3 Combined Operating Costs ..........................................................186
            21.2.4 Unit Operating Cost ......................................................................186
22    ECONOMIC ANALYSIS.......................................................................................188
      22.1 Introduction .................................................................................................188
      22.2 Metal Sale Prices ........................................................................................188
      22.3 Exchange Rates .........................................................................................188
      22.4 Taxes 189
      22.5 Goods and Services Tax ............................................................................190
      22.6 Carbon Trading Scheme ............................................................................190
      22.7 Royalties .....................................................................................................190
            22.7.1 Queensland State Royalty ............................................................190
            22.7.2 Previous Owner Royalty ...............................................................190
      22.8 Native Title Compensation .........................................................................191
      22.9 Other Royalties/Agreements ......................................................................191
      22.10 Revenue Deductions ..................................................................................191
      22.11 Reclamation ................................................................................................191
      22.12 Project Financing ........................................................................................191
      22.13 Assumptions ...............................................................................................191
      22.14 Cash Operating Costs ................................................................................192
      22.15 Revenue......................................................................................................193
      22.16 Cash Flow ...................................................................................................194
      22.17 NPV, IRR and Payback Period...................................................................194
      22.18 Net Present Value (NPV) Sensitivity ..........................................................194
      22.19 After-tax Cash Flow (ATCF) Sensitivity......................................................195
23    ADJACENT PROPERTIES ..................................................................................197
24    OTHER RELEVANT DATA AND INFORMATION ..............................................198
25    INTERPRETATION AND CONCLUSIONS .........................................................200
      25.1 Geology 200
            25.1.1 Merlin - Little Wizard Mo-Re Deposit ...........................................200
            25.1.2 Resource Estimate .......................................................................201
      25.2 Geotechnical, Mining and Mineral Reserve ...............................................201
      25.3 Processing ..................................................................................................202
      25.4 Infrastructure ...............................................................................................202
      25.5 Risks 202
26    RECOMMENDATIONS ........................................................................................204
      26.1 Geological ...................................................................................................204
      26.2 Geotechnical ...............................................................................................205
      26.3 Mining 206
      26.4 Processing ..................................................................................................206
            26.4.1 Concentrate Production ................................................................206
            26.4.2 Concentrate Treatment.................................................................207
            26.4.3 Crushing Facility ...........................................................................207
            26.4.4 Location of the Concentrate Treatment Facility ...........................207
            26.4.5 Sulphur Dioxide Roaster Gas Cleaning .......................................208
            26.4.6 Sell or Toll Treat Molybdenum Concentrates...............................208
27    REFERENCES.....................................................................................................210
28    DATE AND SIGNATURE PAGE..........................................................................213




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                                                     TABLES

Table 1.1        Merlin Mineral Reserve Estimate...............................................................2
Table 1.2        Merlin Project PFS Operating and Cost Data ...........................................2
Table 2.1        Persons Who Prepared this Technical Report ..........................................7
Table 2.2        Persons Who Contributed to this Technical Report ..................................7
Table 4.1        Tenure ......................................................................................................10
Table 6.1        IVA Previous Estimates for Merlin Mo-Re Resource at 0.3% Mo
                 Cut-off.......................................................................................................19
Table 9.1        Little Wizard Resource Intercepts............................................................39
Table 10.1       Summary of Drilling Companies, Date and Method ................................41
Table 10.2       Collar Survey Method ..............................................................................42
Table 10.3       Down Hole Survey Method Summary .....................................................43
Table 11.1       Comparison of Bulk Density Pairs from the Same Sample Interval .......48
Table 11.2       Mount Dore - Merlin - Standard Reference Materials, 28 June,
                 2010 .........................................................................................................51
Table 11.3       Type and Number of Duplicate Samples by Analytical Methods ............61
Table 11.4       Levels of Asymptotic Precision for OG46 Duplicates..............................61
Table 11.5       Levels of Precision at 90th Percentile for Duplicates ..............................75
Table 11.6       Visual Intercept Comparisons for RC vs. Diamond Drilling ....................86
Table 14.1       Cu Equivalence Conversion Factors @ ...................................................102
Table 14.2       Mean Sequential Copper Analyses by the Metallurgical Domains
                 (METDOM), all Samples ........................................................................105
Table 14.3       Mean Sequential Copper Analyses by the Metallurgical Domains
                 (METDOM), >0.25% Cu.........................................................................105
Table 14.4       Individual Domain Codes .......................................................................107
Table 14.5       Combined Domain Codes......................................................................108
Table 14.6       Drilling Data by Drilling Type .................................................................109
Table 14.7       Assayed Drilling by Year, Company and Type......................................110
Table 14.8       Assay Completeness by Drilling Company And Type...........................110
Table 14.9       Dry Bulk Density Averages by Domain and Weathering Type .............113
Table 14.10      Merlin Density Samples .........................................................................113
Table 14.11      Proportion of Sampled, Unsampled and Lost Sample Interval by
                 Company and Drill Type ........................................................................114
Table 14.12      Average Length-Weighted Grades Before and After Compositing .......116
Table 14.13      Top Cuts by Grouped Domains .............................................................118
Table 14.14      Variogram Models ..................................................................................124
Table 14.15      Model Definition .....................................................................................126
Table 14.16      Model Field Values ................................................................................126
Table 14.17      Additional Categorical Value Definition .................................................127
Table 14.18      Wireframe - Block Model Volume Comparison .....................................128
Table 14.19      Default Values for Unestimated Blocks .................................................133
Table 14.20      Proportion of Blocks Estimated .............................................................133
Table 14.21      Global Mean and Variance Comparison ...............................................136
Table 14.22      Mount Dore - Merlin Mineral Resource Breakdown ..............................141
Table 14.23      Merlin Mineral Resource Weathering Type ...........................................144
Table 15.1       Mineral Reserve Estimate showing the Split by Mining Method ...........146
Table 16.1       Recommended Drift and Fill Dimensions ..............................................148
Table 16.2       Recommended Stope Dimensions ........................................................148
Table 16.3       Average Numbers of Heavy Mobile Equipment ....................................154
Table 17.1       Key Process Plant Parameters..............................................................158
Table 17.2       ROM Feed Grades.................................................................................160



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Table 17.3       Metal Recoveries ...................................................................................160
Table 17.4       Plant Production Estimates....................................................................160
Table 18.1       TSF Options Comparison based on A$ per m 3 Tailings........................167
Table 18.2       Operating Workforce and Numbers of Personnel .................................169
Table 19.1       Estimated 2010 World Molybdenum Demand by Application, 2010.....171
Table 19.2       Forecast Molybdenum Oxide Price .......................................................172
Table 19.3       Forecast Average Rhenium Metal Prices 2010 to 2016 .......................174
Table 21.1       Summary of Capital Cost Estimate........................................................181
Table 21.2       Summary of Direct Costs .......................................................................182
Table 21.3       Summary of Indirect Costs ....................................................................183
Table 21.4       Contingency Allowance .........................................................................183
Table 21.5       Employee Numbers ...............................................................................185
Table 21.6       Combined Operating Costs ...................................................................186
Table 21.7       Operating Costs per lb of Molybdenum .................................................187
Table 21.8       Average Cash Cost for Molybdenum ....................................................187
Table 22.1       Key Assumptions ...................................................................................192
Table 22.2       Average Operating Costs ......................................................................192
Table 22.3       NPV Sensitivity to Discount Rate ..........................................................194
Table 22.4       NPV at Various Metal Prices .................................................................195
Table 22.5       ATCF at Various Metal Prices ...............................................................196
Table 25.1       Key Process Plant Parameters..............................................................202


                                                  FIGURES

Figure 4.1       Project Location ................................................................................................ 9
Figure 4.2       Mount Dore Mining Leases ............................................................................. 11
Figure 4.3       Mount Dore Mining Leases and Mineralisation .............................................. 12
Figure 6.1       Historic Workings at Mount Dore Displays Limited Disturbance.................... 16
Figure 7.1       Cloncurry-Selwyn Zone of the Eastern Fold Belt ........................................... 20
Figure 7.2       District Geology for the Mount Elliott to Mount Dore - Merlin Area ................ 24
Figure 7.3       Aerial View Showing Proximity of the Starra .................................................. 25
Figure 7.4       Aerial View showing Local Topography of the Mount Dore Area .................. 25
Figure 7.5       Typical Host Rocks for Mount Dore North ...................................................... 27
Figure 7.6       Geology, Mo & Cu Significant Intersections on Section 7605450 N.............. 28
Figure 7.7       Geology and Drill Hole Collars in the Mount Dore - Merlin Area ................... 29
Figure 7.8       Schematic Cross Section of the Mount Dore Copper Deposit ....................... 31
Figure 7.9       Molybdenite in Matrix Supported Breccia....................................................... 32
Figure 7.10      Irregular and Discontinuous Fracture-Fill Molybdenite .................................. 32
Figure 7.11      Stylolitic Low Grade Molybdenite Mineralisation ............................................ 33
Figure 9.1       Plan of Resource Drilling ................................................................................ 37
Figure 9.2       February 2010 Merlin Drilling and Site ........................................................... 38
Figure 9.3       Massive Molybdenite in MDQ0264 (Little Wizard Zone)................................ 39
Figure 11.1      Scatter Plot of Pairs of Dry Bulk Density Measurements for the Same
                 Sample Intervals ............................................................................................. 49
Figure 11.2      Cross Section Showing High Magnetic Susceptibility Zone on the
                 Footwall of the Merlin ...................................................................................... 49
Figure 11.3      Quality Control Monitoring Charts for Cu Assays, MDL-1 ............................. 56
Figure 11.4      Quality Control Monitoring Charts for Cu Assays, MDM-1 ............................ 56
Figure 11.5      Quality Control Monitoring Charts for Cu Assays, MDH-1 ............................. 56
Figure 11.6      Quality Control Monitoring Charts for Mo Assays, MDM-2 ............................ 57




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Figure 11.7      Quality Control Monitoring Charts for Mo Assays, MDH-2 ............................ 57
Figure 11.8      Quality Control Monitoring Charts for Cu Assays, MEL-1.............................. 57
Figure 11.9      Quality Control Monitoring Charts for Cu Assays, MEM-2............................. 57
Figure 11.10     Quality Control Monitoring Charts for Cu Assays, MXL-1.............................. 57
Figure 11.11     Quality Control Monitoring Charts for Mo Assays, MLB-1 ............................. 58
Figure 11.12     Quality Control Monitoring Charts for Rhenium Assays, MLB-1.................... 58
Figure 11.13     Quality Control Monitoring Charts for Mo Assays, MLL-1.............................. 58
Figure 11.14     Quality Control Monitoring Charts for Re Assays, MLL-1 .............................. 58
Figure 11.15     Quality Control Monitoring Charts for Mo Assays, MLM-1............................. 58
Figure 11.16     Quality Control Monitoring Charts for Re Assays, MLM-1 ............................. 59
Figure 11.17     Quality Control Monitoring Charts for Re Assays in NCSDC70018 .............. 59
Figure 11.18     Quality Control Monitoring Charts for Re Assays in GMO-04........................ 59
Figure 11.19     Quality Control Monitoring Charts for Cu Assays in Field Blanks.................. 60
Figure 11.20     Quality Control Monitoring Charts for Mo Assays in Field Blanks ................. 60
Figure 11.21     Quality Control Monitoring Chart for Re Analyses in Field Blanks ................ 60
Figure 11.22     Cu by OG62 for Core Duplicates .................................................................... 63
Figure 11.23     Mo by OG62 for Core Duplicates ................................................................... 64
Figure 11.24     Re by OG62 for Core Duplicates .................................................................... 65
Figure 11.25     Cu by OG46 for Core Duplicates .................................................................... 66
Figure 11.26     Cu by OG62 for Crush Duplicates .................................................................. 67
Figure 11.27     Mo by OG62 for Crush Duplicates.................................................................. 68
Figure 11.28     Re by OG62 for Crush Duplicates .................................................................. 69
Figure 11.29     Cu by OG46 for Crush Duplicates .................................................................. 70
Figure 11.30     Cu by OG62 for Pulp Duplicates .................................................................... 71
Figure 11.31     Mo by OG62 for Pulp Duplicates .................................................................... 72
Figure 11.32     Re by OG62 for Pulp Duplicates .................................................................... 73
Figure 11.33     Cu by OG46 for Pulp Duplicates .................................................................... 74
Figure 11.34     TH plot for Cu by OG62 .................................................................................. 75
Figure 11.35     TH plot for Mo by OG62.................................................................................. 75
Figure 11.36     TH plot for Re by OG62 .................................................................................. 75
Figure 11.37     TH plot for Cu by OG46 .................................................................................. 75
Figure 11.38     PR plot for Cu by OG62 .................................................................................. 76
Figure 11.39     PR plot for Mo by OG62 ................................................................................. 76
Figure 11.40     PR plot for Re by OG62 .................................................................................. 76
Figure 11.41     PR plot for Cu by OG46 .................................................................................. 76
Figure 11.42     Comparison of ALS Cu by OG46 Versus OG62 Methods, MD1 ................... 77
Figure 11.43     Comparison of ALS Mo by OG46 Versus OG62 Methods, MD1 ................... 77
Figure 11.44     Metallurgical Check Assays for Mo, MD2....................................................... 78
Figure 11.45     Actlabs Re Discovery Check Analyses, MD3................................................. 78
Figure 11.46     Re Neutron Activation Method Validation Analyses, MD4 ............................. 79
Figure 11.47     Re Four Acid Digestion Method Validation Analyses, MD4 ........................... 79
Figure 11.48     Mo Four Acid Digestion Method Validation Analyses, MD4 .......................... 79
Figure 11.49     Cu Check Assays, MD5 .................................................................................. 80
Figure 11.50     Mo Check Assays, MD5 ................................................................................. 80
Figure 11.51     Cu Check Assays, MD6 .................................................................................. 82
Figure 11.52     Mo Check Assays, MD6 ................................................................................. 82
Figure 11.53     Re Check Assays, MD6 .................................................................................. 82
Figure 11.54     Cu Check Assays, MD7 .................................................................................. 82
Figure 11.55     Mo Check Assays, MD7 ................................................................................. 82
Figure 11.56     Re Check Assays, MD7 .................................................................................. 82
Figure 11.57     Cu Check Assays, MD8 .................................................................................. 84
Figure 11.58     Mo Check Assays, MD8 ................................................................................. 84



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Figure 11.59     Re Check Assays, MD8 .................................................................................. 84
Figure 11.60     Mo High Grade Ratio Check Assays, MD8 .................................................... 84
Figure 11.61     Re High Grade Ratio Check Assays, MD8..................................................... 84
Figure 11.62     Cu High Grade Ratio Check Assays, MD8..................................................... 84
Figure 11.63     RC Cu Versus Diamond Drilling Cu................................................................ 85
Figure 11.64     RC Mo Versus Diamond Drilling Mo............................................................... 85
Figure 11.65     RC Sulphur Versus Diamond Drilling Sulphur................................................ 85
Figure 14.1      Plot of Mo vs. Re Assay Values from Drill Cores (Kirby 2009) ...................... 98
Figure 14.2      Typical Laser Ablation - ICPMS Trace of “Dirty and Clean”
                 Molybdenite (Kirby 2009) ................................................................................ 98
Figure 14.3      Mo & Re Grade Distribution from all Samples ............................................... 99
Figure 14.4      Mo & Re Log-Probability Plots by Mo Domains (MODOM) ......................... 100
Figure 14.5      Cu Grade Distribution from Samples............................................................ 100
Figure 14.6      Zn Grade Distribution from Samples ............................................................ 102
Figure 14.7      Sequential Cu Ternary Diagram Colour Coded on Existing Initial
                 Reinterpretation Based on Sequential Copper............................................. 104
Figure 14.8      Merlin Mo vs. Density.................................................................................... 113
Figure 14.9      Distribution of all Sample and Composite Lengths ...................................... 115
Figure 14.10     Distribution of Merlin Domain Sample and Composite Lengths .................. 115
Figure 14.11     Mo & Cu Variogram Models for the Merlin High Grade Mo Domain............ 122
Figure 14.12     Cu & Zn Variogram Models for the Lower Cu & Zn Domains ...................... 123
Figure 14.13     Plan View of Wireframes .............................................................................. 129
Figure 14.14     Section 7,605,450 mN: ROCK (Lithology).................................................... 129
Figure 14.15     Section 7,605,450 mN: METDOM ................................................................ 129
Figure 14.16     Section 7,605,450 mN: MINL (Weathering) ................................................. 129
Figure 14.17     Section 7,605,450 mN: MODOM (Mo Domain)............................................ 129
Figure 14.18     Section 7,605,450 mN: CUDOM (Cu Domain)............................................. 130
Figure 14.19     Section 7,605,450 mN: P_DOM (Polymetallic) ............................................ 130
Figure 14.20     Section 7,605,450 mN: DOM (Combined Domain) ...................................... 130
Figure 14.21     Section 7,605,450 mN: RESCAT (Classification) ........................................ 130
Figure 14.22     Section 7,605,450 mN: Cu............................................................................ 134
Figure 14.23     Section 7,605,450 mN: Mo ........................................................................... 134
Figure 14.24     Section 7,605,450 mN: Zn ............................................................................ 134
Figure 14.25     Section 7,605,450 mN: Re............................................................................ 134
Figure 14.26     Section 7,605,450 mN: Ag ............................................................................ 135
Figure 14.27     Section 7,605,450 mN: Au ............................................................................ 135
Figure 14.28     Section 7,605,450 mN: Pb ............................................................................ 135
Figure 14.29     Section 7,605,450 mN: Fe ............................................................................ 135
Figure 14.30     Section 7,605,450 mN: Co............................................................................ 135
Figure 14.31     Section 7,605,450 mN: S .............................................................................. 135
Figure 14.32     Discrete Gaussian Comparison for Cu......................................................... 138
Figure 14.33     Discrete Gaussian Comparison for Zn ......................................................... 138
Figure 14.34     Discrete Gaussian Comparison for Mo ........................................................ 139
Figure 16.1      High Grade Molybdenum Wireframe (Longitudinal Section - Aug ’10
                 Model) ........................................................................................................... 149
Figure 16.2      High Underhand DAF Mining in a Narrow Orebody ..................................... 151
Figure 16.3      Case 2 Mine Design Layout.......................................................................... 152
Figure 16.4      Annual Production Schedule ........................................................................ 153
Figure 17.1      Crushing Circuit Schematic .......................................................................... 161
Figure 17.2      Concentrate Production Plant Schematic..................................................... 162
Figure 17.3      Concentrate Treatment Plant Schematic ..................................................... 163
Figure 22.1      A$ / US$ Exchange Rate .............................................................................. 189



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Figure 22.2      Value of Sales Revenues ............................................................................. 193
Figure 22.3      Production vs. Unit Operating Cost .............................................................. 193
Figure 22.4      After Tax Cash Flow ..................................................................................... 194
Figure 22.5      NPV Sensitivity at +/-10%............................................................................. 195
Figure 22.6      ATCF Sensitivity at +/-10%........................................................................... 196




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1     SUMMARY

Overview

Ivanhoe Australia Limited (IVA) has a number of mineral prospects in the Cloncurry
area of Queensland. The prospect area is covered by eleven Mining Leases held by
IVA within the Mount Isa Mining District. Access is by air from Mount Isa or
Townsville and by road from Townsville.

Priority is being given to development of the Merlin Molybdenum-Rhenium Project
(“The Project”), which is located some 145 km south-east of Mount Isa and 53 km
north of the Osborne processing complex.

In October 2010, IVA purchased mining leases and the Osborne mine, concentrator
and infrastructure from Barrick (Osborne) Pty Ltd. The purchase included a 2 million
tonne/year copper gold concentrator and associated maintenance and office
facilities, a 24 MW gas/diesel power station, accommodation village for 470 people
and a sealed airstrip.

Mineral Reserves

The Merlin Mo/Re orebody has unusually high molybdenum and rhenium grades
compared to most other molybdenum-containing orebodies. Concentrate that will be
produced by the Merlin project is not a by-product of copper sulphide flotation but
instead is a primary source of molybdenum and rhenium.

An independent Mineral Resource estimate by Golder Associates (Golder) of
Brisbane, in August 2010 for Merlin and Little Wizard using a 0.3% molybdenum cut-
off reported:
•     Indicated Mineral Resource of 6.5 million tonnes @ 1.3% molybdenum and
      23 g/t of rhenium; and
•     Inferred Mineral Resource of 0.2 million tonnes @ 0.9% molybdenum and
      15 g/t of rhenium.

This was reported in a NI43-101 technical report prepared by Golder titled “NI43-101
Technical Report: Mount Dore - Merlin Deposit, NWQld, Australia; Report number
107631016-011- Rev0; Submitted 19 Oct 2010”.

A Pre-Feasibility study (PFS) on the mining and processing of the Merlin deposit for
the recovery of molybdenum, rhenium, copper and silver has been completed. This
PFS has resulted in a Mineral Reserve estimate by AMC Consultants Pty Ltd (AMC)
for the Merlin project as set out in Table 1.1.




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Table 1.1           Merlin Mineral Reserve Estimate
     Category           Reserve (Mt)           Mo (t)        Re (kg)             Mo (%)      Re (ppm)
    Proved                     -                  -             -                  -            -
    Probable                 6.7               75,000       128,000                1.1         19.1
    Total                    6.7               75,000       128,000                1.1         19.1

   The effective date for this mineral reserve is 10 Octob er 2011.

The initial project life based on the known Mineral Reserves is greater than 10 years.

Operating and Cost Data

As a consequence of the purchase of the Osborne infrastructure, IVA’s PFS scope
was adjusted to consider regional aspects of the development of the Merlin deposit.
The PFS has identified that project economics would be optimised by the
development of a new access road between Merlin and Osborne and the
construction a new purpose-built molybdenum/rhenium sulphide flotation
concentrator located adjacent to the existing copper/gold concentrator.

Ore will be crushed at Merlin and transported the 53 km to Osborne for processing.
This pathway will allow synergies between the proposed Merlin molybdenum based
operations and processing and the proposed copper-gold production and processing
activities centralised at Osborne.

The key Merlin project PFS operating and cost data are summarised in Table 1.2.

Table 1.2           Merlin Project PFS Operating and Cost Data
             PFS Data                                               Value
             Ore mined                                              6.7 Mt
             Run Of Mine feed                                       500,000 dt/a
             Average Run of Mine grade                              1.1 % Mo,19 g/t Re
             Concentrate produced                                   225,100 tonnes
             Concentrate grade *                                    30 % Mo, 0.046 % Re
             Molybdenum produced (contained in MoO3)                5,030 t/a
             Rhenium produced (contained in APR)                    7,209 kg/a
             Capital cost estimate to First Production              A$337 M (US$337 M)
             Capital cost for Lif e of Mine                         A$518 M (US$435 M)
            * Note >PFS estimates that 50% of Concentrate can be produced at +50% Mo grade

Further information is provided in Section 21 of this report.

Economic Analysis

As part of its economic evaluation, IVA commissioned a report on the outlook for
molybdenum and rhenium markets from Roskill Consulting Group Ltd (“Roskill”).

Roskill expects the molybdenum market to move into deficit from 2013, with prices
thereafter rising to around US$30 /lb Mo (Roskill, July 2011). In its report, Roskill




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forecasts molybdenum oxide prices to increase from an average price of
US$16.50 /lb for 2011 to US$27.00 /lb (US$23.77 /lb in real terms) in 2016.

Roskill forecasts rhenium metal prices to increase from an average of US$4,300 /kg
in 2011 to US$5,500-5,900 /kg (US$4,842-5,194 /kg in real terms) in 2016 (ibid).

IVA has adopted Roskill forecast real prices for molybdenum and the mid-point of
forecast real prices for rhenium from 2011 to 2016 inclusive, with 2016 prices held
constant thereafter together with a forecast AUD:USD exchange rate of 1.00 for the
first three years of the project (inclusive), reducing to 0.83 thereafter.

These evaluation assumptions together with the PFS assumptions yield the following
economic results for the Merlin project:
•     NPV (at 8% real discount rate): A$690 M.
•     After-tax cash flow: A$1,563 M.
•     Internal rate of return (IRR): 32%.
•     Payback Period (from the date of first production) of approximately four years.

Further information regarding the outlook for the molybdenum and rhenium markets
is provided in Section 19 of this report. Further information regarding the economic
evaluation of The Project is provided in Section 22 of this report.

PFS optimisation opportunities

The PFS has identified a number of optimisation opportunities that will be
investigated during any feasibility study (FS) of The Project.

The key opportunities are briefly outlined below.

Mining cost:
•     The PFS has assumed 15 m sublevels in the underground mine. As a
      consequence of further evaluation following completion of the PFS these have
      now been modified to 20 m sublevels which will reduce the mining cost. Any FS
      will be completed on this basis.
•     The PFS focused attention on traditional mining methods. It has been identified
      that innovative mechanical mining may have potential at The Project and thus
      will be included in any FS scope. A mix of long hole open stoping and drift and
      fill stoping will also be considered as part of the optimisation of mining
      methods.
•     Sensitivity analysis which has been undertaken on The Project indicates that a
      10% reduction in mine operating costs is likely to improve NPV and after-tax
      cash flow by approximately A$47 M and approximately A$85 M respectively.

Molybdenum concentrate grade:
•     In this report, for design purposes, it has been assumed that all of the Mo/Re
      concentrate grade will be 30% Mo, however testwork has shown that two




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      concentrate grades at 50% Mo and 30% Mo can be produced. Further work will
      be completed during the early stages of the FS to optimise the product streams

Capital cost:
•     The molybdenum/rhenium sulphide concentrate produced will be processed
      through a       purpose-built, conventional technology,        roasting and
      hydrometallurgical plant to produce saleable products including molybdenum
      trioxide and ammonium perrhenate (APR). The sale of the Mo/Re concentrate,
      the optimisation of the process design to minimise capital cost as well as the
      potential to produce other Mo products including ferro-molybdenum (FeMo),
      ammonium di-molybdenate and sulphuric acid, will be investigated in the early
      stages of the FS. This has the potential to reduce the capital cost of the
      processing plant.
•     This PFS assumed that the Mo/Re roaster and hydrometallurgical plant is
      located at Osborne. Alternative locations in other countries will be evaluated in
      the early stages of the FS to potentially reduce the capital expenditure and
      optimise the return on investment.
      Sensitivity analysis which has been undertaken on The Project indicates that a
      10% reduction in overall capital costs is likely to improve NPV and after-tax
      cash flow by approximately A$31 M and approximately A$35 M respectively.

Life extension:
•     Exploration within the existing tenements where there have been high grade
      molybdenum finds has the potential to extend the life of the Mo/Re project.

Property
The property consists of five Mining Leases containing molybdenum, rhenium and
copper mineralisation suitable for treatment by flotation and further processes. As
The Project underlies the Mount Dore Copper Heap Leach Project, the leases cannot
be identified separately for those two projects. Exploration drilling for The Mount
Dore Copper Heap Leach Project intersected the Merlin deposit and studies of The
Project have been performed in parallel with studies for the development of the
Mount Dore Copper Heap Leach Project.

Ownership
The five Mining Leases are owned by Cloncurry Mines Pty Ltd and are on the
Starcross Pastoral Holding, which also is owned by Ivanhoe Cloncurry Mines Pty Ltd.
Ivanhoe Cloncurry Mines Pty Ltd is a wholly-owned subsidiary of IVA.

Geology and Mineralisation
Mount Dore is part of a wider exploration focus for IVA over a number of tenements
in the same region. This area lies within the Eastern Fold Belt of the Mount Isa Inlier.
The project area stratigraphy is part of the Soldiers Cap Group and Young Australia
Group, which are part of regional Cover Sequence 3.




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Molybdenum-rhenium and copper mineralisation at Mount Dore is hosted within a
tectonised sequence of metashale, metasiltstone, schist and phyllite belonging to the
Proterozoic Kuridala Formation in the Eastern Fold Belt of the Mount Isa Inlier. This
stratigraphic package lies to the west of the over-thrust Mount Dore Granite and
extends north-south along strike for several kilometres and dips eastward
underneath the granite. A massive, easterly-dipping, intensely-silicified quartzite
ridge in the western side of the area forms the footwall to the deposits.

Copper and Mo-Re mineralisation is hosted within variable proportions of
interfingered black carbonaceous and grey micaceous metasiltstone and grey
metashales with thicker beds of phyllite and schist. These metasedimentary units
exhibit recrystallisation textures but retain relict sedimentary features such as
bedding and folding in outcrop.

Status of Exploration, Development and Operations
There has been exploration drilling performed to support mineral resource and
mineral reserve estimates at The Project. A decline has been commenced. There is
no operating mine or mill at the property.

Conclusions
1.    The Project orebody and associated processing facilities have the capacity to
      produce, on average, 5,030 t/a molybdenum and 7.21 t/a rhenium in saleable
      products for a period in excess of 10 years.
2.    Capital cost to first production is estimated at A$337 M and Life of Mine
      A$518 M.
3.    Economic analysis shows the Project is viable for a range of molybdenum and
      rhenium prices. Based on Roskill’s forecast prices for molybdenum and
      rhenium and an AUD:USD exchange rate of 1.00 for the first three years of the
      project (inclusive) reducing to 0.83 for the remaining mine life, NPV is
      A$690 M, after-tax cash flow is A$1,563 M and IRR is 32%.

Recommendations
1.    Continue development of The Project decline to be able to access and develop
      the mineral reserve to be able to mine 500,000 t/y ore by the third quarter of
      2013.
2.    Proceed to Feasibility Study to optimise the mining method and cost,
      commence detailed design of the Mo/Re sulphide flotation plant and roaster
      and hydrometallurgical plant to produce saleable products.
3.    Continue to explore in the region to extend the project life.




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2     INTRODUCTION

2.1       Terms of Reference and Purpose of this Report

This Technical Report on the Merlin Molybdenum-Rhenium Project (“The Project”) at
Mount Dore in North West Queensland, Australia was prepared by Peter McCarthy,
AMC Consultants Pty Ltd (“AMC”) of Melbourne, Australia, John Horton of Golder
Associates, and Tom Hunter, Jacobs E&C Australia Pty Ltd on behalf of Ivanhoe
Australia Ltd (“IVA”) of Melbourne, Australia. It was prepared in accordance with the
requirements of National Instrument 43-101 (“NI 43-101”), “Standards of Disclosure
for Mineral Projects”, of the Canadian Securities Administrators (“CSA”) for
lodgement on CSA’s “System for Electronic Document Analysis and Retrieval”
(“SEDAR”). This report is required as a disclosure of the results of a Prefeasibility
Study.

Heading numbers in this report follow those listed in the instructions for completing a
NI 43-101F1 Technical Report for lodgement on SEDAR.

The Technical Report is based on reports prepared as part of a Prefeasibility Study
(listed in Section 27 References).

Most of the factual text for the Technical Report is drawn from a Prefeasibility Study
conducted during 2010 and 2011 and compiled by Jacobs E&C Australia Pty Ltd.
The study components were originally compiled as follows:
•     Golder Associates - Geology and Mineral Resource estimate.
•     AMC Consultants - Geotechnical, Mining, Mineral Reserve estimate.
•     Metcom Research - Metallurgical test work.
•     Pocock Industrial - Metallurgical test work.
•     Jacobs – Process plant & implementation.
•     Parsons Brinkerhoff - Power study.
•     Metago Environmental Engineers - Tailings test work.
•     Ivanhoe Australia - Infrastructure, organisational structure, marketing,
      operating and capital cost summaries, financial analysis, option studies.

The Golder Associates work is the subject of NI43-101 Technical Report titled
“Mount Dore - Merlin Deposit NWQld, Australia” Number 107631016-011-Rev0 dated
19 October 2010.

For the purpose of this Technical report AMC independently reviewed and confirmed
the sections prepared by IVA.

IVA was provided with a draft of this report to review for factual content and
conformity with the brief.




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      2.2          Qualifications of Consultants

      The individuals shown in Table 2.1, by virtue of their education, experience and
      professional association, are considered Qualified Persons (QP) as defined in NI
      43-101, for this report. The Qualified Persons meet the requirements of
      independence as defined in NI 43-101. Section responsibilities are also listed below.
      Table 2.1            Persons Who Prepared this Technical Report
                         Qualified Persons responsible for the preparation of this Technical Report
                                                                     Date of
  Qualified                                            Independent    Last         Professional
                   Position          Em ployer                                                          Sections of Report
   Person                                                 of IVA      Site         Designation
                                                                      Visit
                                                                                                      1,2,3,4,5, 13, 15, 16, 17
                                                                                                      (except 17.2), 18, 19, 20,
                  Principal                                          15-16
Mr P L                          AMC Consultants                                 FAusIMM (CP),         21 (except processing
                  Consultant,                              Yes       June
McCarthy                        Pty Ltd                                         BSc(Eng),MGeosc       component), 22, 24, 25
                  Director                                           2011
                                                                                                      (except 25.1), 26 (except
                                                                                                      26.1, 26.4), 27
                                                                     22-27
                  Principal     Golder Associates                               FAusIMM (CP),         6,7,8,9,10,11,12,14, 23,
Mr J Horton                                                Yes       April
                  Geologist     Pty Ltd                                         MAIG                  25.1, 26.1
                                                                     2011
                  Associate     Jacobs E&C                                                            17.2, 21 (processing
Mr T Hunter                                                Yes       No visit   FAusIMM
                  Director      Australia Pty Ltd                                                     component), 26.4

      Other experts upon whose contributions the Qualified Person have relied are shown
      in Table 2.2.

      Table 2.2            Persons Who Contributed to this Technical Report
                         Other Experts upon whose contributions the Qualified Persons have relied
                                                                     Independent     Date of Last
      Expert              Position                  Em ployer                                          Sections of Report
                                                                        of IVA        Site Visit
  Mr W
  Sw eetser                                EHP Consulting Inc.            Yes       No visit          13.2, 17.1.5, 17.1.9

  Mr E
                                           EHP Consulting Inc.            Yes       No visit          17.1.8
  Partelpoeg
  Mr V Wagh                                EHP Consulting Inc.            Yes       No visit          17.1.7, 17.1.10
                                           Pocock Industrial              Yes
                    Technical Executive-
  Mr N Jarvinen                            Parsons Brinkerhoff            Yes       1-2 June 2011     18.2.1
                    Electrical
                                           Metago Environmental
  Dr G Mc Phail     Director               Engineers                      Yes       13 April 2011     18.3, 20


      2.3          Site Visits
      Peter McCarthy visited the site on 15 and 16 June 2011. He inspected facilities at
      Osborne and Mount Dore and made an underground inspection of development for
      The Project.

      John Horton visited the site from 17 February 2010 to 22 February 2010 for the
      purpose of the original resource evaluation and revisited the site from 22 April to
      27 April 2011.

      Site visits were also made during 2010 by members of the technical study teams.




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3     RELIANCE ON OTHER EXPERTS

A solicitor’s report dated 19 October 2010 on the status of the Mount Dore mining
tenements was prepared by Minter Ellison Lawyers and was included as Appendix B
in the Golder Associates NI43-101 report dated 19 October 2010. The Minter Ellison
report concluded that the Mount Dore Mining Licenses appear to be in good standing
and unencumbered. AMC has relied upon this advice and on advice from IVA that
lease renewals were subsequently completed and has not sought further advice on
the standing of the tenements.




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4     PROPERTY DESCRIPTION AND LOCATION

4.1       Property Location

The Merlin deposit is located near Mount Dore in west central Queensland, Australia,
approximately 145 km southeast of Mount Isa and 700 km west-southwest of
Townsville (refer to Figure 4.1).

Figure 4.1        Project Location




                                                              Merlin Deposit



                                                                  Osborne Plant



      Source: IVA, with an effective date of 14 July 2011

4.2       Land Tenure

The Cloncurry Project Area is covered by eleven Mining Leases within the          Mount Isa
Mining District, Queensland, Australia. All leases are owned by Ivanhoe           Cloncurry
Mines Pty. Ltd. and are on the Starcross Pastoral Holding, which also is          owned by
Ivanhoe Cloncurry Mines Pty. Ltd. Ivanhoe Cloncurry Mines Pty Ltd is              a wholly-
owned subsidiary of IVA.

The five leases relevant to the Mount Dore resource area, their grant date, expiry
date and area are listed in Table 4.1. All of the leases and the surrounding area are
covered by an exploration permit EPM 10783 that includes all relevant areas at
Mount Dore if the mining leases were relinquished.




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All of the relevant mining leases are valid for the purpose of mining gold, silver,
copper, cobalt, molybdenum, zinc and lead. ML 2566 is also valid for mining a
number of other minerals. Rhenium is not listed as a specified mineral for the existing
mining leases. IVA has applied for rhenium to be added to the mining leases and
AMC understands this should only be a formality. Otherwise rhenium falls with the
encompassing Exploration Permit for Mining (EPM) held by IVA.

Table 4.1             Tenure
                                                                  Area
Tenure Type              Number               Name                         Granted      Expirers
                                                                  (ha)
                          2688    Mount Dore Extended No 1        125.48   21-Jun-79     30-Jun-20
                          2689    Mount Dore Extended No 2        129.6    27-Apr-78    31-May-29
ML : Mining Lease         2690    Mount Dore Extended No 3        129.6    24-Aug-78    31-May-29
                          2691    Mount Dore Extended No 4        120.46    12-Jul-79     31-Jul-20
                          2566    Marilyn 1                       32.37     6-Dec-73    31-May-29
EPM: Exploration                                              238 sub-
                          10783                                            26-Oct-95     25-Oct-12
Permit for Minerals                                            blocks


The boundaries of the mining leases were resurveyed in 2009 to confirm their
location (refer to Figure 4.2). Figure 4.3 shows the drill hole locations and resource
wireframe with respect to the mining leases.




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Figure 4.2       Mount Dore Mining Leases




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Figure 4.3       Mount Dore Mining Leases and Mineralisation




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The annual rentals for the year beginning 13 September 2010 totalled $201,250.10
for all 20 granted mining leases held by Ivanhoe Cloncurry Mines Pty Limited.

Rentals for The Project mining leases (MLs 2566, 2688, 2689, 2690, 2691) made up
$25,731 of this total.

Cloncurry Shire Council rates for the 6 month period beginning 1 Jan 2011 for all 20
granted Ivanhoe Cloncurry Mining Pty Limited Mining Leases totalled $58,597.

Rates for the Mount Dore group of contiguous mining leases (MLs 2566, 2688, 2689,
2690, 2691, 2692, 2693, 2694, 2733, 2745 and 2746) made up $42,479 of this total
and were charged in one combined invoice for the whole group.

Cloncurry Shire Council rates are also payable for Starcross Pastoral Holding
($3,107.54 for the 6 month period beginning 1 Jan 2011), and rent is also payable for
Starcross to the Cloncurry Shire Council.

Queensland State Mineral legislation imposes a royalty on the sale of minerals. In
the case of The Project the royalty for copper is 4.5%. There is no previous owner
royalty or commitments to third parties based on The Project mineral sales.

The Osborne mine and infrastructure are located on Mining Lease (ML) 90040, the
borefield is on ML 90057 and a rail load out facility (adjacent to Phosphate Hill) on
ML 90068. The Kulthor copper/gold underground deposit is on ML 90158. The
deposit and the majority of surface infrastructure are located on Chatsworth Station,
between Carbo Creek and Little Sandy Creek. The borefield is located on Kheri and
Pathungra Station approximately 25 km south of Osborne.




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5     ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE
      AND PHYSIOGRAPHY

5.1       Accessibility

Access is by chartered aircraft via an all-weather airstrip from Mount Isa and
Townsville, or by road from Cloncurry, 140 km to the north. Currently Mount Dore is
serviced by a charter flight from Townville on two days per week and a number of
smaller charter flights to Mount Isa.

Mount Isa is the largest city and main supplies centre for the region, whereas
Cloncurry is a smaller, local supply town. The population of Mount Isa and
surrounding area is about 35,000; and Cloncurry has a population of 2,400.

IVA announced the strategic acquisition of the Osborne copper and gold mine,
processing facilities and tenements on May 25 2010. Osborne is a fly-in fly-out
operation with its employment base at Townsville. Osborne operates charter flights to
and from site to Townsville, currently on four days per week. Supplies are
transported by road haulage from Mount Isa. Services are sourced from Mount Isa or
Townsville.

5.2       Vegetation and Climate

Vegetation consists of arid spinifex and sparse eucalypt trees. The area has a semi-
arid climate with temperatures averaging from 10° to 25°C in the winter and from 25°
to 40°C in the summer. Average rainfall is 350 mm, most of it occurring during the
summer months of December to March.

The climate at Osborne is typical of the inland arid zones of sub-tropical north west
Queensland. The majority of rainfall comes from summer thunderstorms and
decaying depressions drifting down from the north coast. The average annual rainfall
is 320 mm. Temperatures range from extremely hot (average maximum 38°C) in
summer to mild (average minimum 7.6°C) in winter.

The weather is amenable for mining operations year-round, although construction
activity could be affected in the wet season.

The area forms part of the interior lowlands physiographic region with topography
varying between gently-undulating plateau and local hilly terrain. Surface elevations
vary from about 200 to 400 m above sea level. Eucalypt-spinifex grass-woodland
vegetation predominates on deep soils; grass with scattered scrub is common within
flat, open, soil-filled drainages.

5.3       Water Supply

IVA has rights to extract 260 Ml of water from the Mount Dore aquifer. IVA has a
pipeline to the Burke River Bore Field that can supply up to 100 l/s and has annual
rights to extract 1,000 Ml from this field. IVA has a partly filled licensed tailings
storage facility that was in use when the Selwyn Operation closed in 2003.




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5.4       Infrastructure

Power would be supplied by rental power station, with gas trucked from the Osborne
mine site. An access road to the Osborne mine site would be upgraded. Water would
be supplied from nearby boreholes. There is an existing village and aerodrome at
Mount Dore. More details are supplied in Section 18, Project Infrastructure.

5.5       Mining Surface Infrastruture

All mining surface infrastructure is to be accommodated within the existing mining
leases.




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6     HISTORY

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and reproduced here with some exclusions of Mount
Dore specific information which is not relevant to the Merlin Mo-Re resource.

Relevant history for the Mount Dore and the Merlin Zone has been summarised by
Lazo and Pal (2009).

Copper mining commenced in the area at Mount Elliott (20 km north of Mount Dore)
in the early 1900’s. A mining and copper smelting operation was established that was
linked by rail to Cloncurry. This operation closed in the 1920’s. It was during the early
part of the 1900’s that copper was most likely first mined from Mount Dore from
outcropping oxides, although total recorded production up to 1961 was only
5.9 tonnes of copper. Figure 6.1 displays the previous surface workings and with
limited waste dump confirming previous mining was not extensive.

Figure 6.1       Historic Workings at Mount Dore Displays Limited Disturbance




The first modern drilling was conducted in 1957, and only one hole intersected the
deposit, returning 1.3% Cu over 8.8 metres at a depth of 49.7 metres.

Mount Dore was subsequently subjected to a series of company acquisitions,
mergers and joint ventures. The tenement was acquired by Amoco, which later
became Cyprus Minerals Australia Company (Cyprus), in 1975. In 1984, Mungana
Mines Limited (later on Elders Resources Limited) got involved in the project while
Cyprus took on Sunland Petroleum (later Arimco NL) as partner in Australia. The
project in 1990 became a joint venture between Cyprus (50%), Arimco (25%) and
Elders (25%) and briefly closed in 1999 before Selwyn Mines Limited purchased the
property and recommenced operations on the Starra Line and Mount Elliott. This
operation failed after an attempt to increase production to 2 million tonnes per year
from 700,000 tonnes per year. A fall in the copper and gold price and a slower than
expected ramp up in production caused the banks to foreclose on the owners in
December 2002.




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In September 2003 Ivanhoe Cloncurry Mines Pty Ltd acquired the tenements,
exploration data and pastoral station from the receivers of Selwyn Mines Limited.
Other items purchased at the sale were fixed assets, including the water pipeline,
power station and camp.

Between September 2003 and June 2008, IVA completed an additional 19,273 m of
drilling. During a reverse circulation (RC) drilling campaign, aimed at identifying
additional near-surface copper occurrences to the north of the Mount Dore deposit, a
hole encountered significant molybdenite mineralisation.

This prompted a review of drilling data to determine the potential of the area, which
led to the identification of the 700 m long, north-north-easterly-trending Merlin Zone
that contained molybdenite at the northern end and molybdenum oxide at the
southern end. Following the discovery, additional analytical work was conducted to
investigate the associated rhenium content, which returned high values.

Exploration work conducted on the Mount Dore Project is defined in this report as
that within the area bounded by 7,604,000-7,606,150 mN and 447,000-448,000 mE.
The Merlin Zone lies within the northern area of the overall project, from 7,605,000 to
7,606,150 mN, with the Little Wizard ‘bonanza’ shoot at the far south edge of the
Merlin Zone. The majority of drilling on the Merlin Zone has been done by diamond
drilling (DD), and lesser portion by reverse circulation drilling (RC). Most of the
drilling was done by Ivanhoe Australia Limited, and a minor amount by previous
operators. IVA’s work also included twinning of RC and DD holes, which indicated
that the RC drilling underestimates copper and molybdenum grades-one reason why
RC was discontinued.

The history of exploration and mining at Mount Dore includes:
1900’s       High grade surface copper enrichment mined. Total recorded production
             up to 1961 was only 5.9 tonnes of copper (it is unclear if this is ore or
             metal).
1957         Initial drilling was conducted in 1957, and only one hole intersected the
             deposit, returning 1.3% Cu over 8.8 metres at a depth of 49.7 metres.
1976-80      Cyprus completed 40 diamond and 47 percussion drill holes.
1989-91      Cyprus completed 8 diamond drill holes, plus 54 Airtrack and 15 RAB
             near surface drill holes.
1999         Arimco completed 5 RC and diamond drill holes.
2003         Ivanhoe Cloncurry Mines (ICM fully owned by IVA) acquired the Mount
             Dore tenements from the receivers of Selwyn Mines.
2003-7       IVA undertook widespread exploration with some drilling at Mount Dore.
2007         IVA RC drilling completed to define near surface extent of the
             mineralisation at Mount Dore South.
2008         IVA drilling at Mount Dore south targeting secondary copper zone. QG
             completed an initial resource estimate for Mount Dore South in 2008. IVA




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             announced a copper resource of 80 Mt at 0.6% Cu for the Mount Dore
             deposit.
             Discovery of the Merlin high grade Mo-Re zone below the Mount Dore
             North mineralisation. After which most drilling was completed by diamond
             core (except RC precollars)
2009         IVA drilling at Mount Dore North targeting the Merlin mineralisation.
             acQuire database implemented
2010         Scoping study completed for Merlin, (SRK, 2010)
             Completed nominal 50 m infill drilling and definition of Merlin
             Golder resource estimate for Merlin and Mount Dore

Other historic exploration data is summarised in the IVA exploration handover notes
Pal et al. (2010) and includes:
•      767 soil samples.
•      44 stream sediment samples.
•      1991 airborne magnetic survey.
•      1997 airborne EM GEOTEM survey.
•      1996 ground magnetic survey.
•      1978 airborne radiometrics survey.
•      1992 PhD thesis on Mount Dore breccia hosted copper gold deposit.

Exploration completed by IVA includes:
•      Geological mapping at 1:25,000 and 1:1,000 scales.
•      Onsite rock chip geochemistry.
•      Termite mound geochemistry.
•      Sub-audio magnetic (SAM) survey.
•      Ground gravity survey.
•      Three transects of seismic.
•      Drilling as described later.
•      Petrology and petrography.
•      Ground survey as described later.

Historic estimates of the Mount Dore Cu resource are summarised by Golder 2010.

Table 6.1 provides the previous JORC and NI43-101 classified resource statements
released by IVA for the Merlin Mo-Re deposit. These historic resources were
summarised by Golder only to provide a historical reference and are now superseded
by this report. Further details for previous resource estimates are available in the
original ASX market releases by IVA, as referenced.




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          Table 6.1            IVA Previous Estimates for Merlin Mo-Re Resource at 0.3% Mo
                               Cut-off
Release                                 Mo     Re     Cu      Ag
              Classification      Mt                                                     Notes
 Date                                   %     ppm     %       g/t
 2009           Indicated         7.4   0.8    14     0.1     4.7      IVA internal estimate using inverse distance
21 Apr           Inferred         5.6   0.8    13     0.3     5.1        weighting, JORC statement IVA 2009a
 2009           Indicated         5.2   1.0    16     0.2     3.7     Quantitative Group JORC statement IVA 2009b
 9 Nov           Inferred         3.5   0.8    14     0.3     4.4        & published NI43-101 Report, QG 2010


          A qualified person has not done sufficient work to classify the historical estimate as
          current mineral resources or mineral reserves and Ivanhoe is not treating such
          estimates as current mineral resources or reserves.




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7     GEOLOGICAL SETTING AND MINERALISATION

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and reproduced here for reference.

Descriptions of geological setting presented here are based on work by Lazo and Pal
(2009) and Carter et al. (2009). References are contained in these two reports.

Mount Dore is part of a wider exploration focus for IVA over a number of tenements
in the same region. This area lies within the Eastern Fold Belt of the Mount Isa Inlier
(refer to Figure 7.1).The project area stratigraphy is part of the Soldiers Cap Group
and Young Australia Group (which are lateral facies equivalents), and which are part
of regional Cover Sequence 3. The depositional age for these Groups is
1712-1654 Ma, while the main phase of deformation and metamorphism is
considered as equivalent to the Isan Orogeny at 1530-1480 Ma, associated with the
Williams-Naraku Batholith.

Figure 7.1       Cloncurry-Selwyn Zone of the Eastern Fold Belt




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7.1       Regional Geology

For the Mount Elliott and Mount Dore district the tenement package is a highly
prospective locality hosting four advanced projects (Merlin, Mount Dore, Starra and
Mount Elliott) as well as significant copper-gold and molybdenum-rhenium
exploration projects.

The majority of the copper-gold deposits in the area belong to the IOCG style of
mineralisation. The copper-dominant polymetallic zone at Mount Dore has only low
levels of iron oxide and could represent an iron-oxide poor member of the IOCG
spectrum, with fluids and fluids sources similar to the other IOCG deposits. Merlin,
the molybdenum-rhenium mineralised zone is a late phase overprint on the copper
system at Mount Dore.

The deposits in the region tend to focus on geological and structural elements,
namely being in close proximity to: 1) the contact of the Kuridala and Staveley
Formations, 2) regionally extensive north-south trending shear and fault zones (long-
lived faults that have been the principal conduits for fluid flow and alteration), 3) late
NE and NW trending structures, and 4) proximity to a Williams-Naraku Batholith
intrusion.

Figure 7.2, demonstrates the surface relationship. The plan shows the NS linear belt
of the metasedimentary packages of Stavely and Kuridala Formations, in part
bisected by the Mount Dore Fault Zone (MDFZ), increasingly intense NW-NE faulting
around prospect locations. The whole metasedimentary package is buttressed either
on one or both sides by granites, generally in close proximity to the major prospects.

Stratigraphy

Proterozoic lithologies of interest in the project area are the Staveley Formation and
the Kuridala Formation. The Staveley Formation comprises a >2000 m thick linear
belt of shallow water, well-bedded to brecciated variably calcareous, ferruginous,
feldspathic, micaceous, and siliceous sandstone, siltstone, and phyllite, impure
limestone (marble), and lenses of breccia, together with schist and banded calc-
silicate rocks (mainly near granite), and minor basalt lava, conglomerate, and banded
quartz + hematite +/- magnetite rock. The Stavely Formation has a depositional age
of <1720±20 Ma.

The Kuridala Formation is a tightly folded package of moderately deep-water
turbiditic sediments (schistose greywacke, siltstone and shale) with quartzite,
carbonaceous and pyritic slate and calc-silicate rocks. The thickness of this
sequence is >2000 m, although neither the top nor the base of the Kuridala
Formation is well established. The Kuridala Formation has a depositional age of
<1681±5 Ma.

Intrusives

The series of linear batholith scale granitic intrusives are significant lithologies in the
Mount Elliott to Mount Dore district (refer to Figure 7.2). These include the Wonga
Granite (Gin Creek Granite) (1760-1730 Ma) which was                             emplaced




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contemporaneously (into CS2), and the Williams-Naraku Batholiths (Mount Dore,
Squirrel Hills, Yellow Waterhole, Wimberu Granites and SWAN Diorite)
(1556-1504 Ma) outcrop over a large area in the EFB and were emplaced in late syn-
to post-Isan Orogeny time.

Several dykes and sills intrude the Kuridala and Staveley Formations. In Mount
Elliott, distinct bedding concordant basalt intrusives (metamorphosed to amphibolite)
were noted in the host Kuridala Fm. The same concordant dykes and/or sills were
noted in Mount Cobalt, Lady Ella, and Starra.

Tectonics

A complex regional metamorphic history (comprising six sequences) and peaking
prior to the emplacement of Williams-Naraku intrusives has been proposed. The
Mount Elliott - Mount Dore district has been mapped at greenschist facies to
amphibolite facies metamorphism. The basement and cover sequences were
deformed and metamorphosed by the Isan Orogeny (ca 1600-1500 Ma). Up to seven
deformational events in the Eastern Fold Belt have been postulated. Of these
deformation events four major events impact on the project area:
•     D1: Early thrusting.
•     D2: Upright N-S large scale folding synchronous with the peak of
      metamorphism.
•     D3: Folding of D2 folds. Brittle-ductile deformation with intrusion of.
•     D4: Upright N-S folding. Brittle events producing NE - NW faults are integral to
      the formation of many of the deposits in the region.

Structure

District structure is dominated by the regional-scale Starra, Selwyn and Mount Dore
shear and thrust zone, which runs through the core of the district. These NS striking
structures host a number of deposits (refer to Figure 7.2), and include:
•     Mount Dore Fault Zone (MDFZ): A north-south striking deep seated structure
      along the limb of two major regional folds, resulting in sheared sub-parallel
      series of faults which host a number of prospects including Marilyn, Mount
      Dore, Merlin, Flora, Busker and Metal Ridge.
•     Mount Dore Silicified Zone (MDSZ): A north-south striking silicified ridge that
      extends for at least 10 km both north and south of Mount Dore, and broadly
      marks the western limit of the MDFZ. This forms the footwall to the Merlin and
      Mount Dore (refer to Figure 7.3 and Figure 7.4).
•     The Starra Shear Zone is located 2 km west of and parallel to the MDSZ.
•     The Selwyn Line comprises an intensely alkali-iron-silica-carbonate altered
      section of the Starra shear zone, which is host to multiple high grade gold-
      copper shoots.
•     The Selwyn Shear Zone runs parallel to the Starra Shear and the Starra
      deposits area. It essentially marks the Eastern Haematites, a line of ironstones.




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•     Brecciation is a feature of many deposits in the district. Both hydrothermal
      (Amethyst Castle) and faulted and crackle (Mount Elliott) breccias are
      represented as well as combinations and overprints of the all styles throughout
      the region. The abundance of late brittle fractures may have also contributed to
      the development of secondary copper enrichment as noted in the Mount Elliott,
      Mount Dore, Victoria and Lady Ella.




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Figure 7.2       District Geology for the Mount Elliott to Mount Dore - Merlin Area




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Figure 7.3       Aerial View Showing Proximity of the Starra




Ironstone Ridge is where 5 previous mines have operated and the Mt Dore area. The Quartzite Ridge
forming the footwall to the Mt Dore mineralised zone is clearly visible as a sub-parallel ridge to the
Ironstone.

Figure 7.4       Aerial View showing Local Topography of the Mount Dore Area




Alteration

In the Southern Cloncurry District metasomatic alteration events play a fundamental
part in the formation of the majority of ore deposits within the belt, by either being
directly associated with carrying ore minerals or by rheologically preparing rocks for
brittle fracture or ductile shearing. There are two main regional scale alteration
events:
•      The first of these events is a regional Na-Ca metasomatic event.
•      The second regional alteration event consists of large scale K-Fe-Mg alteration.




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Both these alteration assemblages                are commonly found associated with
mineralisation within the region.

Specific alteration types and episodes noted in the district and deposit related are:
•     Iron-oxide alteration early - shear hosted (Starra) magnetite or hematite rich.
•     Iron-oxide alteration later - veined replacive (Mount Elliott) magnetite.
•     Sodic alteration (widespread) - replacive, veined, stockworking and brecciation
      (Mount Elliott).
•     Sodic-calcic alteration - banded, veined, massive and breccia replacive and
      infilled (Starra, SWAN and Mount Elliott).
•     Potassic alteration (widespread) - veins and breccia clasts replacement and
      matrix infill K-feldspar, carbonate and biotite veining (Mount Dore, Merlin, Lady
      Ella and Starra).
•     At least four fluid sources have been proposed as being involved in the
      regional alteration, metasomatism and formation of ore deposits.

7.2       Local and Property Geology

Molybdenum-rhenium and copper mineralisation at Mount Dore is hosted within a
tectonised sequence of metashale, metasiltstone, schist and phyllite belonging to the
Proterozoic Kuridala Formation in the Eastern Fold Belt of the Mount Isa Inlier (refer
to Figure 7.7). This stratigraphic package lies to the west of the over-thrust Mount
Dore Granite and extends north-south along strike for several kilometres and dips
eastward underneath the granite. A massive, easterly-dipping, intensely-silicified
quartzite ridge in the western side of the area forms the footwall to the deposits.

Copper and Mo-Re mineralisation is hosted within variable proportions of
interfingered black carbonaceous and grey micaceous metasiltstone and grey
metashales with thicker beds of phyllite and schist (refer to Figure 7.7). These
metasedimentary units exhibit recrystallisation textures but retain relict sedimentary
features such as bedding and folding in outcrop. Brief descriptions of these units
which appear to have a spatial correlation with mineralisation are included below and
can be related to the example core photographs in Figure 7.5.

Hanging Wall Granite

The Mount Dore Granite, dominating the eastern section of the project area, is part of
the extensive Williams-Naraku Batholith. This intrusion forms the unmineralised
hanging wall to the Cu and Mo mineralisation and conceals the greater part of the Cu
and Mo ore body.

Metasiltstone

The metasiltstone consists of recrystallised quartz grains with a micro mosaic texture
with occasional incipient K-feldspar grain of hydrothermal origin. Another variety of
the metasiltstone appears to have a slightly silky sheen due to fine muscovite




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(sericite) and is probably derived from clayey or muddy protolith. The metasiltstone is
the predominant unit in the metamorphic sequence.

Figure 7.5       Typical Host Rocks for Mount Dore North




Black Shale

The carbonaceous metashales consists of extremely fine-grained oriented muscovite
(sericite) and the dark colour is attributed to the presence of abundant ultrafine
graphite along laminations (generally <5% of rock). The black shale unit occurs
predominantly within the hanging wall to mineralization, stratigraphically below the




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Mount Dore Granite (refer to Figure 7.6). However, it is also observed as a
mineralization host rock, as well as in a footwall position. This unit bifurcates into at
least three (3) discrete lithological units above the mineralization to the north, and is
interpreted to be offset by vertical structures down dip.

Figure 7.6       Geology, Mo & Cu Significant Intersections on Section 7605450 N




Phyllite

The phyllite unit occurs predominantly as a single unit which broadly follows the
geometry of the bounding granite and quartzite (refer to Figure 7.7). Mineralization
can be contained within the phyllite, but also sits in a footwall position from
7,505,600 mN moving south. This unit has been geologically modelled with the view
to using it as a geological marker unit throughout the deposit.

Quartzite

The footwall unit to the Kuridala Formation forms a narrow linear and north-south
trending ridge west of the project area. This massive, intensely silicified ‘quartzite’
ridge, with little internal texture, dips east and is less than 40 m in true thickness.
This zone may define the Mount Dore Fault and may serve as the boundary between
the host metasedimentary package of the Kuridala Formation and the underlying
siltstone and shale units of the Staveley Formation.




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Figure 7.7       Geology and Drill Hole Collars in the Mount Dore - Merlin Area




                                                                  N




7.3       Mineralisation

Descriptions of mineralisation are based on Lazo and Pal (2009) with some
supplementary figures and observations acquired during the site visit by Golder.

An example cross section of Mount Dore North is provided in Figure 7.6. The lower
Mount Dore North copper and zinc mineralisation overlap the high grade Merlin Mo-
Re mineralisation and is described here for completeness.




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The Mount Dore project is situated within the meta-sedimentary rocks of the
Proterozoic Kuridala Formation to the west of and below the Mount Dore granite
body shown in Figure 7.2.

Mount Dore is a polymetallic deposit containing copper, zinc, silver, gold, lead, cobalt
and molybdenum with rhenium in the Merlin zone all within the Kuridala Formation.

The near surface mineralisation has been extensively oxidised such that the copper
rich zones outcrops as copper oxides above a thick zone of chalcocite mineralisation.
The primary zone consists of chalcopyrite, sphalerite, galena and molybdenite as the
visible ore minerals. Chalcocite extends into the primary zone and may also be
hypogene (ie original mineralisation).

Mineralisation dips to the east beneath the Mount Dore granite body in a zone of
about 180 m true thickness.

Surface oxidation of the primary sulphides produced overlapping zones of copper
rich minerals dominated by a suite of secondary copper sulphide and copper oxide
minerals, as well as native (metallic) copper.

The supergene process at Mount Dore involves the conversion of the primary copper
and other sulphides to predominately chalcocite followed by further oxidation to
produce chrysocolla, native copper, cuprite and pseudomalachite.

7.3.1     Copper Mineralisation

The bulk of the currently known Cu mineralisation in the project area consists of
secondary Cu oxides and carbonates (chrysocolla, cuprite, chalcotrichite,
pseudomalachite, minor to trace azurite and malachite) and native Cu after
chalcocite. This oxide zone is underlain by a transition zone dominated by chalcocite
(replacing pyrite, chalcopyrite, and sphalerite) and trace covellite (Lazo and Pal,
2009). The oxides and native Cu penetrate deeper into the transition zone within
major shears and fault zones. Primary Cu mineralisation was emplaced in breccias
and fractures that were best developed in the metasiltstones and black shales and
are only weakly developed in the schists and phyllites.

Two major episodes of Cu mineralisation have been recognised: an earlier
chalcopyrite-pyrite-sphalerite-bornite assemblage emplaced into brecciated
metasiltstone and black shale with associated K-feldspar ± quartz, and a later
dolomite-hosted breccia with chalcopyrite-pyrite sphalerite. Trace to minor galena,
cobaltite, arsenopyrite and molybdenite are noted in the primary sulphide zone. Both
types of primary Cu sulphide became the source of the secondary enrichment zones
for Cu by weathering, after the unroofing of the granite cover by erosion. Very little
gossan is developed.

Note the overlap of copper mineral species in the transitional zone and the presence
of chalcocite, native copper and sulphides in the Primary zone as graphically
indicated in Figure 7.8.




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Figure 7.8       Schematic Cross Section of the Mount Dore Copper Deposit




                            Source: Modified from Pollard and Taylor 2004

7.3.2     Molybdenum-Rhenium Mineralisation

Fracture-controlled and breccia-matrix molybdenite mineralisation is hosted within K-
feldspar-altered and albitised black shales and siltstones, which lie above and below
the foliated schist and phyllite (Lazo and Pal, 2009). The footwall structure at the
base of the foliated phyllite and schist appears to have the strongest Mo and is
inferred to have developed good open structures due to competency contrast. This
basal contact also appears to have acted as good barrier for the Mo-rich fluids, thus
resulting in ponding of the metals in the favourable structures just below the contact.
The mineralised matrix-breccias contain sub-rounded clasts of K-feldspar and clay-
altered siltstone with very minor clay; with molybdenite partially to completely
replacing the breccia matrix (refer to Figure 7.9). Minor patchy pyrite and chalcopyrite
within the matrix are commonly enveloped by molybdenite. The molybdenite breccias
are between several centimetres to several metres thick but most commonly less
than one metre.

Infill drilling has continued to demonstrate a consistent mineralised structure. At
depth additional structures are defined. The style of mineralisation can vary from a
few narrow zones of massive molybdenite to a broader zone of mixed mineralisation.

The high grade Mo-Re mineralisation zone at Merlin comprises one or more narrow
high grade veins between generally 2 and 10 m in width defined over a strike length
of 1000 m and between depths of 60 and 580 m. This is surrounded by weak low
grade mineralisation, predominantly in the footwall. Cu and Zn mineralisation
increase at depth where Merlin overprints the lower Mount Dore North copper zone.
Molybdenite also occurs as stylolitic fracture fill, disseminations and preferential infill
of folded bedding planes. Molybdenite mineralisation in this form is often one to over
10 m thick and is considered continuous between holes 25 to 50 m apart. Narrow
patchy clay zones (illite, montmorillonite) are noted in some of the fractures infilled by
molybdenite.




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There will always be concerns with the short range continuity and how the zones can
be interconnected. Golder has not observed any indication that the mineralisation is
not largely continuous and can be extracted by conventional stoping methods.
Continuity is currently demonstrated on 25 m to 50 m drill spacing. Exploration plans
currently include an exploration decline with further underground drilling to prove up
mineable units.

Mo-Re mineralisation is very strongly correlated. This is partly related to the
occurrence of Re which can largely only occur within the mineral structure of
Molybdenite, the primary Mo bearing mineral.

Figure 7.9 displays an example molybdenite breccia. On the left hand side there is a
silica-feldspar altered metasiltstone clasts within massive molybdenite matrix fill, from
hole MDQ0137; at a depth of 338.8 m. On the right hand side of Figure 7.9 is a
breccia after feldspar-altered carbonaceous slate with porous matrix partly infilled by
molybdenite and hematite, from hole MDQ0147; at a depth of 238.4 m.

Figure 7.9       Molybdenite in Matrix Supported Breccia




On the left hand side of Figure 7.10 is a specimen with pervasively feldspar altered
rock. On the right hand side of Figure 7.10 is a core specimen of massive
molybdenite, of variable thickness, infills fractures and shear planes with some
brecciation in pervasively feldspar altered host. Both samples are from MDQ0147; at
a depth of 231 m; the core in 4b about 6 cm across.

Figure 7.10      Irregular and Discontinuous Fracture-Fill Molybdenite




Further examples of molybdenite occurrence in Figure 7.11 include on the left a
stylolitic texture at the boundary of breccia and metasiltstone and on the right a calc-
silicate host rock, feldspar-altered (K-spar/albite) and silicified, with molybdenite
stylolites.




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Figure 7.11      Stylolitic Low Grade Molybdenite Mineralisation




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8     DEPOSIT TYPES

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and is reproduced here for reference.

Primary copper occurrences at Mount Dore are typical of many hydrothermal copper
deposits in the Kuridala Formation south of Cloncurry. Mo-Re mineralisation is
classified as late-stage hydrothermal mineralisation that overprints the Mount Dore
mineralisation in part.

Mount Dore

The Mount Dore Deposit has polymetallic (Cu, Au, Ag, Zn, Pb, Mo and Co),
secondary copper and Mo-Re mineralisation along the MDFZ. The deposit is hosted
in tectonised carbonaceous metapelites, metasiltstones, schists and phyllites of the
Proterozoic Kuridala Formation. The Mount Dore Granite to the east has been thrust
over the Kuridala Formation, with east-dipping contact forming the hanging wall to
mineralisation. Similar to the Merlin Mo-Re deposit, the polymetallic mineralisation
and the resulting secondary copper zone have been emplaced along breccias
parallel to the granite contact.

The bulk of the currently known Cu mineralisation in the project area consists of
secondary Cu oxides and carbonates (Chrysocolla, cuprite, chalcotrichite,
pseudomalachite, minor to trace azurite and malachite) and native Cu after
chalcocite. This oxide zone is underlain by a transition zone dominated by chalcocite
(replacing pyrite, chalcopyrite, and sphalerite) and trace covellite. The oxides and
native Cu penetrate deeper into the transition zone within major shears and fault
zones. Primary Cu mineralisation was emplaced in breccias and fractures that were
best developed in the metasiltstones and black metashales and are only weakly
developed in the schists and phyllites.

Merlin

Merlin mineralisation is hosted within carbonaceous metapelites and metasiltstones.
The molybdenite mineralisation occurs as breccia infill, disseminations, stylolites and
irregular fracture infill along and adjacent a major fault contact between the
underlying Staveley Fm and overlying Kuridala Fm. Much of the Mo-Re
mineralisation at Merlin, and the overlying Cu mineralisation at Mount Dore, remain
concealed under extensive granite.

The high-grade, northeast-trending and east-dipping molybdenite mineralisation
occurs at the base of the carbonaceous metapelite unit (Kuridala Formation) and
above the calc-silicate banded unit (Staveley Formation). The contact is complex and
this is manifest as several high grade lenses in some holes. Evidence from drilling
suggests an overlap of the molybdenum-rhenium and the polymetallic mineralisation
phases.

Fracture-controlled and breccia-matrix molybdenite mineralisation is hosted within K-
feldspar-altered and albitised black shales and siltstones, which lie both above and
below the foliated schist and phyllite. The footwall structure at the base of the foliated




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phyllite and schist appears to have the strongest Mo and is inferred to have
developed good open structures due to competency contrast. This basal contact also
appears to have acted as significant barrier for the Mo-rich fluids, thus resulting in
ponding of the metals in the favourable structures just below the contact. The
mineralised matrix-breccias contain sub-rounded clasts of K-feldspar and clay-
altered siltstone with very minor clay, with molybdenite partially to completely
replacing the breccia matrix. Minor patchy pyrite and chalcopyrite within the matrix
are commonly enveloped by molybdenite. Molybdenite also occurs as stylolitic
fracture fill, disseminations and preferential infill of folded bedding planes.




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9     EXPLORATION

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and is here reproduced in its entirety for the
convenience of the reader.

The project history and exploration process is summarised in Item 6.

Exploration work is only described in detail for the area relevant to the resource
estimate at Mount Dore and Merlin within the area defined in Table 14.15 and Figure
9.1. This area is subdivided into three distinct mineralised zones, including:
•     Mount Dore South (south of 7,605,000 mN) represents a deeply weathered
      copper dominated zone which has historically been targeted for exploration for
      near surface copper oxide material suitable for leach extraction;
•     Mount Dore North (north of 7,605,000 mN) is less deeply weathered but
      overlain by a depleted zone; hence the copper does not come near surface and
      has only been defined by more recent extension drilling. Polymetallic
      mineralisation with notable Zn mineralisation occurs in the primary zone.
•     Merlin and lower Mount Dore North (lower sequence north of 7,605,000 mN).
      This lower copper and polymetallic (mainly Zn) mineralisation contains some
      lower grade and some very high grade Mo-Re veins referred to as Merlin.
      Discovery and drill definition of the Merlin Mo-Re mineralisation resulted in the
      definition of both the lower and upper Copper zones in 2009 and 2010.

Historic drilling is mostly diamond core and some near surface percussion and open
hole drilling. Drilling is now dominated by IVA drilling. Initial IVA drilling at Mount Dore
south included some RC drilling, but subsequently concentrated on diamond core
drilling when the high grade Mo-Re resource was discovered. RC is still used
primarily for precollars for deeper diamond drilling.

IVA’s work also included twinning of RC and DD holes (MDQ0153A and MDQ0153),
which indicated that the RC drilling appears to underestimate copper and
molybdenum grades.

A map of the drill hole collar locations is shown on the geology plan in Figure 9.1.

Drill spacing at depth from the historic drilling is too broad to consistently track the
location of the higher grade mineralised zones such that in the past resources have
been estimated within a grade shell at a cut-off grade of 0.2% Cu. To advance Mount
Dore, the drill spacing has been closed up with step outs of between 50 m and 100 m
down dip and across strike from the highest grade existing holes.




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Figure 9.1        Plan of Resource Drilling




                                                                               N




   Resource Model Limit (solid black line), Extend of Indicated Resource Classification (dashed magenta line).
   Source: Golder, 2010.

A diamond drilling program over the secondary copper zone and the northern
primary copper zone at Mount Dore commenced in late 2007, intersecting significant
Cu mineralisation.

Inspection of the drill locations had shown that a large area up dip and along strike
from the main Mount Dore copper mineralisation had not been adequately tested. A




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few open hole percussion holes from the 1970s had been drilled to shallow depths
indicating some weak copper mineralisation in the area.

To test this area on a broad scale an RC percussion program was also carried out to
define the near surface limits of the secondary copper mineralisation and to test the
large area to the north-west of the main Mount Dore body. In total 36 holes were
drilled in this program for 4,566 m.

It was during this drilling that MDQ0153 located strong molybdenum mineralisation at
the very base of the Kuridala sequence beneath copper oxides and sulphides.

Drilling to the north of the main secondary copper zone has confirmed the presence
of significant primary copper, zinc and silver mineralisation with elevated
molybdenum values. RC drilling has found significant shallow secondary copper
mineralisation and molybdenum at depth in MDQ0153 on the northernmost traverse
drilled at Mount Dore. This intersection was twinned by diamond hole MDQ0153A.

Figure 9.2 displays drilling activities and a panorama view of Mount Dore viewed
from around the middle area of Mount Dore North. Looking west the quartzite ridge
occurs in the footwall of the mineralisation. Granite hills lie to the east, behind the
view provided, overlying the mineralisation in the hangingwall. The road traverses cut
into the eastern side of the ridge allowed vertical RC drilling to test the near surface
copper oxide continuation up dip in an area that could not be drilled with angle holes
due to the steep topography.

Figure 9.2       February 2010 Merlin Drilling and Site




                            looking South (left), West (middle) and North (right)

MDQ0153 drilled into the flat area in the middle foreground and discovered Merlin;
other RC holes on the flat were then extended with diamond tails to intersect the
deeper Merlin mineralisation.

During the definition and geotechnical drilling completed to assist the exploration
decline development plan, a small very high grade zone was discovered in mid 2009.
Called Little Wizard, this small zone has now been followed up with 13 mineralised
drill intersections on roughly a 12.5 m drill spacing.

Figure 9.3 displays an example of the massive molybdenite mineralisation with Table
9.1 displaying the current resource intercepts, Zn and Pb values are low and not
included.




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Figure 9.3         Massive Molybdenite in MDQ0264 (Little Wizard Zone)




 This interval 0.9 m @ 46.40% Mo, 446 g/t Re, 1.40% Cu, 0.29 g/t Au, 25 g/t Ag (part of 3.73m @ 15.5% Mo, 95 g/t
                                      Re, 1.4% Cu, 0.2 g/t Au, 26 g/t Ag)

Other exploration work by IVA consisted of termite mound sampling, which outlined
copper and molybdenum anomalies along the stratigraphic belt between the
overlying granite and underlying quartzite. Prior to IVA’s work, Selwyn Mines Limited
(Selwyn) conducted a CSAMT survey (Controlled Source Audio-Frequency
Magnetotellurics) over the entire area, which didn’t return any recognizable
anomalies; however, when IVA re-examined the data and implemented refined data
processing techniques, a broad north-north-easterly trending anomaly that roughly
coincides with the Merlin Zone was outlined. IVA also completed a small SAM (Sub-
Audio Magnetic) and gravity survey over part of the area.

Table 9.1          Little Wizard Resource Intercepts
                                  Depth        Mo       Re       Cu      Au       Ag       Co       Fe       S
   DHID         Length
                            From          To    %      ppm       %      ppm      ppm      ppm       %        %
MDQ0264           3.73      90.85    94.58     15.5     151      1.4     0.2       26      39       0.6     10.4
MDQ0276           1.1        96       97.1     11.8     130      2.4     1.4       39      11       2.3     7.5
MDQ0278            3         87           90   7.8      107      1.4     0.5       25       6       1.3      5
MDQ0279           2.6       90.5      93.1     7.4      96       0.7     0.2       17       4        1      4.6
MDQ0279           0.6       93.3      93.9     15.8     166      3.5     0.3       35       7       0.3     10.5
MDQ0280           1.9       93.3      95.2     22.9     282      8.4      1        55      37       1.5     13.2
MDQ0280           3.6       95.4          99   5.1      65       1.6     0.6       13       9       1.9     3.5
MDQ0281            2         90           92   0.5       3       0.3     0.2       23       3       2.5     0.2
MDQ0287           0.7       74.8      75.5      2       23       0.9     0.5       24       1       0.5     1.3
MDQ0287           0.55      75.75     76.3     26.4     384      3.8     2.4       40      11       0.8      10
MDQ0287           3.6       76.4          80   10       164      3.2     0.8       37      13       3.6     4.8
MDQ0288            8         80           88   2.1      22       2.3     0.7       19      31       1.4     1.9
MDQ0290           2.5       75.5          78   0.6      10        0      0.1       28       1       0.9     0.3
MDQ0335           1.5        82       83.5     0.1       0       0.1      0        1        3       2.9      0
MDQ0386           2.8        76       78.8     0.5       0       2.5     0.1       26      46       1.3     1.2
MDQ0387           5.8        86       91.8      5       61       3.8     1.1       60      26       1.3     3.9




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10     DRILLING

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and is here reproduced in its entirety for the
convenience of the reader.

The mineralisation at Mount Dore and Merlin is largely restricted to the Kuridala
Formation and internal and bounding structures. These roughly dip at 50° to the east
and can be as steep as 70° or as shallow as 30°. The Merlin high grade Mo
mineralisation forms narrow planes or zones typically only a few metres in width
which averages 3.9 m in true width and varies between 1 and 20 m (note a minimum
2 m width is applied for resource domaining). This is contained within a low grade Mo
mineralised dipping zone around 50 m in width. For the copper the higher grades
form roughly planar zones around 4 to 10 m in width and are included within the
lower grade envelope used for resource estimation of between 5 and 30 m in width at
Mount Dore North and a more amorphous shape at Mount Dore South of between
30 and 150 m in true width. Figure 14.17 to Figure 14.19 provide examples of the
mineralisation widths modelled.

Most drill holes are vertical or are inclined up to 60° to the west. This results in
oblique intersections in most cases and apparent intersection widths much larger
than the true width of the mineralisation.

Typical dips of inclined mineralisation of 50° and drilling inclination of 70° would
result in down hole intersections of 10% longer than true width. This scenario is
common for the upper Cu and Mo-Re resource but increases with depth when
targeting the narrow high grade Mo-Re resource where both steeper structures and
occasional vertical drilling can result in down hole lengths up to 100% longer than
true width. All resource estimation was undertaken using 3D modelling methods that
account for the drill intercepts in true space and the different intersection lengths
achieved by the drill orientation. The drill orientation approach undertaken by IVA is
the most practical given the terrain and target depth.

10.1      Drilling Methods

Drilling by pre-IVA companies included a range of drilling types and sample
procedures.

Cyprus completed 32 diamond holes between 1976 and 1980 and used the data
generated to calculate some preliminary geological resource estimates for the
prospect. A number of open hole percussion holes were also drilled in the area.

In 1989 to 1994 Cyprus completed a program of 16 diamond holes and 2 water bores
in the Mount Dore prospect.

Arimco undertook a significant part of the Mount Dore drilling in 1999-2000. To test
near-surface open pit potential, 54 Airtrack and 15 RAB holes were drilled. In
addition 8 diamond holes and 7 RC water bores were sampled. 43 RC holes in 2000
added to this substantial campaign that tested the upper 50 m of the mineralisation.




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This material was deemed unsuitable for the Selwyn plant due to the presence of
malachite and chrysocolla, and the proposed open pit scenario was found to be
uneconomic. Initial drilling by IVA from 2004 used diamond core for definition of the
Mount Dore mineralisation.

An RC program was carried out at Mount Dore south to help define the near surface
limits of the secondary copper mineralisation.

Subsequent to the 2008 RC program drilling has been targeted to intersect expected
mineralisation with diamond core drilling. This preference has been accentuated
since the discovery of the Merlin narrow high grade Mo-Re zone. Subsequent drilling
has been by diamond core except for RC precollars which are restricted to areas
expected to comprise hangingwall waste, i.e. generally granite.

Typically holes cored from surface start at PQ, reducing to HQ and eventually NQ as
the drilling conditions become more difficult with depth.

The discovery of the Merlin deposit in December 2008 changed the emphasis of the
exploration program from shallow infill of the Mount Dore South copper oxide
resource to progressive extension and infill of the Mount Dore North and underlying
Merlin deposits. This resulted in drilling progressively becoming deeper and stepping
out further to the north. The focus on Merlin and Mount Dore North (north of
7605000 mN) required RC precollars where there is significant unmineralised
hanging wall granite expected. Table 10.1 summarises the exploration programs and
drilling methods, and is restricted to only the exploration data relevant to the Mount
Dore and Merlin resources, excluding exploration drilling at other nearby prospects.

Table 10.1        Summary of Drilling Companies, Date and Method
                                    Number of Drill Holes                 Total Drilling Length (m)
Year    Company
                             DDH      WB      RAB/AT        RC    DDH         WB      RAB/AT           RC
1976    Previous Company       8                                  1147
1977    Previous Company      15                                  4729
1978    Previous Company       8                                  3486
1979    Previous Company       1                                  351
1989    Previous Company       8        2                         2472        472
1992    Previous Company       2                                  652
1993    Previous Company       2                                  611
1994    Previous Company       4                                   306
1999    Previous Company       5                 54               1266                  1138
2000    Previous Company       3        7        16         43     725        955        626          4229
2004    IVA                   17                                  3561
2007    IVA                   19                                  7443                                1196
2008    IVA                   67                            38    22938                               11179
2009    IVA                   115                           1     39908                               7815
2010    IVA                   34                                  11800                               2296
Total                         308       9        70         82   101393      1427       1764          26715


The recovery is dealt with in Sections 10.4 and 11.7.




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10.2      Collar Surveys

For the recent drilling by IVA, drill hole set-outs were by hand held GPS, then picked
up by a contract survey company from Mount Isa using differential GPS over a
number of campaigns from July 2009. All IVA drill holes were surveyed with the
exception of 5 drill holes from 2004 that were buried and could not be located. One
hole, started on 25 June 2010 for the most recent drilling had not been accurately
surveyed. Since no sampling of that hole had been completed at cut off of 28 June it
is not considered in the resource drilling subset.

Historic drill hole collars (pre-IVA, from the 1970’s onwards) were picked-up using
traditional theodolite-based methods. Table 10.2 summarises the collar survey
method for the drilling within the Mount Dore resource area. Collar survey data and
methods are considered appropriate for detailed resource evaluation. The majority of
the hole collars in the Merlin Project area have been accurately positioned using the
prevailing industry standards.

Table 10.2       Collar Survey Method
                     Company      Drill Type         Survey Method       Holes    Holes
                     IVA          DDH & RC           Surveyed             286      61%
                     IVA          DDH                GPS coordinates       5       1%
                     pre-IVA      DDH                Surveyed              5       1%
                     pre-IVA      DDH                Unknown*             51       11%
                     pre-IVA      RC & percussion    Unknown*             122      26%
                 * In most case the historic data tagged as unknown was surveyed by theodolite

10.3      Down Hole Surveys

Recent down hole surveying techniques at Mount Dore have been performed by an
electronic Reflex instrument (‘EZ-SHOT’) survey tool which operates in both single
and multi-shot modes. Generally survey shots are taken initially at 30 m intervals in
diamond core holes, but at much greater intervals (or sometimes even not at all) in
the RC pre-collars. The single shot surveys are written down at the drill site and
manually entered into the acQuire database. This device relies on magnetism to
determine the drill hole azimuth, so it will be affected by magnetic minerals. Since
2009 the same survey tool is regularly used to resurvey the drill hole in multi-shot
mode at even 6 m intervals down hole. This data is to the acQuire database.

Pre-IVA, the most common form of down hole surveying was by Eastman single shot
camera. However, as there appears to be very little magnetite, pyrrhotite or other
magnetic minerals at Merlin or the oxide zones at Mount Dore, the azimuths should
be quite reliable. A small number of holes drilled in 1999 in the southern part of the
deposit were surveyed with a Maxibor instrument, with good correlation to the
magnetic single shot surveys during drilling.

Validation of the down hole surveys was undertaken by IVA based on several criteria
including survey intervals greater than 50 m and survey deviations in dip or azimuth
>3°. The down hole surveys were reviewed and historical records checked.




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Two programs of gyroscopic down hole survey were undertaken to both resurvey
earlier suspect IVA drill hole surveys identified in the internal review and also to
locate some of the deeper Merlin drilling more accurately. These were surveyed
using hired GyroSmart (Reflex) and Multishot Ez Track camera (Reflex) and in 2010
a Humphries SR Gyro operated by Surtron Technologies. A total of 24 drill holes
were check surveyed in late 2009 and 57 in mid 2010.

The implementation of the acQuire database allows the replacement of earlier down
hole surveys with more accurate methods by using a priority system. The
implementation of the survey review by IVA and resurveying of many drill holes has
corrected or removed many of the down hole surveys that were previously suspect or
inadequate.

Table 10.3 summaries the down hole survey method for the drilling within the Mount
Dore resource area. There remain some deep drill holes which rely heavily on down
holes surveys that may be too widely spaced to accurately locate the drill hole
location at depth. With drilling as deep as 836 m there will always be issues with
locating drill holes for accurate design purposes. Typically the vein locations are
redefined by underground drilling and mine development. Down hole survey data and
methods are considered appropriate for resource evaluation.

Table 10.3           Down Hole Survey Method Summary
  Survey              Company
                                                         Down Hole Survey Comments
  Method       IVA       Pre-IVA
                                   Collar Sighting using a compass (note this is typically used at the collar to
 CollSight     288          178
                                   start the drill hole trace inclination)
  Eastman      129          5      Eastman single shot camera
  GyroSR       1603                Gyroscopic nonmagnetic instrument
  Maxibor                   393    Maxibor nonmagnetic instrument (1999)
  MSFlexi      4986                Multishot Reflex Ez-Shot (from May 2009)
  Plotting      36          2      Planned orientation
  SSFlexi      844                 Single Shot Reflex (to May 2009)
 Unknown        1           366    Unknown, probably mostly Eastman surveys for historic holes


10.4         Recoveries and Rock Quality

QG, 2010 noted that much of the core from the southern section of Mount Dore is
very broken, with very low Rock Quality Designation (RQD), but the quality of the
core appears much better in the Merlin Zone. The rock at Merlin and Mount Dore
North is generally less weathered and more competent, with less shearing and
faulting than further south, and core recovery is significantly greater as a
consequence.

IVA record RQD and other geotechnical information as part of their routine core
logging. Orientated core is available for some of the drilling that is not vertical and
where reliable orientations could be obtained. The core is sufficiently broken that only
some core is orientated.




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11     SAMPLE PREPARATION, ANALYSIS AND SECURITY

This section is unchanged from sections in the Technical Report prepared by Golder
in October 2010 (Golder, 2010) which were headed “Sampling Method and
Approach” and “Sample Preparation, Analyses and Security” and are here
reproduced in their entirety for the convenience of the reader.

11.1      Methods

All significant sampling conducted on the Mount Dore Project was done using
diamond drill core or RC cuttings. IVA drilling at Mount Dore dominates the available
resource data and has had a total of 329 holes drilled with 290 relevant to the
Mount Dore resource, of which 131 were diamond core, 119 were diamond core with
RC precollars and 40 were RC. Two thirds of the IVA drilling is in the Mount Dore
north area which concentrated on the definition of the Merlin resource at 50 m grid
spacing and some closer infill at Little Wizard. The IVA drilling at Mount Dore
includes 36,700 samples of which 35,000 were included in the Mount Dore - Merlin
resource estimate and the remainder at nearby exploration prospects.

Many RC precollars commence in the overlying granite that was not sampled if there
was no visible mineralisation. IVA samples were taken from HQ core (77%), PQ core
(10%), RC cuttings (12%) and NQ core (1%). Most of the samples were from HQ
core, which was split along the long axis, and one-half of the core over 2 m intervals
made up each sample.

All sampling of diamond drill core and RC cuttings was conducted by IVA personnel
at IVA’s Mount Dore core processing facility several km from the drill sites. Core
recovery within near-surface oxide zones can be quite poor, but below the oxide
zone, overall core recovery was excellent, and only few broken zones were
encountered that could impact accuracy and reliability of results.

Sample quality and procedures in this section concentrate on the recent IVA work
which dominates the resource database. Historic drilling procedures and quality
sampling is poorly documented. IVA is currently indexing a large volume of work
reports from the old Starra mine which may provide further details for the historic
drilling and sampling.

11.2      Procedures

11.2.1    Sample Dispatches

Prior to conducting sampling of diamond drill core and RC cuttings, Sample Cut
Sheets, created by logging geologists or by QAQC technicians, were delivered to
Core Shed Supervisors. The Sample Cut Sheets were used to guide samplers in
compiling dispatches for shipping. These sheets are then used to create the Sample
Shipment Memos (SSMs), which provide a set of instructions for the laboratory. Each
Sample Dispatch sheet comprises details for 70 routine and QAQC samples and
corresponds to one bulka bag for shipping purposes. Dispatches were assigned
numbers that correspond to the first sample in the dispatch and were divided into two
batches that also were assigned numbers corresponding to the first sample. Each




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dispatch contained the following samples for which assays were reported on a
regular basis:
•     59 routine samples (29 per first batch and 30 per second batch).
•     4 standard reference materials (2 per batch, randomly inserted at site).
•     2 field blanks (one per batch, randomly inserted at site).
•     2 pulp duplicates (one per batch, inserted at laboratory in fixed position).
•     2 coarsely-crushed duplicates (one per batch, inserted at laboratory in
      fixed position).
•     1 core duplicate (one in first batch, randomly inserted at site).

In addition to the above samples, pulp duplicate samples were made-up by the
laboratory and returned to the work site to be used for independent check sampling.

All samples were put into cloth or plastic sample bags with numbered tags, and two
of these bags were placed together into a larger plastic bag that was put into a
security sealed woven bulka bag for shipping to the laboratory. When sampling of a
dispatch was completed, Sample Dispatch Sheets and SSMs were given to the
QAQC Technician who arranged transportation from the core shed to the ALS
Laboratory.

Additional quality control measures were implemented during the drilling program:
reduced usage of cloth sample bags by using plastic bags to eliminate dust
contamination issues; flushing of all sample preparation equipment after mineralised
samples; and flagging of high-grade samples.

11.2.2    Sampling of Diamond Drill Core

Prior to logging, the following was done: all driller’s core metreage markers were
checked for errors; core was pieced together; recovery was determined for runs
between markers; metreage lines were drawn across the core and onto box rungs
where depths were written; a continuous line was drawn along the long axis of the
core to guide splitting to minimise sampling biases; and photographs were taken of
all core boxes, with core dry and wet, to facilitate future checking work. After logging,
the core was cut in half along or beside the cutting lines using a rock saw flushed
with water, and it was placed back into the core box with the cutting surface oriented
vertically.

Individual samples were marked at 2 m intervals down drill holes throughout zones
that were designated for sampling by logging geologists. Sample numbers were
written at the start of each sample interval after the metre mark on box rungs on the
left side of the core as one looks down-hole. The right-hand half of the drill core, as
one looks down-hole, was collected for assaying, and if a core duplicate was to be
collected, then it was taken using the left-hand half of the drill core. If core was
excessively broken, a stainless steel spoon was used to collect one-half of the chips
along the side.




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11.2.3    Sampling of Reverse Circulation Cuttings

The cyclone output from one metre intervals was split at the cyclone to obtain an
eighth (10 to 15%) split subsample, which was placed in cloth bags. Because a
splitter was not available, for some of the earlier samples, they were bagged in their
entirety, and then several spear samples were taken from each bag and placed
together in cloth bags. The cloth bags were allowed to dry; the lumpy portion of the
cuttings was broken inside the bag; the bag was tossed several times to mix the
cuttings; and, finally, the bag was weighed. A scoop, comprising about 250 ml, was
taken from each cloth bag for compositing, and scoops from two consecutive
intervals were combined for a two metre composite. The composited samples were
compiled into dispatches as described earlier.

11.3      Bulk Densities

Investigation of the bulk density database has revealed more density determination
methods are present than previously described. Work is ongoing to more fully
describe the available data.

The description of the density measurement process undertaken by IVA from 2007 to
mid 2009 is as follows.

IVA undertook bulk density determinations (BD) at regular intervals of 5 to 10 m. A
10 to 20 cm length of core was used, with the selected BD sample halved to allow
measurement while the rest of the core was sampled and processed. The BD
determination method used a wax dip approach where the samples are weighed dry
and during immersion in water to determine their bulk density relative to that of water.
The BD samples are not generally returned to the original core tray but retained
separately in a specimen library. A correction is made for the buoyancy and mass of
the paraffin wax. The method is a high quality approach well accepted and used in
the industry, but was discontinued in mid-2009 as it was found to be time intensive
and measured relatively small sticks of core after cutting. QG (2008) reviewed the
process used at the time and considered the method, scale calibration, reference
measurements (a piece of basalt in every 10 samples), data handling were all
suitable, though some improvements could be made.

The systematic density measurement process was discontinued during mid 2009 due
to a number of reasons including, safety concerns with the use of hot wax,
inadequate coverage of the narrow Merlin zones and concerns over the adequacy of
small specimen measurement for the Merlin mineralisation which varies in sulphide
content. Trials of different density determination methods were undertaken in late
2009 and early 2010. From mid 2010 a program of density measurements was
resumed using a standard water immersion method.




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The acQuire database supplied to Golder from IVA contained 10,390 density
measurements for the area relevant to the Mount Dore resource. The existing data is
appears to fall into three groups, one of which correlates to the available description
for the project. These include:
•     6,645 density values with weight data for dry, waxed and wet measurements.
      These correspond to the standard measurements taken by IVA using the
      described wax method process by IVA for only IVA drilling. The database
      contains weight data that supports the density values, using an assumed wax
      density of 0.8 m/t3 which corresponds to paraffin wax. This wax was changed
      to a different wax with a lower more safe melting temperature which has a
      density closer to 1 m/t3. This yellow wax was sighted by Golder. The change
      over period is currently under investigation so that the calculated density
      values can be corrected. This small correction is not likely to be material to the
      current estimate.
•     1,507 density values with weight data for wet and dry measurement. These
      appear to correspond to measurements taken in 2008 using a water
      immersion method by IVA for only IVA drilling plus a few older drilling holes.
      There are 31 density values within this set that do not correspond to the
      supporting weight data. The rest are supported by recorded weight
      measurements. The source and description of this data is currently being
      investigated. The data appears to be consistent with the other available data
      and considered suitable for resource evaluation.
•     2,238 density values with no supporting weight data for only older drilling
      between 1997 and 1994. This data is not documented but is presumably older
      density records that were uploaded to the current database. IVA is currently
      determining the source of this data. It is consistent with the other available
      data and is retained for resource purposes as it covers historic drilling areas.

Additional density data was provided to Golder separately for recent work and is yet
to be imported into the acQuire database. This includes two sets of data, including:
•     A trial program of bulk density methods was undertaken in late 2009 by IVA to
      derive a method more suitable for measuring drill core at Merlin, which can
      contain both massive molybdenite and clayey material, see Hillier, 2009. This
      program used several techniques and confirmed the previous waxed
      immersion method was suitable. It also indicated other methods performed
      comparably, including water immersion, direct volume measurement and
      vacuum sealing. This trial program indicated that density derived from
      stoichiometric calculations were generally too high. The trial program included
      multiple subsamples from some Merlin and Little Wizard Mo intercepts from 58
      specimens. These correspond to 30 separate sample intervals.
•     Bulk density data collected in 2010 has used a standard water immersion
      method with data collected for 893 measurements. The measurements were
      obtained for existing core and cover a range of drilling, including pre-IVA
      drilling.
In 1,687 cases there are multiple density measurements for specimens within the
same sample interval. Comparison of these density pairs provides a measure of




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variability presented in Table 11.1 as the half absolute relative difference (HARD).
The relatively low HARD values for all the comparisons, indicates reasonable
precision for each method. There are insufficient comparisons between the four main
methods or periods of data collection to draw any statistically significant conclusions.
A plot of all the data comparison in Figure 11.1 indicates generally a tight cluster of
data around 2.5 t/m 3 with no evidence that any particular data sets are significantly
different.

Table 11.1            Comparison of Bulk Density Pairs from the Same Sample Interval
Method Comparison                         Pairs      Mean 1       Mean 2   Difference   HARD
2008-2009 wax duplicates                  318         2.70         2.56       NA        1.46
2008-2009 wax vs. 2010 immersion           8          2.62         2.62      0.01       1.04
2008 immersion duplicates                  55         2.53         2.41       NA        1.38
2008 immersion vs. 2010 immersion          5          2.48         2.43      0.05       1.37
historic duplicates                       702         2.52         2.38       NA        1.46
2010 immersion duplicates                 599         2.63         2.64       NA        0.73



11.4        Magnetic Susceptibility

Magnetic susceptibility is measured for each 1 m of drill core. This has previously
been used to identify the red rock unit in the lower sequence of the Kuridala
Formation that contains the mineral magnetite. Data entry errors were noted in data
collected between March 2008 and December 2009. These errors were largely
corrected but imply some doubt as to the integrity of the data available.

Nonetheless, magnetic susceptibility values can be grouped into readings that are
high and low. It can be seen that the area footwall to the molybdenum mineralisation
has a high magnetic susceptibility signature. The contact is approximately the
contact between the hangingwall carbonaceous meta-sediment and underlying calc-
silicate, see Figure 11.2.

11.5        Logging

Geological logging comprises recording the core size, sampling intervals, the
lithology zone (lith zone), the dominant lithology (lith1 and lith2), major silicate
minerals, the dominant alteration and alteration minerals, the opaque minerals in
order of abundance, brecciation and the nature of clasts and matrix, the oxidation
horizon and the major geotechnical properties which includes rock strength along
with the joint breaks and fracture logging. Measurement of vein characters and
structures is dependent on the availability of the down hole drill / orientation lines.

Drill hole logging is direct to pocket acQuire on laptops. Validation of lithology,
mineralisation etc. codes is done upon data entry, and the data is then downloaded
onto the main server every day, reducing the risk of data loss on the laptops.

Logging to date has been setup to log a defined lithological sequence. Recent
petrological work indicates greater influence of magmatic units, magmatic fluids and




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alteration being responsible for some if not all of the variations previously noted
within the Kuridala formation. Relogging and reworking of the logging has been
commenced and only completed for 100 m spaced sections to date.

Figure 11.1 Scatter Plot of Pairs of                Figure 11.2 Cross Section Showing
Dry Bulk Density Measurements for                   High Magnetic Susceptibility Zone on
the Same Sample Intervals                           the Footwall of the Merlin

                                                                              Granite




                                                                  Quartzite



11.6      Database Management

IVA has been using acQuire as the company geological database since 2009.

Assay data is imported electronically from the laboratory (ALS). Assay priorities are
assigned by assay method if different methods were used for the same interval.
“Best” assay is the highest priority. Repeats are not averaged to produce the “best”
assay; which is good practice. The original two-acid digest used by ALS (OG46) was
not performing properly leading to a low-bias for the molybdenum and rhenium
assays by up to 20%. This method was replaced with a four-acid digest method
(OG62) where Mo values are above 500 ppm. These different methods do not affect
the copper assays. Further information is provided by Sketchley (2009).

Down hole surveys have priorities based on method used, but at Mount Dore the
main method used so far is by electronic single shot. The results of the down hole
surveys are written onto survey forms by the driller, and entered by hand into the
database (seven degrees is added to the magnetic azimuth to align the reading with
the local grid). QG (2010) noted there is potential for transcription errors, but the data
is also available electronically, so further validation is possible.

Collar surveys also have priorities commencing with drill hole set-outs by GPS which
are then superseded when picked up by a contract survey company using differential
GPS.

Drill hole logs are generally entered directly into acQuire loggers at the core shed
allowing the entry forms to perform basic validation during logging.




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QAQC data routinely loaded into database with site personnel dedicated to QAQC.

QG (2009) audited the data base for 7 IVA drill holes (approximately 5% of the 110
core and RC holes drilled at Merlin) checking the original electronically-supplied
assays from the laboratory against data in acQuire database. Also the bulk density
measurements and down hole surveys readings in the database were checked
against the original worksheets and documents. Comments from this audit indicate
the database management was acceptable with only minor issues such as assay
rounding, a few minor transcription errors, a missing correction for magnetic
declination in one hole and some manual smoothing azimuths for vertical down hole
surveys and one density calculation error.

Golder has not audited the database at this stage due to timing issues but considers
a more formal audit of both new and old drilling data is required in preparation to the
feasibility study.

11.7      Adequacy of Sampling

There were no diamond drilling or core recovery factors that might have resulted in
sampling biases for samples from competent rocks below the oxide zone. In addition,
the sampling procedures that were set up and followed for diamond drilling have
provided samples that adequately represent the whole original drill core.

However, for RC drilling in incompetent rocks, poor recoveries could affect how
representative grades are. Although modern RC drilling systems typically use face-
centred bits to eliminate contamination by channelling cuttings back up the centre of
the drill stem, recoveries are variable and low enough that they could affect the final
assay results. This is supported by analytical results from twinning of RC and
diamond drill holes that indicate a significant low bias for the RC drilling samples,
which is discussed in a later section of this report.

As there was a tendency to drill more vertical to steeply-inclined holes, recovered
lengths of some mineralised intersections are greater than the true thicknesses.
There could be grade biases related to drilling non-orthogonal intersections.

11.8      Quality Assurance and Quality Control Procedures

Quality control procedures used by IVA comprise inserting of Standard Reference
Materials (SRMs), Field Blanks (FBs), and duplicates (DPs) into sample dispatches.
In addition, the analytical laboratory used internal reference materials and pulp
replicates. SRMs are used to measure accuracy; FBs, to check for contamination
and mix-ups; and DPs to monitor precision at several stages of sample preparation.

SRMs used by IVA for exploration work on the Mount Dore project are either
commercial standards purchased from Geostats Pty. Ltd. (GSTAT), Ore Research
and Exploration Pty. Ltd. (OREAS), and CDN Resource Laboratories Ltd. (CDN), or
matrix-matched standards developed by Ore Research and Exploration Pty. Ltd., and
CDN Resource Laboratories Ltd. A total of 20 different SRM types were used for
2,285 assay determinations, comprising several elements (refer to Table 11.2). The




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SRMs comprise series for oxide copper; sulphide molybdenum; sulphide copper and
gold; and sulphide molybdenum, rhenium, copper, and gold.

Table 11.2       Mount Dore - Merlin - Standard Reference Materials, 28 June,
                 2010
 ICM            Supplier             Au          Cu         Mo       Re      Number   Material
 Name            Name               (g/t)        (%)        (%)     (ppm)    SRMs      Type
MDL-1      GSTAT GMB307-15            -         0.389        -        -       638       Ox
MDM-1      GSTAT GBM307-15            -         1.092        -        -       551       Ox
MDH-1         GSTAT BM7               -         1.397        -        -       211       Ox
MDM-2       GSTAT GMO-01              -           -        0.098      -       242       Sx
MDH-2       GSTAT GMO-03              -           -        0.533      -       124       Sx
MEL-1         CDN CGS-15            0.570       0.451        -        -       258       Sx
MEM-1         CDN CGS-11            0.730       0.683        -        -        4        Sx
MEM-2         OREAS 50Pb            0.841       0.744        -        -        37       Sx
MEH-2         OREAS 54PA            2.900       1.550        -        -        7        Sx
MEL-3          OREAS MM             0.248       0.444        -        -        4        Sx
MEL-4          CDN CM-5             0.294       0.319        -        -        4        Sx
MXL-1         OREAS 52Pb            0.307       0.334        -        -        37       Sx
MXM-1         OREAS 53Pb            0.623       0.546        -        -        19       Sx
MXL-2          OREAS 50c            0.836       0.742        -        -        7        Sx
MLB-1           CDN MM              0.016       0.156      0.130     0.19      41       Sx
MLL-1           CDN MM              0.030       0.156      0.081     1.48      52       Sx
MLM-1           CDN MM              0.030       0.145      0.886    13.98      24       Sx
MLH-1           CDN MM              0.047       0.137      1.610    23.10      17       Sx
MLE-1           CDN MM              0.038       0.098      8.580    122.50     8        Sx
 ALS            GMO-04                -           -          -       4.40     513       Sx
 ALS          NCSDC70018              -           -          -      31.20     430       Sx
  FB              IVA              <0.010      <0.010      <0.001   <0.01     1625    Granite
Notes:    OREAS MM = Mount Elliott matrix-matched series
          CDN MM = Merlin matrix-matched series
          ALS = Internal laboratory SRMs

As rhenium SRMs are uncommon, initial monitoring of rhenium performance was
totally reliant upon ALS’ use of commercially-purchased SRMs, which comprised two
types that were used in 943 determinations (refer to Table 11.2). Standard deviation
data were provided by the manufacturer for the higher value SRM, and the standard
deviation for the lower value SRM was calculated by ALS using actual data. Several
major checking programs, encompassing several labs, were completed to ensure
that rhenium performance was checked independently.

Tolerance limits for SRMs were set at two and three standard deviations (2SD and
3SD) from the Round Robin (RR) mean value of the SRM. A single batch failed if any
of the following criteria were met: a single SRM value was greater than the RR 3SD
limits; two consecutive SRM values were greater than the RR 2SD limits on the same
side of the mean; and a FB value was over 0.06 g/t Au, 0.06% Cu, 0.005% Mo, or
0.2 ppm Re. These limits were used because, if an analytical system were under
control, there would be only a 1% chance that an SRM would return assays




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exceeding 3SD, or a 0.625% chance that two SRMs in adjacent batches would
exceed 2SD on the same side of the mean. In all cases, a barren override can be
applied for batches deemed to have insignificant values.

Round robin programs also were conducted on barren material that was obtained
from fresh, unaltered, non-mineralised rock of the Mount Dore Granite about five
kilometres to the north. A total of 115 samples were tested by ALS laboratories at
Mt. Isa, Townsville, and Brisbane, and the material was accepted as a suitable blank
for monitoring contamination and sample mix-ups. A total of 1,625 field blanks were
used (refer to Table 11.2).

11.9      Laboratories

All routine samples collected for resource estimation were prepared and analysed at
independent sample preparation and analytical laboratories. No aspects of the
sample preparation and analytical work conducted at the laboratories were done by
an employee, officer, director, or associate of IVA.

Samples were shipped from the work site to the Australian Laboratory Services Pty.
Ltd. (ALS) facility in Mt. Isa (ALS-MI) where they were prepared, and some base
metal analytical work was completed. All gold analytical work was conducted by ALS
at Townsville facility (ALS-TV), and the remaining analytical work was done by ALS
at Brisbane facility (ALS-BR).

Check assay samples were sent to Activation Laboratories Ltd. (Actlabs) in Perth,
Western Australia, and Ancaster, Ontario, Canada; Genalysis in Perth, Western
Australia, and Townsville, Queensland; and Becquerel in Mississauga, Ontario,
Canada. Additional checking work, part of a metallurgical testing program, was
conducted by Burnie Research Laboratory in Burnie, Tasmania.

ALS operates in accordance with ISO/IEC 17025 under Nata accreditation No. 825;
Actlabs, in accordance with ISO/IEC 17025 under Nata accreditation No. 266 and
SCC accreditation No. 266; Genalysis, in accordance with ISO/IEC 17025 under
Nata accreditation No. 3244; and Becquerel, in accordance with ISO/IEC 17025
under SCC accreditation No. 422. Burnie has no accreditation.

11.10     Sample Preparation

Sample dispatches comprising 70 samples, including quality control samples, were
delivered to the ALS-MI laboratory for sample preparation. The ALS-MI facility has
one 23 m 3 drying oven; two 200 mm wide jaw crushers (Jacques and Essa); one riffle
splitter with 18 slots at 23 mm wide and 200 mm long; three Boyd crushers with
attached rotary splitters; and ten LM2 pulverisers. Occasionally, when ALS-MI has
become overloaded with samples, they were shipped to ALS-TV for preparation. The
ALS-TV facility is set up similar to the MI facility and has the following equipment: two
15 m 3 drying oven; two 200 mm wide Jacques jaw crushers; three coarse riffle
splitters; three Boyd crushers with attached rotary splitters; six LM2 pulverisers; and
twenty LM5 pulverisers.




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All crushing and pulverizing work was conducted at work benches set into a plenum
ventilation system, which creates a negative pressure to remove dust. Compressed
air was used for cleaning. When a backlog arose at ALS-MI due to shipping
difficulties during the rainy season, a small number of samples were sent to ALS-TV
for sample preparation.

The following protocol was used for preparing samples:
•       Sorting - All samples are sorted and checked.
•       Weighing - All samples are weighed.
•       Drying - Samples were dried at 100ºC as necessary.
•       Coarse-crushing - Samples were crushed using a 12 mm jaw setting, which
        produces a product of approximately 70% passing 6 mm. This step was
        omitted for samples comprising RC cuttings.
•       Splitting - Samples were ½ split through a 1 cm riffle splitter to obtain
        subsamples weighing approximately 3 to 4 kg. The remaining material was
        archived.
•       Fine-crushing - Splits from coarse-crushed samples were fine-crushed to 90%
        passing 2 mm. Screen tests were initially done at 5 mm and later were
        changed to 4 mm.
•       Splitting - Fine-crushed samples were split to obtain a 1 kg subsample and the
        remaining material was archived.
•       Pulverizing - Splits from fine-crushed samples were pulverised to 90% passing
        75 µm.
•       Subsampling - Three subsamples, weighing about 100 g each, were obtained
        by taking multiple scoops from the pulverised samples and placing into Kraft
        bags. The remaining material was archived.
•       Cleaning - After processing each sample, all equipment was flushed with
        barren material, if a request for mineralised samples was made by a logging
        geologist on the SSMs. If barren flushes were used, crushers were blown out
        only after the flush had gone through, whereas pulverisers were blown out
        before and after the flush. If flushes were not used, all equipment was blown
        out after each sample.
•       Coding - Each sample was assigned an internal ALS bar code, and stickers
        with a code were printed for all sample bags.

Screen tests were conducted on crushed and pulverised material taken from one
sample out of each batch of samples. Results were reviewed as part of this report
compilation, and only a few anomalous crushing results related to schistose material
were noted.

11.11      Analyses

All samples were analysed by ALS for molybdenum, copper, rhenium, and gold using
assay methods and a multi-element geochemical suite. The key difference between




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the two methods is precision: the assay method uses larger acid volumes and a large
volumetric finish, whereas the geochemical method uses small volumes. Measuring
larger volumes leads to a better precision than measuring smaller volumes. All over-
range geochemical samples were re-analysed using assay methods. Primary
analytical methods comprise Inductively Coupled Plasma - Optical Emission
Spectrometry (ICP-OES), Inductively Coupled Plasma - Mass Spectrometry (ICP-
MS), and Atomic Absorption Spectrometry (AAS). Analytical work for some samples
was done at ALS-MI for copper and molybdenum assays and for multi-element
geochemical suites; at ALS-TV, for gold assays; and at ALS-BR, for some copper
and molybdenum assays and multi-element geochemical suites. ALS-MI has two
Varian Vista Pro ICP-AES instruments; ALS-TV, two Varian Spectra 220 AAS
instruments and one Varian Spectra 240 AAS instrument; and ALS-BR, nine ICP-
AES and four ICP-MS instruments.

Copper and molybdenum were initially determined by an assay method, using a two-
acid digestion (ALS code OG46). Following internal ALS checking work that was
conducted at IVA’s request to improve assay precision, the procedure was changed
to an assay method using a four-acid digestion (ALS code OG62). As part of the
switch in procedures, all molybdenum assays over 0.05% that had been completed
by OG46 were redone using OG62. Analyses for the multi-element geochemical
suite were done by two-acid (ALS code ME-ICP41) or four-acid digestions (ALS code
ME-ICP61). The detection limit for copper was initially 0.01% for the OG46 method,
but it was later changed to 0.001%; the latter detection limit was carried through for
the OG62 method. Molybdenum detection limits were 0.001%.

Rhenium was initially determined by a geochemical method, using a two-acid
digestion (ALS code ME-MS42). This method was later changed to a geochemical
method that used a four-acid digestion (ALS code ME-MS61). Since IVA requested
that assay methods be used for rhenium to improve precision, the method was
changed to a four-acid digestion under ALS code OG62, and rhenium assays were
done by this new method. When molybdenum assays over 0.05% that had been
completed by OG46 were redone by OG62, rhenium was included. The detection
limit for rhenium was 0.001 ppm for all methods.

The following procedures were used for copper, molybdenum, and rhenium
analyses:
•     OG46 - copper and molybdenum assays, using two-acid digestion:
      -      0.4 g sample.
      -      31.25 ml of HCL and 5.75 ml nitric acids.
      -      Digested at 115°C for 10 minutes and then at 150°C for 60 minutes.
      -      Diluted to 100 ml with distilled H2O.
      -      Analysed by ICP-OES / ICP-MS.
•     ME-ICP41 / ME- MS42 - multi-element geochemical analyses, using two-
      acid digestion:
      -      0.5 g sample.




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        -     3 ml of HCL and 1 ml of Nitric acids.
        -     Digested at 115°C for 45 minutes.
        -     Diluted to 12.5 ml with distilled water.
        -     Analysed by ICP-OES / ICP-MS.
•       OG62 - copper, molybdenum, and rhenium assays, using four-acid
        digestion:
        -     0.5g sample.
        -     2 ml of Nitric, 1 ml of Perchloric, and 2 ml of Hydrofluoric acids.
        -     Digested at 185°C for 2½ hours.
        -     Leached with 30 ml of 50% HCL to dissolve soluble salts.
        -     Diluted to 100 ml with distilled water.
        -     Analysed by ICP-OES / ICP-MS.
•       ME-ICP61 - multi-element geochemical analyses, using four-acid
        digestion:
        -     0.5 g sample.
        -     2 ml of Nitric, 1 ml of Perchloric, and 2 ml of Hydrofluoric acids.
        -     Digested at 185°C for 2½ hours.
        -     Leached with 5 ml of 50% HCL to dissolve soluble salts.
        -     Diluted to 12.5 ml with distilled water.
        -     Analysed by ICP-OES / ICP-MS.

Gold was determined by a 30 g fire assay fusion, cupelled to obtain a prill, digested
with Aqua Regia, and finished by AAS with a detection limit of 0.01 g/t.

All analytical results were posted by ALS to Webtrieve, a client accessible website. In
addition, results were reported digitally by email to individual IVA personnel as
requested in CSV, SO2, and PDF formats.

11.12       Monitoring

Digital assay results were imported into an acQuire database, and the laboratory’s
values for SRMs and FBs were compared to the established SRM pass-fail criteria.
Sample batches that passed were given a Priority 1 status, whereas batches that
failed were given a Priority 3 status, and the laboratory was asked to re-assay until
they passed, at which point they were re-assigned a Priority 1 status. Monitoring
charts that recorded the performance of individual SRMs with respect to RR
tolerance limits were maintained to monitor performance.




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11.12.1 Standard Reference Materials

All SRMs used by IVA for monitoring copper, molybdenum, and rhenium
performance had round robin programs completed that provided statistical tolerance
limits for passing and failing data. Monitoring charts with final SRM assays for more
than 20 determinations are presented in Figure 11.3 to Figure 11.18. Overall
performance has been good with only a low number of failures. Copper had an initial
slightly-high bias that diminished with time; molybdenum was initially biased slightly
low with failures due to digestion issues that were rectified with a procedural change;
and rhenium initially had poor precision, but it improved with time. Only a few failures
remain from batches deemed to be barren.

Figure 11.3      Quality Control Monitoring Charts for Cu Assays, MDL-1




Figure 11.4      Quality Control Monitoring Charts for Cu Assays, MDM-1




Figure 11.5      Quality Control Monitoring Charts for Cu Assays, MDH-1




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Figure 11.6      Quality Control Monitoring Charts for Mo Assays, MDM-2




Figure 11.7      Quality Control Monitoring Charts for Mo Assays, MDH-2




Figure 11.8      Quality Control Monitoring Charts for Cu Assays, MEL-1




Figure 11.9      Quality Control Monitoring Charts for Cu Assays, MEM-2




Figure 11.10 Quality Control Monitoring Charts for Cu Assays, MXL-1




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Figure 11.11 Quality Control Monitoring Charts for Mo Assays, MLB-1




Figure 11.12 Quality Control Monitoring Charts for Rhenium Assays, MLB-1




Figure 11.13 Quality Control Monitoring Charts for Mo Assays, MLL-1




Figure 11.14 Quality Control Monitoring Charts for Re Assays, MLL-1




Figure 11.15 Quality Control Monitoring Charts for Mo Assays, MLM-1




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Figure 11.16 Quality Control Monitoring Charts for Re Assays, MLM-1




    Figure 11.17 Quality Control                       Figure 11.18 Quality Control
  Monitoring Charts for Re Assays in                 Monitoring Charts for Re Assays in
            NCSDC70018                                           GMO-04




11.12.2 Field Blanks

Monitoring charts for copper, molybdenum, and rhenium analyses, using
Field Blanks (FB), are presented in Figure 11.19 to Figure 11.21. Results show a low
incidence of contamination and mix-ups for gold and copper; however, some carry-
over issues were noted with molybdenum and rhenium, following extremely high
grade samples. This issue was investigated in detail, and the cause was found to be
mineralised rock fragments that were caught up inside a ledge on a coarse jaw
crusher. Rhenium shows a similar pattern to molybdenum as the two elements are
strongly associated. The issue was rectified with the laboratory, and all affected
dispatches were re-assayed. Several samples with elevated molybdenum and
rhenium values remain, because coarse reject material was unavailable for re-
assaying. The significance of the remaining values was deemed to be negligible
compared to the grades of surrounding samples.




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Figure 11.19 Quality Control Monitoring Charts for Cu Assays in Field Blanks




Figure 11.20 Quality Control Monitoring Charts for Mo Assays in Field Blanks




Figure 11.21 Quality Control Monitoring Chart for Re Analyses in Field Blanks




11.13     Duplicates

Three types of duplicate samples were collected: field, coarse, and pulp. Field
duplicates were obtained from drill core or RC cuttings; crush duplicates, from
crushed samples; and pulp duplicates, from pulverised samples. In addition, lab
replicate data were provided by ALS. The type and number of duplicate samples
classified by analytical methods is tabulated in Table 11.3. These data were
compiled for copper, molybdenum, rhenium, and gold, and statistically analysed by
scatter (SC), quantile-quantile (QQ), relative difference (RD), Thompson-Howarth
(TH), and percentile rank (PR) plots. Results are presented only for IVA duplicates
from samples used mostly for resource modelling: duplicates from core samples
assayed for molybdenum, copper, and rhenium, using method OG62, and for copper,
using method OG46.




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Table 11.3            Type and Number of Duplicate Samples by Analytical Methods
                            Au AA25          Cu OG62         Cu OG46          Mo OG62          Mo OG46     Re OG62
           Field            426              331             153              346              44          339
IVA-       Crush            855              649             196              687              70          675
HC         Pulp             857              661             176              689              70          678
           Total            2138             1641            525              1722             184         1692
           Field            102              8               110              8                12          8
IVA    -   Crush            141              17              124              17               9           17
RC         Pulp             140              16              121              16               8           16
           Total            383              41              355              41               29          41
ALS        LabRep           1260             1409            649              1460             160         1419
Notes: HC = core drilling w ith half core sampling; RC = reverse circulation drilling.


SC plots compare values on X and Y axes, whereas QQ plots rank values first and
then compare them. RD plots compare absolute or relative differences between pair
values and the mean grade of a pair. These statistical plots are used to search for
biases, which would indicate sampling issues. SC, QQ, and RD plots with descriptive
statistics for field, crush, and pulp duplicates taken from core samples and assayed
for copper, molybdenum, and rhenium are shown in Figure 11.22 to Figure 11.33.
Other than a few obvious sample mix ups, all plots show symmetrical patterns about
zero, indicating unbiased sampling.

TH plots show grade on the X axis and precision on the Y axis. Practical detection
limits occur where the precision is 100%; however, bearing in mind that the precision
here is really the difference between pair values, 100% precision really means 100%
variability. Precision levels for TH plots should ideally be better than or close to 15%
for coarse duplicates and 5% for pulp duplicates-this is because these can be
controlled by the sub-sampling protocol. Core duplicates precision cannot be
controlled by the sub-sampling protocol; however, it should ideally be better than or
close to 30%, but it is not unusual for it to be several tens of percent higher. TH plots
for field, crush, and pulp duplicates taken from core samples and assayed for copper,
molybdenum, and rhenium are shown in Figure 11.34 to Figure 11.37. The levels of
precision are presented in Table 11.4 and are within or close to ideal ranges.

Table 11.4            Levels of Asymptotic Precision for OG46 Duplicates
                                             (from Thompson How arth Plots)
                                                                                  Core Samples
                            Element
                                                         N         Core           N       Crush       N          Pulp
                  Molybdenum (0.01%)                  330       39%          649         24%         649       4%
 OG62             Rhenium (0.1ppm)                    330       49%          550         31%         594       2%
                  Copper (0.01%)                      330       30%          649         10%         650       2%
 OG46             Copper (0.01%)                      154       35%          176         22%         176       5%
       Notes: N is based on groups of 11 with lowest values trimmed; asymptotic precision is at higher grades


PR plots rank data above a threshold by absolute relative difference. Thresholds for
PR plots are chosen at values where the precision begins to decrease dramatically
as the detection limits are approached. These plots are used to compare precision at




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various stages in the sampling and preparation process to ensure that it improves
with decreasing particle size and that it is within reasonable limits for the style of
mineralization. Precision levels for PR plots at the 90th percentile should ideally be
better than 15% for coarse duplicates and 10% for pulp duplicates-this is because
these can be controlled by the sub-sampling protocol. Although the core duplicates’
precision cannot be controlled by the sub-sampling protocol, it would ideally be better
than or close to 30% at the 90th percentile; however, it is not unusual for it to be
several tens of percent higher. PR plots for field, crush, and pulp duplicates taken
from core samples and assayed for copper, molybdenum, and rhenium are shown in
Figure 11.22 to Figure 11.33.




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Figure 11.22 Cu by OG62 for Core Duplicates




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Figure 11.23 Mo by OG62 for Core Duplicates




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Figure 11.24 Re by OG62 for Core Duplicates




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Figure 11.25 Cu by OG46 for Core Duplicates




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Figure 11.26 Cu by OG62 for Crush Duplicates




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Figure 11.27 Mo by OG62 for Crush Duplicates




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Figure 11.28 Re by OG62 for Crush Duplicates




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Figure 11.29 Cu by OG46 for Crush Duplicates




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Figure 11.30 Cu by OG62 for Pulp Duplicates




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Figure 11.31 Mo by OG62 for Pulp Duplicates




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Figure 11.32 Re by OG62 for Pulp Duplicates




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Figure 11.33 Cu by OG46 for Pulp Duplicates




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   Figure 11.34 TH plot for Cu by OG62                             Figure 11.36 TH plot for Re by OG62




   Figure 11.35 TH plot for Mo by OG62                           Figure 11.37 TH plot for Cu by OG46




 Figure 11.38 to Figure 11.41. Levels of precision for these plots are presented in
 Table 11.5. The levels of precision are within or close to ideal ranges for copper and
 molybdenum, but not rhenium. The lower level of precision for rhenium is related to
 lower value data and also may be a result of using ICP-MS.

 Table 11.5           Levels of Precision at 90th Percentile for Duplicates
                                                (from Percentile Rank Plots)
                                                                         Core Samples
                     Element
                                                N           Core            N         Crush             N    Pulp
             Molybdenum (0.01%)           42             67%            104          18%          96        9%
 OG62        Rhenium (0.1ppm)             132            40%            358          91%          268       67%
             Copper (0.01%)               268            44%            553          17%          548       7%
 OG46        Copper (0.01%)               142            63%            179          29%          151       11%
Notes: All data trimmed below cut-off, which is the visual limit of poor precision above detection limit




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Figure 11.38 PR plot for Cu by OG62                 Figure 11.40 PR plot for Re by OG62




Figure 11.39 PR plot for Mo by OG62                 Figure 11.41 PR plot for Cu by OG46




11.14     Checking Programs

11.14.1 Programs

Nine external assay checking programs were completed for Mount Dore assays,
including:
•       MD1 - Digestion method checks for copper and molybdenum.
•       MD2 - Metallurgical checks for molybdenum.
•       MD3 - Discovery checks of rhenium.
•       MD4 - Analytical method checks for molybdenum and rhenium.
•       MD5 - Routine checks for copper, molybdenum, rhenium, gold, and zinc.
•       MD6 - Routine checks for copper, molybdenum, rhenium, gold, and zinc.
•       MD7 - Routine checks for copper, molybdenum, rhenium, gold, and zinc.
•       MD8 - Routine checks for copper, molybdenum, rhenium, gold, and zinc.
•       MD9 - High grade ratio checks for rhenium and molybdenum.

In addition to the assay checking programs, a twinning program for comparing RC
and diamond drilling samples was conducted to assess variations between the two
methods of data collection.




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11.14.2 Set MD1 - Digestion Checks

ALS conducted internal checking work to compare copper and molybdenum assays
that were obtained by the OG46 (two-acid) and the OG62 methods (four-acid).
Results indicated little difference between the two methods for copper assays
(refer to Figure 11.42), whereas molybdenum assays showed that OG46 assays can
be up to 20% lower than the OG62 assays (refer to Figure 11.43). Differences are
related to higher temperatures used for the OG62 method. The results of this work
were used to support changes to the analytical protocol and re-assaying of all
samples that were analysed by OG46 that initially returned over 0.05% molybdenum.

Figure 11.42 Comparison of ALS Cu                  Figure 11.43 Comparison of ALS Mo
by OG46 Versus OG62 Methods, MD1                   by OG46 Versus OG62 Methods, MD1




11.14.3 Set MD2 - Metallurgical Checks

As part of a metallurgical testing program, a total of 16 half-core samples were
submitted to Burnie Research during February 2009, as part of a metallurgical testing
program. Analytical test work included X-Ray Fluorescence (XRF) analyses for
molybdenum, which were compared to ALS OG62 assays (refer to Figure 11.44).
Results indicate that the ALS OG62 assays compare quite favourably with the Burnie
XRF assays.

11.14.4 Set MD3 - Discovery Checks

A total of 349 pulp samples were submitted to Actlabs during November 2008, for
initial checking of rhenium analyses by neutron activation. The results of the rhenium
validation by neutron activation are shown in Figure 11.45. A linear regression
analysis using Actlabs analyses as the independent umpire analyses indicates that
ALS analyses agree well.




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Figure 11.44 Metallurgical Check                   Figure 11.45 Actlabs Re Discovery
Assays for Mo, MD2                                 Check Analyses, MD3




11.14.5 Set MD4 - Method Validation Checks

Following the discovery of the Merlin Zone, a validation program was conducted to
compare rhenium and molybdenum analyses using different analytical methods. The
validation work was conducted because IVA was unable to introduce molybdenum
and rhenium reference materials into the sample stream immediately following the
Merlin discovery, consequently a gap existed in QAQC monitoring until commercial
molybdenum standards were purchased and matrix-matched rhenium and
molybdenum standards were developed. A total of 536 samples, which was originally
analysed at ALS, were sent to Genalysis, Actlabs, and Becquerel laboratories for
molybdenum and rhenium analyses. The wet digestion and dry preparation methods
that laboratories used are as follows:
•     ALS:
      -     Molybdenum assays initially done by two acid but later switched to
            four acid.
      -     Molybdenum geochemical analyses (four acid).
      -     Rhenium assays (four acid).
      -     Rhenium geochemical analyses (four acid).
•     Genalysis:
      -     Molybdenum assays (four acid).
      -     Rhenium assays (four acid).
•     Actlabs:
      -     Molybdenum assays (four acid).
      -     Rhenium assays (four acid).
      -     Rhenium analyses (neutron activation).
•     Becquerel:
      -     Rhenium analyses (neutron activation).
Several statistical comparisons were completed. The primary objective was to
validate the original rhenium analyses that used a four acid digestion wet assaying
method with neutron activation analyses using pressed powder pellets. A secondary




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objective was to validate rhenium and molybdenum assays with four acid digestion
methods using different laboratories.

The results of the rhenium validation by neutron activation are shown in
Figure 11.46. A linear regression analysis using Becquerel and Actlabs analyses as
the independent umpire analyses indicates that ALS analyses agree well with
Actlabs, but appear to be biased high compared to Becquerel.

The results of the rhenium validation by four acid digestion methods are shown in
Figure 11.47. A linear regression analysis using Genalysis and Actlabs analyses as
the independent umpire analyses indicates that ALS analyses appear to be slightly
high.

The results of the molybdenum validation by four acid digestion methods are shown
in Figure 11.48. A linear regression analysis using Genalysis and Actlabs analyses
as the independent umpire analyses indicates that ALS analyses agree well.

Figure 11.46 Re Neutron Activation                 Figure 11.48 Mo Four Acid Digestion
Method Validation Analyses, MD4                    Method Validation Analyses, MD4




Figure 11.47 Re Four Acid Digestion
Method Validation Analyses, MD4




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11.14.6 Set MD5 - Routine Checks

IVA operated an ongoing routine checking program that comprised four pulp
duplicate samples per dispatch of 70 submitted samples, which were taken from
fixed positions at the laboratory and returned to the work site. A total of 1,729
samples from the Mt. Elliott, Mount Dore, and Recon projects were sent to Genalysis
for checking of assays previously completed at ALS. Of these samples,
approximately 30% were from the Mount Dore project. Samples were compiled
together, divided into groups of 18, and two QC samples consisting of a SRM and a
FB were inserted randomly into each group to create batches of 20 samples. SRMs
were certified for gold and copper, but not for molybdenum and rhenium because this
was one of the first routine check assay programs conducted that included
Mount Dore samples. Copper and gold assays were conducted on all samples, and
samples from Mount Dore were check assayed for molybdenum, rhenium, and zinc.
When assays were received from the laboratory, IVA’s standard quality control
monitoring criteria were applied, which has been described in a previous section.

The results of the copper check assaying by Genalysis are shown in Figure 11.49. A
linear regression analysis using Genalysis as the independent umpire analyses
indicates that ALS copper analyses are overall biased low. However, the bias is
grade dependent and appears to be related to Genalysis assays as they are actually
biased high above 8% and biased low below 8%, which is noted on the QC
monitoring charts.

The results of the molybdenum check assaying by Genalysis are shown in
Figure 11.50. A linear regression analysis using Genalysis as the independent
umpire analyses indicates that ALS molybdenum analyses are overall biased high.
However, this bias appears to be related to Genalysis assays as the observation is
affected by a small number of higher grade data and is not supported by the Set
MD2 metallurgical checks and Set MD4 method validation checks, which have more
data at higher grades.

Figure 11.49 Cu Check Assays, MD5                   Figure 11.50 Mo Check Assays, MD5




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11.14.7 Set MD6 - Routine Checks

A second set of routine check samples from the Mount Dore project was compiled
using the same procedures as described under Set MD5. A total of 451 samples
from the Mount Dore, Mt. Elliott, and Recon projects were sent to Genalysis for
checking of assays that were previously completed at ALS. Approximately 75% of
the samples were from the Mount Dore Project. Copper, molybdenum, rhenium, gold,
and zinc check assays were conducted on all samples.

The results of the copper check assaying by Genalysis are shown in Figure 11.51. A
linear regression analysis using Genalysis as the independent umpire analyses
indicates that ALS copper analyses agree well.

The results of the molybdenum check assaying by Genalysis are shown in
Figure 11.52. A linear regression analysis using Genalysis as the independent
umpire analyses indicates that ALS molybdenum analyses appear to be biased high.
However, this bias appears to be related to Genalysis assays as the observation is
affected by a small number of higher grade data and is not supported by the Set
MD2 metallurgical checks and Set MD4 method validation checks, which have more
data at higher grades.

The results of the rhenium check assaying by Genalysis are shown in Figure 11.53.
A linear regression analysis using Genalysis as the independent umpire analyses
indicates that ALS rhenium analyses agree well.

11.14.8 Set MD7 - Routine Checks

A third set of routine check samples from the Mount Dore project was compiled using
the same procedures as described under Set MD5. A total of 290 samples from the
Mount Dore project were sent to Genalysis for checking of assays that were
previously completed at ALS. Copper, molybdenum, rhenium, gold, and zinc check
assays were conducted on all samples.

The results of the copper check assaying by Genalysis are shown in Figure 11.54. A
linear regression analysis using Genalysis as the independent umpire analyses
indicates that ALS copper analyses agree well.

The results of the molybdenum check assaying by Genalysis are shown in
Figure 11.55. A linear regression analysis using Genalysis as the independent
umpire analyses indicates that ALS molybdenum analyses are overall biased low.
Considering that the IVA check assay sample submission did not have any SRMs
above 2000 ppm, which is where biased data are, then the bias could be related to
Genalysis data.

The results of the rhenium check assaying by Genalysis are shown in Figure 11.56.
A linear regression analysis using Genalysis as the independent umpire analyses
indicates that ALS rhenium analyses agree well.




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Figure 11.51 Cu Check Assays, MD6                   Figure 11.54 Cu Check Assays, MD7




Figure 11.52 Mo Check Assays, MD6                   Figure 11.55 Mo Check Assays, MD7




Figure 11.53 Re Check Assays, MD6                   Figure 11.56 Re Check Assays, MD7




11.14.9 Set MD8 - Routine Checks

A fourth set of routine check samples from the Mount Dore project was compiled
using the same procedures as described under Set MD5. A total of 314 samples
from the Mount Dore project were sent to Genalysis for checking of assays that were
previously completed at ALS. Copper, molybdenum, rhenium, gold, and zinc check
assays were conducted on all samples.




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The results of the copper check assaying by Genalysis are shown in Figure 11.57. A
linear regression analysis using Genalysis as the independent umpire analyses
indicates that ALS copper analyses agree well.

The results of the molybdenum check assaying by Genalysis are shown in
Figure 11.58. A linear regression analysis using Genalysis as the independent
umpire analyses indicates that ALS molybdenum analyses agree well.

The results of the rhenium check assaying by Genalysis are shown in Figure 11.59.
A linear regression analysis using Genalysis as the independent umpire analyses
indicates that ALS rhenium analyses agree well.

11.14.10     Set MD9 - High Grade Ratio Checks

During submission of the routine check assays during drilling of the Merlin Zone, it
became apparent that higher grade molybdenum and rhenium assays were not being
checked adequately. In addition, some routine assays showed anomalously high or
low Mo to Re ratios, which was disconcerting considering the strong association of
molybdenum and rhenium. In order to address these concerns, a separate high
grade ratio checking program was initiated. The program comprised 285 samples in
groups of six to twenty-four, taken from 20 drill holes. The selection was based on
strings of higher grade molybdenum assays, particularly where anomalous ratios
were observed. Samples were compiled together, divided into groups of 18, and two
QC samples, consisting of an SRM and a Field Blank were inserted randomly into
each group to create batches of 20 samples. All samples were assayed for
molybdenum, rhenium, copper, gold, and zinc at Genalysis. IVA’s standard quality
control monitoring criteria, described in a previous section, were applied to all
assays. This particular set of check assays utilised matrix-matched SRMs and
included an external standard with extremely high molybdenum and rhenium grades.

The results of the molybdenum check assaying by Genalysis are shown in
Figure 11.60. A linear regression analysis using Genalysis assays as the
independent umpire values shows that ALS molybdenum analyses are overall biased
low. However, this observation is affected by a high bias with Genalysis that was
noted with matrix-matched SRMs in the 1% to 8% range, which returned values up to
10% above the round robin mean. There are no matrix-matched SRMs above 8% Mo
and, although an external SRM at 56% Mo does not show this bias, there are no
check samples at that grade.

The results of the rhenium check assaying by Genalysis are shown in Figure 11.61.
A linear regression analysis using Genalysis assays as the independent umpire
values shows that ALS rhenium assays agree well.

The results of the copper check assaying by Genalysis are shown in Figure 11.62.
Two sets of check assays were provided: one by ICP-OES, and the other by AAS. A
linear regression analysis using Genalysis assays as the independent umpire values
shows that ALS copper assays agree well.




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Figure 11.57 Cu Check Assays, MD8                    Figure 11.60 Mo High Grade Ratio
                                                            Check Assays, MD8




                                                     Figure 11.61 Re High Grade Ratio
Figure 11.58 Mo Check Assays, MD8                           Check Assays, MD8




                                                     Figure 11.62 Cu High Grade Ratio
Figure 11.59 Re Check Assays, MD8                           Check Assays, MD8




11.14.11     Twinning

Analytical data from RC hole MDQ0153 were compared to those from diamond drill
hole MDQ0153A to assess variations between the two methods of data collection.
The two holes were collared 3.3 m apart at the same elevation and drilled at minus




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90°. The diamond drill hole is down dip from the RC hole, and lithological contacts
are calculated to be about 4 m lower in depth; however, grade breaks are at similar
depths.

Three correlatable copper zones are present: a thicker upper zone, 36 m wide; and
two narrower lower zones, 10 and 18 m wide (refer to Table 11.6 and Figure 11.63).
A thick zone of molybdenum, 30 m wide, underlies the lowermost copper zone (refer
to Table 11.6 and Figure 11.64). Four sulphur zones correspond to the three copper
zones and one molybdenum zone, and most zone boundaries are at the same
depths, except for the upper sulphur zone, which has the lower boundary further up
hole (refer to Table 11.6 and Figure 11.65).

                        Figure 11.63 RC Cu Versus                                                                                                                                                   Figure 11.65 RC Sulphur Versus
                            Diamond Drilling Cu                                                                                                                                                        Diamond Drilling Sulphur
                                                                                              Ivanhoe Australia Limited                                                                                                                                      Ivanhoe Australia Limited
                                                                                          Merlin Zone - Copper Assays                                                                                                                                  Merlin Zone - Molybdenum Assays
                                                                Twinning of MDQ0153(RC) vs MDQ0153A(DD)                                                                                                                                     Twinning of MDQ0153(RC) vs MDQ0153A(DD)
                                                                                                     Copper RC             Copper DD                                                                                                                          Sulphur RC             Sulphur DD
             4.0                                                                                                                                                                                            7.0


             3.5                                      Zone 1                                                                      Zone 2                             Zone 3                                                      Zone 1                                                   Zone 2                   Zone 3             Zone 4
                                                                                                                                                                                                            6.0


             3.0
                                                                                         Zone of strong                                                                                                     5.0
                                                                                          oxide copper
             2.5
                                                                                                                                                                                                            4.0
  Cu (%)




                                                                                                                                                                                                    S (%)




                                                                                            Probable
             2.0                                                                         zone of strong
                                                                                          oxide copper                                                                                                      3.0
             1.5

                                                                                                                                                                                                            2.0
             1.0


                                                                                                                                                                                                            1.0
             0.5


             0.0                                                                                                                                                                                            0.0
                                                                                                                                                                                                                  18

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                                                                                                                                                                                                                                                                    90

                                                                                                                                                                                                                                                                         98

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                                                                                                                                                                                                                                                                                                 130

                                                                                                                                                                                                                                                                                                       138

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                                                                                                                                                                                                                                                                                                                          162

                                                                                                                                                                                                                                                                                                                                170

                                                                                                                                                                                                                                                                                                                                       178

                                                                                                                                                                                                                                                                                                                                             186

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                   18

                             26

                                       34

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                                                                                                                                                 138

                                                                                                                                                        146

                                                                                                                                                               154

                                                                                                                                                                      162

                                                                                                                                                                            170

                                                                                                                                                                                  178

                                                                                                                                                                                        186

                                                                                                                                                                                              194




                                                                                                     Down Hole Distance (m)                                                                                                                                         Down Hole Distance (m)




                        Figure 11.64 RC Mo Versus
                           Diamond Drilling Mo
                                                                                              Ivanhoe Australia Limited
                                                                                     Merlin Zone - Molybdenum Assays
                                                                Twinning of MDQ0153(RC) vs MDQ0153A(DD)
                                                                                         Molybdenum RC                           Molybdenum DD
             70000

                                                                                                                                                                            Zone 4
             60000



             50000
  Mo (ppm)




             40000



             30000



             20000



             10000



                   0
                        18

                                  26

                                            34

                                                      42

                                                                50

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                                                                                                                                                                            170

                                                                                                                                                                                  178

                                                                                                                                                                                        186

                                                                                                                                                                                              194




                                                                                                      Down Hole Distance (m)




RC grade profiles are smoother than those from diamond drilling, which is expected
because of homogenization of chips from larger samples in RC and the actual
heterogeneity of mineralization from smaller diamond drill core samples. There
appears to be no down-hole contamination below higher grade zones. Overall
intercept values from RC are typically lower than those from diamond drilling, except
for the upper copper zone. RC copper intercepts are 15% to 30% lower than
diamond drilling; molybdenum, 196% lower; and sulphur, 5% to 66% lower. The large
differences are probably related to loss of fine copper and molybdenum
mineralization during RC drilling and sampling.




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Table 11.6       Visual Intercept Comparisons for RC vs. Diamond Drilling
                (MDQ0153 - MDQ0153A) (copper = tan, molybdenum = grey, sulphur = yellow).




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The upper copper zone appears to be an exception to the grade comparison pattern
as RC copper is 42% higher than diamond drilling; however, a closer look at the
down-hole grade profiles for this zone indicates that several RC copper spikes
correlate to diamond drilling copper troughs, which has skewed the intercept
comparison. When these spikes/troughs are removed, the RC copper intercept drops
to 15% below the diamond drilling intercept, similar to the other copper zones.

The RC copper spikes occur in two groups, comprising three samples each. The
lower group correlates to a thin siltstone - phyllite unit that is characterised by non-
sulphide copper minerals, based on core logging-this is supported by lower sulphur
analyses. As the upper group does not have a separately-logged interval, no direct
comments can be made about lithological associations. However, it is interesting to
note that the upper group has a sulphur spike in the diamond drilling analyses,
suggesting the presence of secondary sulphates such as jarosite or gypsum rather
than pyrite, which is less likely to be present in strongly oxidised rocks.

The effect of extreme values on intercept comparisons between the two drilling
methods was investigated further by removing from the calculation pairs of data with
large differences. After approximately 25% of values were removed, RC intercept
values were still typically lower than those from diamond drilling, although not as
extreme. RC copper intercepts were 20% to 305% lower than diamond drilling;
molybdenum, 95% lower; and sulphur, 0% to 66% lower.

Although modern RC drilling systems typically use face-centred bits to eliminate
contamination by channelling cuttings back up the centre of the drill stem, recoveries
are still low enough that they could affect the final assay results. The twinning
comparison of RC and diamond drilling indicates that significant low biases in RC
analytical results mean that care should be exercised if RC is to be used as a tool for
obtaining analytical data for resource modelling.

11.14.12     Security and Chain of Custody

All bulk bags for shipping samples were sealed with individually numbered tamper-
proof security tags and transported by IVA vehicle to the ALS Laboratory in Mt. Isa.
SSMs, corresponding to the shipment dispatches, were sent electronically to the
laboratory; shipments were examined upon arrival at the laboratory; and the SSMs
were returned back to IVA, marked with a confirmation of the state of the security
seals on boxes, the presence of all samples comprising each batch, and laboratory
report numbers assigned to each batch. Following completion of assaying, samples
were stored at the laboratory and then transported back to the work site by IVA
vehicle for long-term storage.

11.15      Adequacy of Sample Preparation, Analytical, and Security Procedures

The sample preparation, analytical, and security procedures were adequate to
ensure high quality drill-hole assay data acceptable for geological modelling and
reliable resource estimation. There is an unbroken chain of custody from the site to
the analytical laboratory; sufficient reference materials have been used to control
analytical processes; appropriate analytical procedures were used that take rock




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matrices into account and provide acceptable levels of precision; and sufficient
checking work has been done to demonstrate that the data are unbiased.

Data were imported into the database via automated processes that have essentially
eliminated transcription issues. Record duplication issues do arise when querying the
database to obtain data sets for statistical analyses and plotting-these issues are
covered by manual checking of exports in spreadsheets through comparison of data
and conditional statements. Since the database has some built in quality control
plotting routines; a mixture of manual and only limited automated routines was used
for compiling and plotting. The manual portion of the work is required to maintain an
ongoing working data verification process. There are no limitations to the verification
process.




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12    DAT A VERIFICATION

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and is here reproduced in its entirety for the
convenience of the reader.

IVA has applied modern drilling, sampling and surveying methods to derive the data
used as the basis for the current resource estimate. Golder has reviewed the data
and considers it suitable for resource estimation purposes.

Previous independent verification of the mineral resources is available as:
•     IVA drilling and resource estimates have achieved similar copper grades and
      volumes to those defined by previous workers at the Mount Dore South
      copper resource; see Table 2, Golder 2010, for a listing of previous results.
•     QG (2010) reviewed eight holes and confirmed that the range of grades
      observed in the assay log correlated well with the observed abundance of
      MoS. In particular high grade >1% Mo correlated with Mo breccia in the core.
•     QG (2010) undertook some database validation for the previous
      NI43-101 report, including an audit of 5% of the Merlin drilling, which
      highlighted some minor issues that were resolved at the time.

Other reviews include the following:
•     The majority shareholder IVN (Ivanhoe Mines) regularly reviews the
      operations of IVA for procedures, laboratory and QAQC, see Sketchley
      (2008abc, 2009, 2010abc). These reviews are completed every 6 months and
      provide a high level of technical assistance to the site geologists as well as a
      regular review of processes and procedures. IVA has demonstrated a resolve
      to address the concerns previously raised and to improve its practises.
•     IVA undertakes extensive QAQC work amounting to 23% of all assay results.
      This proportion is above typical industry practise and demonstrates a
      commitment to quality.
•     IVA demonstrates a commitment to sample security in its sample dispatch
      processes.
•     IVA reviewed all surface and down hole survey work, IVA (2010). Following
      previous reviews IVA committed to resurveying all collars and many down hole
      surveys.
•     Where possible collar surveys and tenement locations are located by an
      independent surveyor.
•     Poorly defined or unusual down hole surveys have been resurveyed using
      more thorough gyroscopic methods to remove survey uncertainty in most
      cases. This remains an area of constant review and reassessment due to the
      length of some of the deeper drilling.
•     In 2009 the drilling data were migrated to acQuire database. This provides on
      entry data validation and improves the data integrity and capture processes.
      The process of the migration included a review of the database and inclusion




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      of meta-data required. IVA has been progressing through the issues raised
      during the migration addressing some issues.

Golder reviewed the procedures and processes available while on site which
included drilling, sampling, logging, standard down hole survey, database and
interpretation processes. Due to timing of the site visit Golder did not observe density
measurement or other periodic process such as survey and gyroscopic down hole
survey processes, but understand the processes used and believe they are suitable.

Golder completed the following verification and audit processes:
•     Golder (2010) reviewed a complete drill section and observed significant
      sections of massive molybdenite mineralisation where high grade Mo assays
      were present. These observations confirm the general tenor of the Mo
      mineralisation indicated.
•     ALS provided 642 batches of assays with over 44,000 assay records covering
      Mount Dore drilling from November 2007 to August 2010. The assay data were
      provided directly by ALS without possible interference from IVA. These were
      compiled into assays for each sample and method for 32,140 samples relevant
      to the Mount Dore resource estimate (excluding samples for QAQC and
      surrounding exploration). Preferred assays for Au, Ag, Cu, Co, Mo, Re, Zn, Pb,
      S and Fe were compiled from the different assay methods using Golders’
      independent assumption of the preferred precedence of analysis methods. The
      data were then compared directly to the assay data provided by IVA and used
      for the resource estimate. The results indicated a very low proportion of
      discrepancies which highlighted only a few areas for further validation and
      correction. Most differences related to the preferred assay method for low grade
      samples which is not material to the resource estimate. Significant
      discrepancies equating to the cut-off grade for resource reporting included 1
      sample for Re, 3 samples for Mo, 2 samples for Au, 1 sample for Zn, 0 samples
      for Cu and do differences at all for Ag, Co, Pb, Fe, S.
•     12 pages of hardcopy assay certificates (~360 assay) from pre IVA drilling were
      validated against the drill database. This represents 3% of the pre IVA drilling
      data and is limited to a sample of the available certificates. No issues were
      determined except for rounding of Cu analyses to 3 decimal places which is not
      material to the resource estimate. IVA is currently indexing several thousand
      documents from the now closed Starra mine. This may provide access to more
      historic work and allow a more systematic audit of the historic data.
•     20 pages of hardcopy assay certificates (~800 assay) from early 2004 IVA
      drilling were validated against the drill database. This data was assayed by
      SGS a different laboratory to that subsequently used by IVA for all analyses.
      This represents one quarter of the early IVA drilling data. Comparison of a
      regular 10% sample of the data indicated correct data entry for all Au and
      sequential Cu analyses. Very similar values were recorded for most total Cu
      analyses, indicating reanalysis of Cu at ALS for these samples. IVA undertook
      this reanalysis due to issues with the Cu analyses at SGS and the desire to
      analyse for additional elements including Mo, Re, Co, Ag, Zn, Pb, S and Fe.




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      This reanalysis was confirmed and demonstrated by IVA explaining all the
      differences noted.
•     Lodewyk Surveyors provided point survey data directly to Golder without
      possible interference from IVA. The data were compiled and compared directly
      to the collar data provided by IVA and used for the resource estimate. Of the
      322 collar records that could be matched. Golder noted 18 discrepancies which
      were all attributed to either multiple surveys or 1 mm differences attributable to
      rounding error. Two remaining drill hole errors of 2 m are related the mix-up of
      the collars on one drill site resolved later during resurveying of the down hole
      orientation. This comparison of the collars verifies the location of the majority of
      drilling data and confirms the collar locations as inspected by an independent
      licence surveyor.
•     Golder independently surveyed some collars during the site visit with a hand
      help Garmin GPS60csx. The comparison of the collar surveys to the GPS
      coordinates and track log confirmed the location of the prospect and layout of
      the project drilling, concentrating on the northern area relevant to Merlin.
•     Golder compiled 1089 results from the detailed gyroscopic down hole surveys
      from 21 drill holes. These were completed by IVA in 2009 and by Surtron
      Technologies Pty Ltd in 2010. Some filtered values were confirmed as
      erroneous data correctly removed.
•     129 pages of hardcopy density measurement worksheets were reviewed that
      cover measurements between 3/2009 and 9/2009. 1 in 10 were selected and
      checked against the data used in the resource estimate. The records supported
      the density values calculated. 5 data entry errors from 309 indicate a low data
      entry error rate with all errors discovered being minor and insignificant to the
      use of the data. Some significant blocks of data were identified as requiring data
      entry. The incomplete entry of data is an ongoing area IVA needs to address on
      all of their projects. This should not affect the resource evaluation as there is a
      sizable database of existing data which indicates low variability in the density
      values at Mount Dore. Golder calculated all density values independently and
      did not use any density values that did not have supporting measured weights.

Additional comments include:
•     A full review of the older historical drill data is not yet complete. Some assay
      values were verified. IVA is currently indexing a large quantity of historic reports
      from the closed Starra mine where additional supporting documents for the
      older drilling should be located for future use. Note that the older drilling does
      not contribute to the Merlin Mo-Re resource and only significant contributes to
      the Mount Dore south Cu resource.
•     Though there are no independent density samples available, the values
      supplied by IVA are typical for those expected for the type rocks which have a
      low observed porosity and high silica content. The values indicate little
      variability in dry bulk density and are similar to surrounding properties that
      Golder has worked on. A variety of density measurement methods have been
      used resulting in comparable results.




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Golder considers the drilling data base, its collation using specialised third party
database software and the processes meets industry standard practises. The
resource drilling data is suitably supported and maintained for resource estimation
purposes.

Ongoing work should include:
•     Detailed review of the acQuire database structure (now in progress).
•     Physical database audit of both the oldest drilling (pre 2007) as further records
      are discovered and indexed.
•     Selected independent bulk density samples.




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13     MINERAL PROCESSING AND METALLURGICAL TESTING

13.1      Metallurgical Testwork

Preliminary metallurgical testwork was carried out by Burnie Research Laboratories
(AMMTEC 2009) to determine crushing, grinding and flotation characteristics, and
with sighter tests for ferric leaching, roasting and calcine leaching to determine metal
deportment though the process. This work was referenced in the SRK NI 43-101
report (SRK 2010). No further work on crushing and grinding was undertaken for the
PFS.

Metallurgical testwork on fresh samples were carried out by Metcon Research
(Metcon, 2011, three reports) to refine the flotation plant flowsheet. The testwork
results were interpreted by EHP Consulting, Inc. and incorporated into the process
design criteria used for the PFS (EHP Consulting, 2009).

Samples of Molybdenum Rougher Flotation Tails were prepared by KD Engineering
and Metcon and shipped to Pocock Industrial Inc (Pocock) in slurry form for solids
liquid separation (SLS) testing in late December 2010. Pocock is located in Salt Lake
City, Utah.

The SLS testwork was reported in Pocock (2011). EHP and Jacobs reviewed
available information for concentrate treatment plants and their recommendations
were incorporated into the design criteria for the concentrate treatment plant.

13.2      Summary of Testwork

The samples were selected to ensure coverage over the depth and strike length of
the deposit. In addition all lithologies and mineralisation styles were covered. In all
approximately 1.1 tonnes of material was sent to Metcon.

Metcon testing was subdivided into three phases:
•      Phase I focused on optimising conventional flotation conditions and running
       locked cycle testing on a composite representative of the entire ore body. The
       primary goal was to evaluate the potential of eliminating the ferric chloride
       leach and post roast leach from the flowsheet using conventional molybdenum
       flotation techniques.
       The test work indicated the ferric chloride leach could be eliminated, but the
       molybdenum concentrate grade could not be upgraded above 35-40%
       molybdenum using conventional techniques. A post roast leach would be
       required. The primary contaminants were gangue and carbon. All indications
       were that the gangue reporting to the concentrate was associated with the
       carbon.
•      Phase II expanded the test-work to non-conventional techniques for
       concentrate grade improvement; the evaluation of higher grade copper
       composites to test the robustness of copper depression; and the potential of
       producing a saleable grade by-product copper concentrate.




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      No technique was found to improve the molybdenum concentrate grade above
      the 35 to 40% molybdenum range. The testing showed that a three stage
      copper depression circuit eliminates the need for a ferric chloride leach even
      with high copper content ores. A flowsheet incorporating a regrind can be used
      to produce a saleable grade by-product copper concentrate. The copper
      concentrate contains 3 to 5 g/t gold and 70 to 90 g/t silver.
•     Phase III looked at 18 variability samples and an initial evaluation of roasting
      requirements.
      In nine of the variability samples, saleable grade molybdenum concentrate was
      produced. The culprit responsible for concentrate downgrading was identified
      as graphite. The distribution of molybdenum to throwaway tail streams
      averaged 5 to 7%. The distribution of molybdenum to recycled tail streams
      averaged 10 to 15%. Accounting for recovery from the recycle streams, Metcon
      anticipates overall molybdenum recoveries of 90%.




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14     MINERAL RESOURCE ESTIMATES

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and is here reproduced in its entirety for the
convenience of the reader.

14.1      Resource Domains

Geological domains are used to distinguish and divide areas of different
mineralisation style or statistical nature. Domains attempt to create zones of
statistical stationarity where structure and controls are constant. When achieved the
domains are then suitable for statistical and geostatistical analysis and for estimation.

14.2      Grade Domains

Initial analysis by QG (2008) for the copper mineralisation at Mount Dore South
resulted in the simple approach of domaining based on geological units, oxidation
profile and economic copper cut off. Domaining became more complicated with the
discovery of the Merlin Mo-Re resource and the evaluation of the Mount Dore North
mineralisation.

Mount Dore North is more complicated due to overlapping Cu-Mo-Zn zones,
including:
•      Mo-Re mineralisation at Merlin generally lies just below the carbonaceous
       metasediments at the contact with silicified calc-silicate rocks, locally identified
       by the limit of red rock (haematite +/- K feldspar) alteration;
•      Relatively little copper mineralisation has been intersected in the upper
       sequence of the carbonaceous metasediments whereas at depth copper
       becomes more prevalent and overlaps with the Mo mineralisation while also
       extending into the hanging wall of the Mo mineralisation;
•      Zn mineralisation becomes more prevalent towards the north. This is
       particularly the case for the upper copper mineralisation (above Merlin) at
       Mount Dore North where the Zn is apparently leached from most of the oxide
       zone but occurs as a broader domain than Cu, both overlapping with and
       linking the two main upper Cu mineralisation lenses;
•      Zn mineralisation also occurs in the more distal hanging wall extensions to the
       lower Cu mineralisation. Again overlapping with the Cu (and Mo) lenses and
       extending both above and further into the hanging wall.

Finally, Mount Dore north and south areas contrast markedly in oxidation depth. The
southern area is deeply weathered with the fresh zone poorly defined or tested.
Occasional Zn in the fresh zone of the south indicates that Zn domains may persist
to some extent. The northern areas has a shallower oxide and transition zones as
well as deeper drilling that targeted the depth extent of the Merlin mineralisation,
allowing the Mo and Zn zones to be well defined.

Towards the north a depletion zone for Cu is evident near surface whereas Cu
extents to surface in the southern and south-western most areas.




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Previous work by QG (2009) at Mount Dore North identified domains based on
geology, oxidation and mineralisation cut-offs including:
•     >0.5% Mo for high grade Mo at Merlin.
•     >0.01% Mo for anomalous Mo around and mostly below Merlin.
•     >0.25% Cu for potentially economic copper mineralisation.
•     >0.25% Zn for anomalous Zn mineralisation.

This required the use of multiple partially overlapping mineralisation domains to flag,
assess and estimate different elements. This represented the first pass in
understanding and domaining the Mount Dore resource. Golder undertook geological
review and statistical analysis that confirmed the statistical basis for the original
domains.

Golder reviewed the domaining process by first reassessing the Principal Component
Analysis (PCA) used to assist QG in its original domaining. The original work by QG
(2009) identified four principal element groupings, including:
•     Mo-Re-S.
•     Cu-Ag-Au-Co.
•     Pb-Zn.
•     Fe.

Golder undertook a global PCA analysis for all samples from the Kuridala formation,
including both mineralisation and background samples. This was used to assist
domaining and determine grade relationships.

Review of cross sections indicates that there is some correlation between some
elements in some areas which are not evident elsewhere. For example the high Mo
grade Merlin domain does contain Cu in places and can display some correlation of
high Cu grades with higher Mo grade that would suggest a relationship which is not
present when there is no Mo mineralisation (away from Merlin). This conclusion is
further supported by the analysis of correlation statistics which indicate weak
correlation between some elements across the element groups defined by QG for
some domain subsets.

There is sufficient evidence both from the PCA and geological review that separate
estimation of the four main element groups is not warranted as it runs the risk of
removing local correlations between elements from different groups i.e. it decouples
the estimation of elements that may still have local correlation. The most significant
risk is in the narrow high grade Mo domain where previous estimation of other
elements and density using broader shapes could significantly dilute and smooth the
other elements compared to Mo and Re.

Golder adapted the domaining process to produce a set of interpretations that were
largely concentric and able to define a domain on precedence. The changes to the
interpretation included:




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•      Addition of the carbonaceous meta-sediment footwall. This boundary appears
       to be the structural location for much of the high grade Mo mineralisation. Its
       interpretation allows better control and understating of the high grade Mo
       zones.
•      The Cu cut-off was retained at 0.25% Cu to define the core Cu zone but
       relaxed to include contiguous grade down to 0.1% Cu. This relaxation
       improved the interpretability of the Cu domains after infill drilling made some
       zones difficult to follow at a rigorous 0.25% Cu cut-off.
•      The previous Zn domain cut-off did not consider any economic factors or the
       occurrence of other potentially economic elements such as Pb, Au, Ag and
       Co. The domain was found to largely include the Cu mineralisation below the
       oxide zone with correlation between Zn and Cu evident in the Cu domain
       though Zn extended beyond the Cu. The interpretation of the Zn domain was
       revised using a Cu equivalent (ECu) grade using the same approach and cut-
       off as used for Cu. This allowed the interpretation of a concentric zone
       surrounding the Cu domain and including interstitial or additional areas mostly
       on the basis of the additional value potential from Zn, though considering all
       other elements.
•      Retention of the Mo domains using 0.01% Mo and 0.5% Mo cut-offs.

14.3      Mo Domaining (MODOM)

Domaining based on Mo grade is based on two cut-offs at 0.01% Mo and ~0.5% Mo
to define background, medium and high grade domains.

The high grade domain is defined by a significant grade contrast and the narrow
continuous structural location and is referred to as the Merlin resource. It comprises
narrow zones of massive molybdenite carrying high grade Mo and Re grades either
as a single narrow zone or a closely spaced cluster of narrower intercepts.

The Merlin mineralisation consists of a distinct molybdenite rich narrow breccia zone
with grades typically >10% Mo over short intervals. The regular 2 m core sampling
that was adopted for Cu mineralisation and used throughout the Merlin drilling
campaign up until 2010, has resulted in non selective sampling and produced a
range of Mo grades from the typical rich intercepts to more moderate grades in 2 m
samples. This has contributed to considerable edge dilution and the smoothing of the
grade distribution. Despite this dilution the Merlin resource is sufficiently high grade
that a cut-off 0.5% Mo offers excellent discrimination of the high grade Mo zone into
four defined narrow structural domains that display sufficient continuity to warrant
separate consideration.

The definition of the Merlin structural domains effectively accounts for all high Mo
and Re grades from the remaining material, minimising any chance of smearing or
overestimating Mo and Re in lower grade areas. The narrow structural domains offer
a potential target as for narrow selective mining options.

Re displays a close relationship with Mo and its occurrence and maximum
concentration is intimately related to molybdenite occurrence and hence Mo grade.




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The segregation of the very high grade Mo and Re samples is well justified to
improve the volume - grade estimates. A very good correlation exists indicating an
intimate association of the two metals, see Figure 14.1.The “dirty” molybdenite is the
Re host however the intimate relationship with “clean” molybdenite does not offer a
potential for physical separation of the two phases during milling (refer to Figure
14.2).

Figure 14.1      Plot of Mo vs. Re Assay Values from Drill Cores (Kirby 2009)




Figure 14.2      Typical Laser Ablation - ICPMS Trace of “Dirty and Clean”
                 Molybdenite (Kirby 2009)




The Merlin deposit extends north of 7,605,050 mN. A small high grade extension
known as Little Wizard at around 7,605,000 mN is the southern and upper most
extension of the Merlin structure. It represents a small higher grade domain that is
treated separately due to its potential economic influence on the project and the
bonanza grades present. The main Merlin domain defined displays a distinct spatial
relationship with the base of the logged carbonaceous meta-sediment, within the
Kuridala Formation. At depth this zone appears to split into several structural
domains, some follow the base of the carbonaceous meta - sediment and others
extend in to the hanging wall mineralisation, where additional Cu and Zn
mineralisation becomes more prevalent.




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The medium Mo grade domain forms a broad zone, largely including and
surrounding the high grade Mo domain, though it is situated predominantly below
and only extending a few meters above the high grade Mo domain into the hanging
wall. It comprises disseminations or stylolitic veinlets and disseminations that are
small and lower grade than the Merlin domains. To define a domain for the medium
Mo grade significant interburden is included to enable all anomalous Mo and Re
grades to be included in the domain and limit any potential smoothing of grades
beyond the limited extent on Mo mineralisation. There is a smaller zone of
anomalous Mo mineralisation in the hanging wall Cu mineralisation package (just
below the granite contact) that is identified in a few drill holes. Both the grade and
occurrence of this mineralisation is not significant and it has not been modelled
separately.

The statistical basis for domaining the high grade Mo zone that is the Merlin deposit
is well founded at a cut-off of around 0.5% Mo on the basis of the distribution of
grades, see Figure 14.3. Note that the histogram bin size changes to allow all the
data to be plotted as the high grade tail is very long.

Figure 14.3                       Mo & Re Grade Distribution from all Samples
                            10%                                                              10%
                                                     KF Waste                                                KF Waste
                            9%                                                               9%

                            8%                       Other Domains                           8%              Other Domains

                            7%                       Mo Med Grade                            7%              Mo Med Grade
    Percentage of Samples




                                                                     Percentage of Samples




                            6%                       Mo High Grade                           6%              Mo High Grade

                            5%                                                               5%

                            4%                                                               4%

                            3%                                                               3%

                            2%                                                               2%

                            1%                                                               1%

                            0%                                                               0%
                                   >12
                                    10




                                                                                                    10
                                                                                                    11
                                                                                                    12
                                                                                                    13
                                                                                                    14
                                                                                                    16
                                                                                                    18
                                                                                                    20
                                                                                                    22
                                                                                                    24
                                                                                                    26
                                                                                                    28
                                                                                                    30
                                                                                                    32
                                                                                                    34
                                                                                                    36
                                                                                                    38
                                                                                                    40
                                                                                                    42
                                                                                                    44
                                  0.00
                                  0.05
                                  0.10
                                  0.15
                                  0.20
                                  0.25
                                  0.30
                                  0.35
                                  0.40
                                  0.45
                                  0.50
                                  0.55
                                  0.60
                                  0.65
                                  0.70
                                  0.75
                                  0.80
                                  0.85
                                  0.90
                                  0.95




                                     4
                                     6
                                     8




                                                                                                     0

                                                                                                     1

                                                                                                     2

                                                                                                     3

                                                                                                     4

                                                                                                     5
                                                                                                     6
                                                                                                     7
                                                                                                     8
                                                                                                     9
                                   1.0
                                   1.2
                                   1.4
                                   1.6
                                   1.8
                                   2.0
                                   2.2
                                   2.4
                                   2.6
                                   2.8




                                                                                                   0.5

                                                                                                   1.5

                                                                                                   2.5

                                                                                                   3.5

                                                                                                   4.5




                                          Mo %                                                      Re g/t


   Source: Golder 2010

Mo is generally situated below the oxide weathering surface and generally displays
little effective change across the weathering boundaries. Molybdenite appears to
remain unoxidised in near surface high grade areas such is Little Wizard and there is
no expected loss of recovery due to oxidation that would warrant metallurgical
domaining.

Figure 14.4 displays the log-probability plots for all samples based on the Mo
domaining and basic geology. In this figure the three Merlin high grade Mo domains
are combined (MO_345) but separated from the higher grade Little Wizard domain
(MO_6) and the medium grade domains (MO_12). There is good separation of the
distributions indicating effective domaining. Furthermore separation of the
background mineralisation based on geology is also warranted.

Note the change in slope of the log probability plots between the high and medium
grade Mo domains is a further indication of the change in grade distribution between
the two styles of Mo occurrence.




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Figure 14.4                           Mo & Re Log-Probability Plots by Mo Domains (MODOM)




14.4      Cu Domaining (CUDOM)

Domaining for copper remains the same as previously establish by QG (2009 and
2010) and is based on a 0.25% Cu cut-off. This cut-off is supported by the grade
distribution displayed in Figure 14.5 where there is an inflection in the grade profile at
around 0.2% Cu. At this cut-off interpretable domains can be defined that have
significant continuity. At Mount Dore South this includes a single wide shallow
dipping zone within predominantly oxidised or partial oxidised material. At
Mount Dore North defined narrow zones are evident with zones of lower grade or
interburden. In total 6 Cu mineralisation structures can be defined with three in the
upper mineralisation package and three in the lower mineralisation package. The two
packages are separated by a wide zone of largely barren material. Though
occasionally the Zn mineralisation will close this gap the copper remains fairly distinct
as two defined packages.

Figure 14.5                           Cu Grade Distribution from Samples
                                      40%
                                                                               KF Waste
                                      35%                                      Mo Domain
                                                                               Zn Domain
                                      30%                                      Cu Domain
              Percentage of Samples




                                      25%

                                      20%

                                      15%

                                      10%

                                      5%

                                      0%
                                            >3.5
                                             0.0
                                             0.1
                                             0.2
                                             0.3
                                             0.4
                                             0.5
                                             0.6
                                             0.7
                                             0.8
                                             0.9
                                             1.0
                                             1.1
                                             1.2
                                             1.3
                                             1.4
                                             1.5
                                             1.6
                                             1.7
                                             1.8
                                             1.9
                                             2.0
                                             2.1
                                             2.2
                                             2.3
                                             2.4
                                             2.5
                                             2.6
                                             2.7
                                             2.8
                                             2.9
                                             3.0
                                             3.1
                                             3.2
                                             3.3
                                             3.4




                                                                Cu %




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For the upper mineralisation package the upper two Cu domains occur consistently
throughout the Mount Dore North area, whereas the low third domain only occurs on
a few sections and has some continuity issues that are yet to be resolved
adequately.

The uppermost Cu domain in the lower package is, at times the same as the Mo
domain but elsewhere the Cu mineralisation appears to separate and follow its own
path. This indicates separate mineralisation event that in places follow the same
structural feeder. The lowermost two Cu zones are relatively small and are not
interpreted or domained separately as they are adequately defined by the Mo
mineralisation shapes.

The 0.25% Cu cut-off was relaxed to allow the inclusion of contiguous mineralised
samples down to a lower cut-off of 0.15% Cu. This improved the interpretability of the
domain structures which becomes more difficult as infill drilling progresses. This
process also includes some edge dilution softening the hard domain boundary
approach used for estimation.

14.5      Polymetallic Domaining (P_DOM)

Previous estimates have domained Zn separate to Cu and Mo and there are good
indications that Zn mineralisation is different in many respects, including:
•      Zn and Pb are partially leached in the oxide and transition zones resulting in
       lower grades near surface.
•      Zn and Pb are significant at Mount Dore North but the mineralisation appears
       to become progressively weaker south of 7,605,000 mN and into Mount Dore
       South.
•      In the upper mineralisation package at Mount Dore North Zn occurs with the
       Cu mineralisation but also in and around the Cu domains linking them together
       into a broad zone of mineralisation and helping to define the upper package as
       one zone.
•      In the lower mineralisation package at Mount Dore North Zn occurs mainly
       down dip in and around the Cu which also appears to get stronger down dip.
       Zn rarely occurs in any significant amount towards surface and the dip trend is
       stronger than Cu and appears in places to persist beyond the Cu
       mineralisation, particularly into the hanging wall.

Initial Zn was domained on the same basis as previously undertaken by QG (2009),
based on a 0.25% Zn cut-off, see Figure 14.6. This approach is justified but becomes
very complicated when considering the overlap between the other domains for Cu,
high grade Mo and low grade Mo shapes.




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Figure 14.6      Zn Grade Distribution from Samples
                                                     30%
                                                                                        KF Waste
                                                                                        Mo Domain
                                                     25%
                                                                                        Cu Domain
                                                                                        Zn+Cu Domain




                             Percentage of Samples
                                                     20%                                Zn Only Domain


                                                     15%


                                                     10%


                                                     5%


                                                     0%




                                                           >3.5
                                                            0.0
                                                            0.1
                                                            0.2
                                                            0.3
                                                            0.4
                                                            0.5
                                                            0.6
                                                            0.7
                                                            0.8
                                                            0.9
                                                            1.0
                                                            1.1
                                                            1.2
                                                            1.3
                                                            1.4
                                                            1.5
                                                            1.6
                                                            1.7
                                                            1.8
                                                            1.9
                                                            2.0
                                                            2.1
                                                            2.2
                                                            2.3
                                                            2.4
                                                            2.5
                                                            2.6
                                                            2.7
                                                            2.8
                                                            2.9
                                                            3.0
                                                            3.1
                                                            3.2
                                                            3.3
                                                            3.4
                                                                           Zn %



This complicated approach of overlapping domains which resulted in inevitable
inconsistencies in interpreted structure and over domaining of the resource with
inherent difficulties in validating the estimation results. To resolve the complexity the
low grade Mo domain and Zn domains were replaced with a single polymetallic
domain to capture outer areas that have potentially economic value from other
elements. This used a Cu equivalent grade calculation to create an outer shell that
encompassed all the areas of potential value, in most cases these are dominated by
the occurrence of Zn in the Mount Dore North upper and Mount Dore North lower
(down dip) and Mo for Mount Dore North lower (middle to up dip).

Copper equivalence was calculated using the available recovery and price
assumptions from an early draft of Merlin scoping study (SRK, 2010) and is
summarised in Table 14.1. Note that final public version of the SRK study had
revised prices that are not reflected in Table 14.1 or the geological interpretations.
This is not considered significant as the prices used reflect the general comparative
value and are suitable for geological domaining. The conversion factors are rounded
to one or two significant places to provide general factors commensurate to the level
of study. These factors may require revision as the metallurgical results, processing
options and prices are refined, but are suitable for low grade geological domaining as
used here.

Table 14.1       Cu Equivalence Conversion Factors@
                  Element           Metal Price                  Units    Recovery   Rounded CuEq Factor
                    Mo                                     15    US$/lb    87.1%                   5.1
                     Cu                                    3     US$/lb    84.6%                    1
                     Re                               241.13     US$/oz    75.7%                   0.1
                     Zn                                    0.7   US$/lb    80.0%*                  0.22
                     Pb                                          US$/lb    80.0%*              0.22*
                     Ag                                    12    US$/oz    80.0%*              0.006
                     Au                                    700   US$/oz    90.0%*                  0.36
                 * Assumption
                 @ Note prices and recoveries were based on values available at an early stage of geological
                 interpretation and may differ from current project assumptions

14.6      Geology Domains (ROCK)

A review of the geological logging by IVA highlighted some inconsistency due to the
number of geologists involved and changes in the logging procedures over time. A




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relogging program was undertaken in mid 2010 to resolve the inconsistencies and
bring all drill holes up to the same standard of coding.

Geological domaining was undertaken by IVA site geologist and is based on 100 m
spaced sections logged to date. The interpretations are not snapped to drill holes
since not all drilling has been relogged and the geological boundaries are not critical
to the mineralisation interpretations.

Structural modelling is ongoing but the current interpretation includes some faulting
to account for the larger offsets noted during the relogging.

14.7      Weathering Domains (MINL)

Previously logging of oxidation has used a mixture sulphide mineral occurrence and
material characteristics and has changed slightly over time as the Mount Dore project
has progressed and numerous geologists have undertaken logging.

In 2010 IVA undertook a relogging program to provide a more standardised basis for
the definition of the oxidation profile based on physical characteristics, without the
partial integration of assumed leaching characteristics. The logging has been
completed on 100 m sections and focused on the definition of two surfaces to define
the three principal weathering types, defining the MINL domains. These surfaces
include:
•      Base of complete oxidation (BOCO). This is the contact between completely &
       partially oxidised material types based on degree of degradation of the core,
       conversion of feldspars to clay and internal limonitic dissemination throughout
       the core.
•      Base of partial oxidation (BOPO). This is the contact between partially
       oxidised and fresh material types and is based on the amount of oxidation on
       joint and fracture planes, presence of clays on fractures and presence or
       absence of calcite or vughs throughout the core.

The oxidation surface models are relatively approximate as they are only compiled
on 100 m sections, there is some intermixing of materials locally, oxidation logging
includes some subjective decisions and there is sufficient faulting and other variation
to complicate the simple surface model.

14.8      Metallurgical Domains (METDOM)

Golder previously reviewed the copper sequential analyses (Golder, 2010) where a
ternary plot of the three assayed copper components defines four distinct zones of
different mineralogy. Figure 14.7 reproduces the ternary diagram of the three
sequential copper assay components Cu_SOL (Cu in acid), Cu_CN (Cu in Cyanide)
and Cu_Res (Cu in residual material). The four populations defined include:
•      Insoluble minerals (bottom right red), representing relatively insoluble copper
       as possibly Cu-phosphate, Mn wads, native copper or tied up in clays
       (possibly kaolinite & illite).




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•     Highly acid soluble minerals (top red), representing copper oxide, carbonates
      and silicate.
•     Highly cyanide soluble minerals (bottom left blue), representing copper
      supergene minerals such as chalcocite.
•     Insoluble minerals (bottom right blue), representing primary copper minerals
      such as chalcopyrite and bornite.

These grouping were classified and cross sectional interpretations completed to
indentify the four groups as sequential zones. Figure 14.7 displays the samples
colour coded for the four domains interpreted and provides a reasonable definition of
the zone given some intermixing is inevitable.

Figure 14.7      Sequential Cu Ternary Diagram Colour Coded on Existing Initial
                 Reinterpretation Based on Sequential Copper
                Ternary Diagram for Sequential Copper      Cu_SOL
                by Sequential Copper Domaining
                                                                    Acid Soluble Cu Oxides:
                                                                    Mainly Chrysocolla + Cuprite, Chalcotrichite



                                                                                              Fresh
                                                                                              transitional
                                                                                              lower oxide
                                                                                              upper oxide
                                                                                              Axis
                   Supergene Cu Sulphides:                     4

                   Chalcocite, Covellite

                                                                                          Insoluble Remnant Cu:
                                                                                          Native Cu




                            Cu_CN               Primary Cu Sulphides: Chalcopyrite, Bornite          Cu_RES




During the relogging completed by IVA in 2010, on 100 m spaced cross sections, the
base of predominantly leachable copper was also interpreted. This was primarily
interpreted from the dominance of the sulphide copper minerals and the review of
sequential copper assays where present. It is noted that this interpretation lies below
the lowermost surface defined from the ternary diagram, presumably because a
lower effective cut-off for the residual copper has been applied. Since the IVA
interpretation includes an assessment of the logged mineralogy that persists beyond
the available copper sequential analyses this surface has been used to assist the
lateral extension of the copper sequential domains.

Together these surfaces define five domains available; to assist metallurgical
assessment of the copper leachability. There are insufficient copper sequential
analyses across the entire resource to allow effective estimation of the leachable
copper components. Table 14.2 presents the average copper sequential analyses for
each of the metallurgical domains for all available samples and also only those inside
the copper resource domains in Table 14.3.

The copper sequential assays report slightly lower total results than the original
sample; but are within expected error margins for the assay method. The result
indicate reasonable zonation is achieved with:




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•          METDOM 1 - upper residual zone have a mix of some residual and
           otherwise mostly oxide (acid soluble) copper minerals.
•          METDOM2 - oxide copper zone dominated by acid soluble oxide copper
           minerals.
•          METDOM 3 - supergene zone dominated by cyanide soluble minerals
           such as chalcocite.
•          METDOM 4 - transitional zone marked by the increase in residual copper
           minerals such as primary sulphides.
•          METDOM 5 - fresh zone as logged in core with only slightly higher
           copper sulphide content.

Table 14.2           Mean Sequential Copper Analyses by the Metallurgical Domains
                     (METDOM), all Samples
    MET-     Sampl     Cu %     Cu %    Cu %      Cu %     Cu %     Diff-                   %
                                                                            % Sol   % CN
    DOM        es      Orig      Sol     CN        Res     Total   erence                  Res
     1        419      0.35     0.19     0.03     0.12      0.34    0.01    55%     9%     35%
     2        1435     0.82     0.58     0.16     0.06      0.81    0.01    73%     20%    7%
     3        851      0.60     0.21     0.31     0.07      0.59    0.01    36%     52%    12%
     4        541      0.43     0.11     0.20     0.11      0.42    0.01    27%     47%    26%
     5        521      0.41     0.10     0.18     0.12      0.40    0.01    25%     46%    29%


Table 14.3           Mean Sequential Copper Analyses by the Metallurgical Domains
                     (METDOM), >0.25% Cu
    MET-     Sampl      Cu %    Cu %    Cu %      Cu %     Cu %     Diff-                   %
                                                                            % Sol   % CN
    DOM        es       Orig.    Sol     CN       Res      Total   erence                  Res
     1        181       0.64    0.38     0.06      0.18     0.62    0.02    62%     10%    29%
     2        1059      1.06    0.76     0.22      0.07     1.04    0.02    73%     21%    6%
     3        525       0.88    0.31     0.45      0.10     0.86    0.01    36%     52%    12%
     4        262       0.77    0.21     0.36      0.18     0.75    0.02    28%     47%    24%
     5        295       0.63    0.16     0.28      0.17     0.61    0.02    26%     46%    29%

14.9         Domain Definitions

The definition of the codes and associated wireframe models for each individual
domain is defined in Table 14.4.

14.10        Combined Domains (DOM)

Although each domain for Cu, Mo and polymetallic mineralisation are defined and
flagged separately these are combined for estimation and analysis purposes using
an order of precedence that includes:
•          MODOM: Mo domains take precedence followed by.
•          CUDOM: Cu domains then.
•          P_DOM: polymetallic domains.




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Since the polymetallic domains are defined by a copper equivalent cut-off they
provide outer shells that encapsulate much of the copper domains. The wireframe
ignored areas adequately modelled by the copper domains and targeted only
significant areas where additional value is provided by other elements, most notably
Mo or Zn.

The precedence of the Mo domains over Cu domains does result in the copper
domains being split up where the Cu and Mo domain cross or change from being
coincident. There are areas where Mo and Cu display significant correlation. This is
not always consistent and suggests that if the Cu and Mo mineralisation are not
related to the same mineralising event then they may be related in part due to the
use of the same structures or due to scavenging or remobilisation of the original
mineralisation during the formation of the secondary mineralisation. Since Mo
provides the most dominant economic target for underground mining and tends to be
the most spatially restricted of the two it is used to subdivide the Cu domain into that
which occurs with Mo and that which does not.




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Table 14.4       Individual Domain Codes
Dom ain Field/                                                                      Wireframe     Wireframe
                                    Dom ain Description
   Value                                                                              Name        Projection
   MODOM                                                   Mo Mineralisation
      4                                 Little Wizard                                1006m4         inside
      5                               Merlin - main vein                             1006m5         inside
      6                                Merlin - vein 6                               1006m6         inside
      7                                Merlin - vein 7                               1006m7         inside
   CUDOM                                                   Cu Mineralisation
      1                     Upper sequence - vein 1 (north only)                     1006c1         inside
      2                     Upper sequence - vein 2 (north only)                     1006c2         inside
      3                     Upper sequence - vein 3 (north only)                     1006c3         inside
      5                           Low er Sequence - vein 5                           1006c5         inside
      6                     Low er Sequence - vein 6 (south only )                   1006c6         inside
      7                     Low er Sequence - vein 7 (south only )                   1006c7         inside
   P_DOM                                             Poly-metallic Mineralisation
      1                               Upper sequence                                 1006p1         inside
      5                         Low er sequence main zone                            1006p5         inside
      7                        Low er sequence second zone                           1006p7         inside
    MINL                                                 Geology - Oxidation
      2                         Completely oxidised material                     g1006BOCO      Z-up (ie base)
      3                          Partially oxidised material                     g1006BOPO      Z-up (ie base)
      4                                Fresh material                                               default
    LITH                                                 Geology - Lithology
      1                                    Granite                                  g1006gra    X-up (footwall)
      2                             Kuridala Fm. Phyllite                           g1004phy    X-up (footwall)
      3                     Kuridala Fm. Black Shale & Siltstone                     g1006kf    X-up (footwall)
      4                               Quartzite (SQT)                               g1006sqt    X-up (footwall)
      5                              Stavely Formation                                              default
  METDOM                                                 Metallurgical Dom ain
      1              Copper sequential zone - upper residual copper              met1006cus1    Z-up (ie base)
      2            Copper sequential zone - oxide / acid soluble copper          met1006cus2    Z-up (ie base)
      3             Copper sequential zone - cyanide soluble copper              met1006cus3    Z-up (ie base)
      4           Copper sequential zone - start of lower residual copper            Default    Z-up (ie base)
      5                        Logged primary sulphide zone                      met1006sul     Z-down (ie top)


The combined domains for waste and ore materials (excluding weathering) are
defined in Table 14.5. The domain codes roughly define a top down sequence and
split the upper and lower mineralisation sequence to allow separate statistical
analysis. The change in style of mineralisation from South to North Mount Dore at
around 7,605,000 mN is not well defined by drilling in that area making it difficult to
determine how the upper and lower sequence of mineralisation that is evident in
Mount Dore North join to the single sequence defined in Mount Dore South. The
current interpretation infers that the lower mineralisation sequence continues and the
upper sequence disappears.




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Table 14.5       Combined Domain Codes
Dom ain    Description                           Condition
1          Waste – Granite                       No mineralisation and ROCK = 1
2          Waste – Phyllite                      No mineralisation and ROCK = 2
3          Upper Ore sequence: Cu                CUDOM = 1,2,3
4          Upper Ore sequence: Polymetallic      CUDOM = 10 and P_DOM=1
5          Waste – Kuridala                      No mineralisation and ROCK = 3
6          Low er Ore sequence: Mo (Merlin)      MODOM = 4,5,6,7
7          Low er Ore sequence: Cu               MODOM = 0 and CUDOM = 5,6,7
8          Low er Ore sequence: Polymetallic     CUDOM = 0 and MO_DOM = 0 and P_DOM=5,7
9          Waste - Quartzite and Suavely         No mineralisation and ROCK = 4 or 5

14.11     Domain Boundary

Domain boundaries may be defined as ‘soft’ or ‘hard’. Previous statistical
assessment of the domain boundary behaviour by QG (2009, 2010) indicated abrupt
grade changes across the domain boundaries which indicate hard boundaries were
required.

This behaviour is still evident visually for Mo at the Merlin high grade domains and is
also evident for Cu in the Cu domains and is the basis of the existing interpretation.
The subdivision of the copper domains by the Mo domains and subdivision of the
polymetallic domains by the copper domains are areas where soft boundaries for
some elements could be considered. In both case the same hard boundaries are
used for all elements to ensure any local correlations with the principal economic
elements are preserved.

14.12     Data Preparation

14.12.1 Database Preparation

The Mount Dore drilling data was supplied by IVA in a Microsoft Access database
export format, as used by IVA for Micromine software presentation and processing.
The data was briefly validated at several stages for cross table conformity and
integrity. Minor corrections identified by Golder were undertaken by IVA on an
ongoing          basis.         The            final     dataset             delivered
“acQuire_MD_Drilling_Database_Jun_27.mdb” included all data available up until
27 June 2010. Drilling at Mount Dore stopped several weeks prior to this allowing the
assays to be completed for the Prefeasibility Study.

The data provided included all drilling in and around the Mount Dore area and
included 760 drill holes. This includes many shallow air core drill holes to the south
and drilling at another small prospect to the east. This data was initially subset for the
modelling area defined in Table 14.15 and displayed in Figure 9.1, to provide
491 drill holes relevant to the Mount Dore and Merlin resource study. Twenty-two of
these drill holes have no available assays for the following reasons:




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•     Precollars drilled but extensions abandoned or cancelled (MDQ0087,
      MDQ0144, MDQ0215, MDQ0215A, MDQ0333, MDQ0334, MDQ0347,
      MDQ0349).
•     Geotechnical drill holes for the Merlin decline (MDQ0262, MDQ0263,
      MDQ0265, MDQ0266, MDQ0268 to MDQ0270, MDQ0272, MDQ0291,
      MDQ0311, MDQ0382, MDQ0385).
•     Metallurgical drill holes (MDQ0337, MDQ0338, MDQ0400).
•     Water test drill hole (MDQ0401).
•     No assays available (MDQ0368).

Drill holes with outstanding assays that missed the final database cut off date include
MDQ0368 and extensions to MDQ0220 and MDQ0251. All drilling was used for
interpretation; however unassayed drill holes and intervals listed above were
excluded from the resource estimate, leaving a total of 469 assayed drill holes.

Table 14.7 summarises drilling data by drilling type for collar information. This table
indicates that sampling is dominated by diamond drilling and includes:

Table 14.6       Drilling Data by Drilling Type
                              Pre-IVA DDH                         12.5%
                              Pre-IVA percussion                  7.5%
                              IVA DDH                              60%
                              IVA RC                               20%

RC drilling by IVA includes an initial RC drilling program in 2008 of 39 drill holes.
Other RC drilling by IVA includes pre-collar drilling for diamond drill tails, with the RC
mostly in expected waste zones in the granite overburden. Pre-IVA drilling by
previous companies includes a number of drilling types including RC, percussion and
air core drilling.

QG (2008) reviewed RC drilling and spear sampling of wet samples undertaken of
IVA and raised some concerns regarding the sample quality. No test work has been
completed to date to validate the spear sampling used in 2008. Subsequent RC pre-
collar drilling has used on rig riffle splitters.

Comparing the collar and assay results in Table 14.8 indicates the degree of
completeness of assaying for each drill program. As drilling has become deeper,
selective sampling processes has resulted in a lower proportion of drilling being
assayed.




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Table 14.7           Assayed Drilling by Year, Company and Type
                                                                                Total              % of
                                  Drill     Hole Name      Hole Name                       RC
 Year     Company      Type                                                     Depth             Drilling Cumula
                                 Holes         Min            Max                          (m )             tive %
                                                                                  (m )
 1976      Pre-IVA     DDH         8       SHQ-76-10       SHQ-76-9              1147              1.0%     1.0%
 1977      Pre-IVA     DDH        15       SHQ-77-15      SHQ-77-31              4729              4.2%     5.2%
 1978      Pre-IVA     DDH         8       SHQ-78-32      SHQ-78-39              3486              3.1%     8.2%
 1979      Pre-IVA     DDH         1       SHQ-79-40      SHQ-79-40               351              0.3%     8.5%
 1989      Pre-IVA     DDH         8       MDQ-89-41      MDQ-89-48              2472              2.2%    10.7%
 1989      Pre-IVA      WB         2       MDWB-89-1      MDWB-89-2               472              0.4%    11.1%
 1992      Pre-IVA     DDH         2       MDQ-92-49      MDQ-92-51               652              0.6%    11.7%
 1993      Pre-IVA     DDH         2       MDQ-92-50      MDQ-93-53               611              0.5%    12.2%
 1994      Pre-IVA     DDH         4       MDQ-94-54      MDQ-94-57               306              0.3%    12.5%
 1999      Pre-IVA     DDH         5       MDQ-99-58      MDQ-99-62              1266              1.1%    13.6%
 2000      Pre-IVA      AT        54        MDAT-1          MDAT-9               1138              1.0%    14.6%
 2000      Pre-IVA     DDH         3       MDQ-00-63       SMD-76-6               725              0.6%    15.2%
 2000      Pre-IVA     RAB        16        MDRAB-1        MDRAB-9                626              0.5%    15.8%
 2000      Pre-IVA      RC        23         MDRC1          MDRC9                2030              1.8%    17.6%
 2000      Pre-IVA      RP        20         RP-10         RP-MD-2               2199              1.9%    19.5%
 2000      Pre-IVA      WB         7        MDWB1           MDWB8                 955              0.8%    20.3%
 2004        IVA       DDH        17      MDHQ-04-0064   MDHQ-04-0080            3561              3.1%    23.5%
 2007        IVA       DDH         6      MDHQ-07-104     MDHQ-07-99             2109       0      1.9%    25.3%
 2007        IVA     RC- DDHT     13      MDHQ-07-101      MDQ0103               5334     1196     4.7%    30.0%
 2008        IVA       DDH        25        FBD0001       MDQ0188a               8650       0      7.6%    37.6%
 2008        IVA        RC        38        MDQ0151        MDQ0207               4872     4872     4.3%    41.9%
 2008        IVA     RC- DDHT     42        MDQ0088        MDQ0209              14288     6307    12.5%    54.4%
 2009        IVA       DDH        63        MDQ0121        MDQ0342              12835       0     11.3%    65.7%
 2009        IVA        RC         1        MDQ0219        MDQ0219                243      243     0.2%    65.9%
 2009        IVA     RC- DDHT     52        FBD0008        MDQ0310              27073     7572    23.8%    89.6%
 2010        IVA       DDH        16        FBD0013        MDQ0387               3303       0      2.9%    92.5%
 2010        IVA     RC- DDHT     18        MDQ0293        MDQ0378               8497     2296     7.5%   100.0%
DDH =     Diamond drilling            WB = Water Bore                   RC =      Reverse Circulation
RP =      Percussion                  AT = Air Track (Air Core)         RAB =     Rotary Air Blast
RC- DDHT= RC pre-collar w ith diamond drilling tail

Table 14.8           Assay Completeness by Drilling Company And Type
Company        Drill Hole Type     Sample Type     Drilled Length (m)      Assays        Assay Completeness (%)
 Pre IVA             AT                   AT             1138               569                   100%
                     DDH                  DDH            15745              5769                   61%
                     RAB                  RAB             626               299                    96%
                     RC                   RC             2030               996                    98%
                     RP                   RP             2199               793                    86%
                     WB                   WB             1427               431                    69%
   IVA               DDH                  DDH            30457             14858                   79%
                     RC                   RC             5115               2373                   92%
                 RC- DDHT                 DDH            32855             16159                   94%
                 RC- DDHT                 RC             22337              1529                   14%
  Total                                                  113928            43776                   69%




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14.12.2 Recovery

Drilling by IVA and previous companies have used some selective sampling
processes, namely to avoid sampling and assaying long runs of footwall or hanging
wall waste where no logged mineralisation occurs. This introduces a potential bias in
the interpretation and estimation process.

The drill data were prepared to best identify loss from unsampled core using the
available recovery information and some assumptions. The process involved:
•     Missing unsampled drill hole intervals were generated in an Access database
      software from the collar information that includes the total hole depth.
•     All unsampled intervals were broken down by any available recovery logs
      (based on core run intervals).
•     Assigned length weighted core recovery averages to each drill hole interval
      (accounting for sample intervals that can span recovery or core run intervals).
•     Recovered core was then designated as either sampled or unsampled.
•     Since the recovery logs are in places incomplete (ie the database does not
      identify 0% recovery core runs), lost core was identified as unsampled
      intervals. Lost core was flagged if the recovery was either:
      -      0% recover.
      -      <90% recovery where the recovered length was <1 m (assumed the
             interval to sample).
      -      Missing recovery and the interval <3 m where up and down hole
             intervals do have recovery details (ie missing recovery logs in holes that
             do have recovery available).
      -      Interval <1 m and no recovery logs for the drill hole.
      -      No recovery logs at the end of hole.
•     Unsampled drilling was identified as either:
      -      Recovered length >2 m.
      -      Recovery >40%.
      -      Sample type was RC or other percussion type and sample interval >2 m.
      -      Remaining intervals >3 m with no recovery logs.
•     Finally lost or unsampled flags were copied down hole to neighbouring
      unflagged intervals to fill the remaining sample intervals.

Table 14.11 summarises the tagged values for the samples as either:
•     Some recovery and sampled, ie where assay are available.
•     Some recovery and not sampled, ie where defaults should be assigned for
      economic values as no mineralisation was logged to warrant sampling.
•     Lost where no assay could be taken and the interval treated as missing.




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The average recovery of 96% for all IVA diamond drilling should be revised down by
1% to account for the lost core with no recovery records entered into the database.
Recovery data is available for 85% of the IVA drilling.

QG (2009) notes poor recovery near surface at Mount Dore South in the earlier IVA
drilling. The available data does not suggest this observation and further
investigation is required.

14.12.3 Dry Bulk Density

Assignment of density values to the assay table in acQuire was noted to result in
some irregularities where a single density measurement from a ~20 cm specimen
was assigned to intervals up to 16 m long. Subsequent compositing would result in
multiple duplicates of the same data potentially biasing any analysis.

To avoid this issue the sampled or unsampled drill intervals >2 m were broken into
2 m subintervals using a Visual Basic macro, prior to compositing. Dry bulk density
values were then averaged for the available drill hole intervals before being assigned
to the drill hole interval. This provides a database where closely spaced density
specimens are averaged for the same sample interval providing:
•     A better basis for statistical analysis and sample support for estimation
      purposes.
•     A database where density can be paired with assay data.
•     An avoidance of any duplication of the individual density values.

Table 14.9 summarises the dry bulk density data for each combined mineralisation-
lithology domain and weathering type. Bulk density varies slightly with the degree of
oxidation; but there is relatively little variation in density for the different lithology and
mineralisation domains.

The consistency of dry bulk density samples taken across Mount Dore-Merlin can be
attributed to the relative lack of other sulphide minerals such as pyrite and the
pervasive alteration that has made most lithology type similar. Even Merlin (Dom=6)
which contains high grade molybdenite, a very dense mineral, does not appear to
display larger density values on average. A closer inspection of the relationship
between Mo and density does indicate a weak positive relationship (refer to Figure
14.8); this does not appear to have resulted in a higher average density for the
domain that would reflect the high Mo grade. Table 14.10 indicates that if the Merlin
samples are weighted by the density values the average grade increases by 3%.
This would indicate the potential bias if a blanket average density was assigned to
the entire domain.




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Table 14.9       Dry Bulk Density Averages by Domain and Weathering Type
 Dom ain       Dom ain                        Average Density                                      Number of Sample Intervals
  (DOM)       Description             Oxide          Trans                            Fresh       Oxide          Trans        Fresh
    1        Waste: Granite           2.36           2.49                             2.58         35            195           159
    2        Waste: Kuridala          2.27           2.55                             2.68         56            283           590
    3          Upper Cu                              2.54                             2.68                       450           568
    4           Upper Zn              2.24           2.59                             2.67         26            353           570
    5        Waste: Kuridala          2.43           2.50                             2.69         23            643           821
    6          Low er Mo                             2.54                             2.64                        53           373
    7          Low er Cu              2.54           2.46                             2.67         72            1631          853
    8           Low er Zn                            2.48                             2.66                       140           667
    9        Waste: Footwall                         2.55                             2.63                        83           297
  Total                               2.39           2.50                             2.67        212            3831         4898



Table 14.10       Merlin Density Samples                                Figure 14.8                Merlin Mo vs. Density
             Density                                                            3.4
Samples                     Mo Mean      Mo Wt Mean
              Mean
                                                                                3.2
  482          2.63            1.55           1.60
                                                                                 3
                                                             Dry Bulk Density




                                                                                2.8

                                                                                2.6

                                                                                2.4

                                                                                2.2
                                                                                      0       5    10      15      20    25       30
                                                                                                          Mo %



14.12.4 Default Grades

As discussed in Section 14.12.2 selective sampling of drill core can result in
estimation bias when only samples with visible mineralisation are selected for
assaying. This bias is present in only a few early drill holes within the mineralised
zones but is more systematic in the peripheral waste zones, where the estimation of
low grade mineralisation, waste and possible dilution material may become biased if
the selective sampling is not addressed. For grade estimation purposes all
unsampled intervals were assigned default grade of 0.001 for all potentially economic
elements including Cu, Mo, Re, Zn, Pb, Au, Ag and Co. Other non economic
elements such as Fe and S were left as missing. Table 14.11 summarises the
proportion of each domain that is unsampled and set to default grades and that
designated as lost core and left as missing and ignored during compositing.

14.12.5 Domain Flagging

Samples for the domains described in Section 14.1 were flagged in Vulcan software
for samples lying inside the mineralisation solid wireframe models or above/below
the oxidation, lithology and metallurgy surface wireframe models.




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Table 14.11      Proportion of Sampled, Unsampled and Lost Sample Interval by
                 Company and Drill Type
                                  Sampled     Lost    Unsampled   Sampled   Lost   Unsampled
    Company          Drill Type
                                     m         m          m         %        %        %
                        DDH        55060      755        7484      87%      1%       12%
       IVA
                        RC         7743        21        19701     28%      0%       72%
                        DDH        9655        28        6062      61%      0%       39%
Previous Companies
                     Percussion    6619        7          794      89%      0%       11%
      Total            Total       79076      811        34041     69%      1%       30%

14.12.6 Compositing

The majority of the drilling has been sampled on regular 2 m intervals. Only some
sub-sampling or irregular intervals were undertaken by IVA. Limited early drilling prior
to IVA was sampled on less regular intervals.

Though the open pit estimation for Cu oxide resource could justify a larger composite
size, 2 m was retained for all estimation and analysis as the preferred composite
length as this will support selective open pit and underground mining options.

Compositing the available data to 2 m presents potential variance deflation issues in
some selected areas as there are a few drill holes where sub-sampling down hole at
a geological contact has not always been followed by a regular sampling protocol. In
most cases sampling returned to the 2 m regular intervals set from the drill collar, but
is some case the sampling returned to 2 m intervals from the geological contact. This
presents an issue for generic compositing routines to achieve composite files that do
not result in variance deflation somewhere. An adapted “smart” compositing
approach was adopted whereby samples were only composited in Vulcan software if
the samples were not within the desired length of 1.9 to 2.0 m, noting that all
intervals larger than 2 m were already split to a maximum of 2 m. This was
implemented by first assessing the database using a Visual Basic program to
determine if the samples required compositing.

Lost core defined in Table 14.11 was treated as missing during compositing where
as unsampled intervals were assigned default values for all elements with potential
economic significance (other elements were treated as missing).

Majority rules were used to preserve the flagged domain codes. The composites
were assessed visually and statistically along the mineralisation boundaries to
ensure the original sub 2 m intervals still honoured the boundary interpretations
which were based on the original sample intervals. This involved adjusting the
interpretation for Cu and polymetallic domain to reflect 2 m selectivity and enforcing
the Mo domain boundary to break the compositing.

Figure 14.9 displays the progression of the sample database from original intervals
(top left), split samples (top right) and final 2 m composites (bottom). For the entire
database there are relatively few sub 2 m composites remaining after compositing.




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Figure 14.9                             Distribution of all Sample and Composite Lengths
                                 60%                                                                                                                                       90%
                                        Original Drilling Intervals                                                                                                                                                  Split Drilling Intervals
                                                                                                                                                                           80%
                                 50%
                                                                                          Unsampled Interval (31%)                                                                                                                                 Unsampled Interval (31%)
  Proportion of Total Drilling




                                                                                                                                            Proportion of Total Drilling
                                                                                                                                                                           70%

                                 40%                                                      Sampled Intervals (69%)                                                          60%                                                                     Sampled Intervals (69%)

                                                                                                                                                                           50%
                                 30%
                                                                                                                                                                           40%

                                 20%                                                                                                                                       30%

                                                                                                                                                                           20%
                                 10%
                                                                                                                                                                           10%

                                 0%                                                                                                                                         0%




                                          10
                                          20
                                          30
                                          40




                                                                                                                                                                                     10
                                                                                                                                                                                     20
                                                                                                                                                                                     30
                                                                                                                                                                                     40
                                        >50




                                                                                                                                                                                   >50
                                       <0.05




                                                                                                                                                                                  <0.05
                                           3
                                           4
                                           5
                                           6
                                           7
                                           8
                                           9




                                                                                                                                                                                      3
                                                                                                                                                                                      4
                                                                                                                                                                                      5
                                                                                                                                                                                      6
                                                                                                                                                                                      7
                                                                                                                                                                                      8
                                                                                                                                                                                      9
                                         0.1
                                         0.2
                                         0.3
                                         0.4
                                         0.5
                                         0.6
                                         0.7
                                         0.8
                                         0.9
                                         1.0
                                         1.1
                                         1.2
                                         1.3
                                         1.4
                                         1.5
                                         1.6
                                         1.7
                                         1.8
                                         1.9
                                         2.0
                                         2.1
                                         2.2
                                         2.3
                                         2.4
                                         2.5
                                         2.6
                                         2.7
                                         2.8
                                         2.9




                                                                                                                                                                                    0.1
                                                                                                                                                                                    0.2
                                                                                                                                                                                    0.3
                                                                                                                                                                                    0.4
                                                                                                                                                                                    0.5
                                                                                                                                                                                    0.6
                                                                                                                                                                                    0.7
                                                                                                                                                                                    0.8
                                                                                                                                                                                    0.9
                                                                                                                                                                                    1.0
                                                                                                                                                                                    1.1
                                                                                                                                                                                    1.2
                                                                                                                                                                                    1.3
                                                                                                                                                                                    1.4
                                                                                                                                                                                    1.5
                                                                                                                                                                                    1.6
                                                                                                                                                                                    1.7
                                                                                                                                                                                    1.8
                                                                                                                                                                                    1.9
                                                                                                                                                                                    2.0
                                                                                                                                                                                    2.1
                                                                                                                                                                                    2.2
                                                                                                                                                                                    2.3
                                                                                                                                                                                    2.4
                                                                                                                                                                                    2.5
                                                                                                                                                                                    2.6
                                                                                                                                                                                    2.7
                                                                                                                                                                                    2.8
                                                                                                                                                                                    2.9
                                              Interval Length (m)                                                                                                                                                                 Interval Length (m)
                                                                                   120%
                                                                                                                   2m Composited Drilling Intervals
                                                                                   100%
                                                                                                                                                                                         Unsampled Interval (31%)
                                                    Proportion of Total Drilling




                                                                                   80%                                                                                                   Sampled Intervals (69%)


                                                                                   60%



                                                                                   40%



                                                                                   20%



                                                                                    0%


                                                                                             10
                                                                                             20
                                                                                             30
                                                                                             40
                                                                                           >50
                                                                                          <0.05




                                                                                              3
                                                                                              4
                                                                                              5
                                                                                              6
                                                                                              7
                                                                                              8
                                                                                              9
                                                                                            0.1
                                                                                            0.2
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                                                                                            0.5
                                                                                            0.6
                                                                                            0.7
                                                                                            0.8
                                                                                            0.9
                                                                                            1.0
                                                                                            1.1
                                                                                            1.2
                                                                                            1.3
                                                                                            1.4
                                                                                            1.5
                                                                                            1.6
                                                                                            1.7
                                                                                            1.8
                                                                                            1.9
                                                                                            2.0
                                                                                            2.1
                                                                                            2.2
                                                                                            2.3
                                                                                            2.4
                                                                                            2.5
                                                                                            2.6
                                                                                            2.7
                                                                                            2.8
                                                                                            2.9




                                                                                                                                    Interval Length (m)


Recent sampling protocols by IVA allow for sub-sampling and assaying <2 m
intervals in and around the Merlin high grade Mo domains. Figure 14.10 compares
the distribution of original samples and final 2 m composites for the high grade Mo
domains at Merlin. This indicates that some 1 and 1.5 m samples remain which
require length weighting during estimation to avoid estimation biases that may result
from the shorter intervals.

Figure 14.10 Distribution of Merlin Domain Sample and Composite Lengths
                                                                                   100%
                                                                                                       Narrow Merlin Molydbenite Domains
                                                                                   90%

                                                                                                       Original Intervals
                                                    Proportion of Total Drilling




                                                                                   80%

                                                                                   70%
                                                                                                       2m Composites
                                                                                   60%

                                                                                   50%

                                                                                   40%

                                                                                   30%

                                                                                   20%

                                                                                   10%

                                                                                    0%
                                                                                          0.1
                                                                                                0.2
                                                                                                      0.3
                                                                                                            0.4
                                                                                                                  0.5
                                                                                                                        0.6
                                                                                                                              0.7
                                                                                                                                    0.8
                                                                                                                                          0.9
                                                                                                                                                     1.0
                                                                                                                                                                           1.1
                                                                                                                                                                                 1.2
                                                                                                                                                                                       1.3
                                                                                                                                                                                             1.4
                                                                                                                                                                                                   1.5
                                                                                                                                                                                                         1.6
                                                                                                                                                                                                               1.7
                                                                                                                                                                                                                     1.8
                                                                                                                                                                                                                           1.9
                                                                                                                                                                                                                                 2.0
                                                                                                                                                                                                                                       3.0




                                                                                                                                    Interval Length (m)


For only the mineralised domains, where selective sampling and the effect of default
grades is minimal, it is possible to compare the length weighted average grades for
the original samples and 2 m composites, see Table 14.12. The metal content is
expected to increase due to both the addition of default grades to unsampled
intervals and the handing of lost core which has been ignored and effectively
assigned grades from neighbouring samples. The results for Cu indicate a small
decrease in average grade with a similar metal content for Cu which is assayed




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consistently across the deposit except for a few older drill holes where default grades
have reduced the mean grade. Significant decreases in the mean grade and minor
increase in metal content for Mo and Re and to a lesser extent Pb and Zn are a
result of the addition of a significant proportion of default grades; mostly from early
Mount Dore South drilling that was completed before the discovery of the Merlin and
polymetallic zones further to the north. The default samples are mostly situated in the
oxide copper domains of Mount Dore South where Zn and Mo mineralisation is not
recognised to date. Though such mineralisation could exist the addition of the default
grade will ensure neighbouring grades are not over extrapolated past drilling where
logging did not recognise such mineralisation.

Table 14.12      Average Length-Weighted Grades Before and After Compositing
                 Original Intervals                 2 m Composites                  Difference
Element               Mean      Complete                Mean         Complete-   Mean
            Length                         Length                                           Metal
                      Grade      -ness                  Grade          ness      Grade
  Total     42272                           42265
  Cu %      41263     0.377       98%       42226       0.372          100%      -1.3%       1.0%
 Au ppm     40267     0.081       95%       42226       0.078          100%      -3.6%       1.1%
 Ag ppm     38922     4.421       92%       42226       4.113          100%      -7.0%       0.9%
 Co ppm     36289     65.40       86%       42226       56.67          100%      -13.3%      0.8%
  Fe %      32276     3.096       76%       32606       3.101          77%       0.2%        1.2%
  S%        31097     0.860       74%       31426       0.856          74%       -0.5%       0.5%
  Pb %      36735     0.046       87%       42226       0.041          100%      -12.1%      1.1%
  Zn %      36734     0.267       87%       42226       0.233          100%      -12.5%      0.6%
Mo ppm      32354     0.072       77%       42226       0.057          100%      -21.5%      2.5%
 Re ppm     25116     1.509       59%       42226       0.911          100%      -39.6%      1.5%
 Density    10590     2.585       25%       11445       2.582          27%       -0.1%       8.0%

14.12.7 Declustering

Declustering was used remove potential biases in sample statistics that can arise
from variable drill density. This can arise from purpose driven infill drilling of higher
grade areas and local clustering and drill spacing variations due to inherent issues
with placing drill holes and completing a deep drilling program used to define the
Merlin mineralisation.

Normal cell declustering without any boundaries can present issues for structural
deposits where the edge cells become overweighted as the cell size is increased. A
modified cell declustering algorithm was used that weights the cells to the block
model volume within each cell. This allows domains to be declustered separately
without introducing any boundary issues.

The cell size was optimised for a cubic window size of 100 m and the origin offset 10
times. The calculated weights were used in addition to the sample length to weight
samples for declustered statistics and log-probability plots.




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14.12.8 Top-Cutting

The presence of outliers (or ‘extreme’ values), and the need to apply ‘top-cut’ values
(or ‘capping’, where samples above a certain threshold are assigned the top-cut
value) to sample populations was assessed using a number of techniques:
•     Examination of grade distributions using box plots, histograms and probability
      plots.
•     Statistical assessment of the grade distributions.
•     Examination of the spatial locations of identified outlier samples and the
      relative impact of the default grades.

The combination of domains and oxidation creates a large number of combinations
where cutting and grade estimation division could be considered. Statistical analysis
of these domain combinations indicates that some elements display different
behaviour and can be arranged in groups suitable for both estimation and grade
cutting as follows:
•     Mo and Re are both highly correlated to the occurrence of Molybdenite, see
      Golder 2010. This is a mineral that will often persist in the weathering zone
      though most of the Merlin deposit dissipates before reaching the surface or
      the oxide zone. The only significant occurrence of the Merlin Mo mineralisation
      in the oxide is at Little Wizard (MODOM=4) where massive high grade
      molybdenite occurs along with other sulphide minerals that have withstood
      any oxidation. As a result it is not considered likely that Mo or Re should be
      affected by the weathering profile with lower grades in the oxide related to
      weaker peripheral mineralisation rather than oxidation effects. Note IVA is yet
      to analyse for the presence of Mo oxide minerals as most core visually
      indicates that molybdenite persists in the oxide zone.
•     The Little Wizard zone (MODOM=4) within the Merlin high grade Mo domains
      indicates significantly higher grades populations for Cu, Mo, Re and Au and
      lower grades for Co, Zn and Pb, see Golder 2010. Though this domain is
      relatively small, it’s potential economic significant for the mine development
      indicates that separate assessment and grade cutting is warranted.
•     S, Zn and Pb display significant changes in grade distribution through the
      weathering profile between oxide, transition and fresh material type (MINL=2,
      3, 4 respectively). In each case separate cutting and grade estimation is
      warranted.
•     Cu, Co, Au and Ag display reasonable similarity between transition and fresh
      materials with different grade distributions evident in the oxide, see Appendix
      A, Golder 2010. The oxide is generally lower grade although for Au there are
      some cases where small higher grade pods are noted in the oxide; these are
      thought to be local Au accumulations but could also be a result of some
      remnant supergene enrichment. For these elements the oxide is treated
      separately to the transition and fresh zones.
•     Cu displays some indication that the transition zone is higher grade sub-
      population both in north (upper sequence) and south Mount Dore. This was
      evident during sectional interpretations and in statistical plots broken down by




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         weathering. It is not evident if this is a natural variation where the transition
         just happens to coincide with the central and higher grade portion of the
         mineralisation or if supergene enrichment of Cu in the transition has produced
         some of the higher grade zones down open fractures. The differences are
         sufficient to warrant separate grade cutting for Cu in the transition and fresh
         zones.

 Top cuts are those defined in Table 14.13. Distributions before top cuts can be
 compared to those after cutting in Appendix A, Golder 2010.

 Table 14.13      Top Cuts by Grouped Domains
  Dom ain                       Cu     Au     Ag    Co    Fe   S     Pb     Zn      Mo     Re   Density
                Description
 Condition                      %     ppm    ppm   ppm    %    %     %      %      ppm    ppm    t/m 3
 MODOM=4        Little Wizard   5.5   1.5    50    70     25   14    0.02   0.02   20     300     3.2
MODOM=5,6,7        Merlin       3     0.7    50    400    25   10    0.4     3     10     150     2.9
  DOM=3,4       North Upper
                Seq. Oxide      1      2     15    200    25   0.4   0.1    0.6    0.1    0.2     2.7
   MINL=2
  DOM=3,4       North Upper
                                4     0.75   50    500    25   10    0.6     2     0.5     2      3.2
   MINL=3       Seq. Trans
  DOM=3,4       North Upper
                Seq. Fresh      2     0.75   50    500    25   10    0.6    3.5    0.5     2      3.2
   MINL=4
  DOM=7,8       North Low er
                                1      2     15    200    25   0.1   0.1    0.6    0.1    0.2     2.7
   MINL=2       Seq. Oxide
  DOM=7,8       North Low er
                 Seq. Trans     3     0.75   50    400    25   5      1      1     1.0    10      3.2
   MINL=3
  DOM=7,8       North Low er
                                3     0.75   50    400    25   10    0.6    3.5    1.0    10      3.2
   MINL=4        Seq. Fresh
     DOM=7         South
                                4      2     15    900    25   0.1   0.1    0.2    0.1    0.2     2.7
     MINL=2        Oxide
     DOM=7         South
                                5.5   0.75   50    900    25   5     0.2     1     0.1     5      3.2
     MINL=3        Trans
     DOM=7         South
                                4     0.75   50    1200   25   10    0.3     3     0.1     5      3.2
     MINL=4        Fresh
DOM=1,2,5,9        Waste
                                0.2   0.6    20    100    25   0.1   0.3    0.1    0.04   0.2     2.7
  MINL=2           Oxide
DOM=1,2,5,9        Waste
                                0.3   0.5    20    200    25   1     0.4    0.4    0.1     1      2.8
  MINL=3           Trans
DOM=1,2,5,9        Waste
                                0.3   0.5    20    200    25   2.5   0.4    0.8    0.1     1      3.2
  MINL=4           Fresh


 The approach taken to grade cutting was to implement a consistent statistical
 approach by cutting all economic elements to roughly the 99th percentile based on
 the original assay grades, without the inclusion of default values. The values were
 rounded and simplified to maintain similarity between domains where justified. The
 approach included:
 •       Little Wizard and Merlin were considered separately for all elements.
 •       Copper mineralisation for Mount Dore South and North (using 7,604,900 mN)
         were considered separately in case of there being significant difference due to
         the change in weathering depth.




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•       Mineralisation for Mount Dore North was subdivided into the upper and lower
        mineralisation sequences.
•       Weathering was considered as previously described.

14.13      Data Analysis

14.13.1 Density Data

The dry bulk density data was supplied in a separate table to the other assay data,
and was sometimes represented as a discrete point in a drill hole as opposed to a
down hole interval. In general, the length of core used for the density determination is
about 10 cm, and it is common practice to assign this density value to the
corresponding down hole sample length used for assaying. Although this has been
done for some density data, density is left as points elsewhere. Compositing to an
equal length cannot be achieved if original down hole intervals are not known.

QG notes that the samples for density were taken on essentially equal support,
therefore compositing is not strictly necessary.

It is common in resource estimation to density-weight assay data during compositing,
but this usually requires the density data to be as extensively collected as the assay
data. Due to the mismatch between the frequency of density and assay data in this
case, the assay compositing has not been density-weighted.

The density data was flagged with the same domain codes as for the assayed
variables, and six separate zones for geostatistical analysis and estimation were
established:
•       Kuridala formation (outside Mo domains), oxide.
•       Kuridala formation (outside Mo domains), transitional.
•       Kuridala formation (outside Mo domains), fresh.
•       Medium grade Mo (including hanging wall), oxide + transitional.
•       Medium grade Mo, fresh.
•       High grade Mo (all in fresh).

14.13.2 Statistics

Summary statistics for the composited data for all domains, for the combined
domains, along with log-probability plots by individual and combined domains are
presented in Appendix A. Table 14.21 also displays the changes from cutting
including the metal loss and decrease in variance. As expected, many of the
domains that previously had very high coefficients of variation (CV) have significantly
lower CV’s after top grade cutting.

Multivariate statistics are presented in Appendix A as Pearson correlation coefficients
by domain in Appendix A, Golder 2010. This demonstrates some correlation is
present for elements in some domains which are not present globally. Notably the




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correlation between Mo-Re and other elements like S-Au-Ag-Cu increase within the
high grade Mo domain at Merlin compared to other domains. This strengthens the
rationale for ensuring that high grade domains such as those for Mo and Cu are
estimated with all elements constrained in the same envelope as there is sufficient
positive correlation present between many elements.

14.13.3 Variography

Cut 2 m composites were used for all variogram analysis. Traditional semi-
variograms provided erratic results in some instances and a combination of semi-
variograms and inverted correlograms were used as a basis for all variogram
modelling. All semi-variograms were scaled to the domain variance to enable the
different variogram and correlogram types to be overlayed at the same scale.

Variogram maps for the major element domains were assessed to ensure that the
variograms support the shape and structure of the domains, which predominantly
strike N-S and dip at average of 50° towards the east. None of the variograms
indicated any significant trend within the plane of mineralisation. This lack of plunge
was also the case for Mo and Re where the overall grade trends suggest a northerly
plunge might be present. The Oxide Cu suggests a greater continuity N-S than down
dip. Despite this indication the geological interpretation still indicates reasonable
down dip continuity for Cu, though this may be lost for Zn which is leached and
occurs only in deeper patches in the oxide zone.

Variogram models were primarily based on the inverted experimental correlograms
and are plotted in black in all figures. In all cases the experimental semi-variogram
was overlain (in green) and used to assist the interpretations. The semi-variograms
suffer from being both erratic and not converging to the expected sill due to
occurrence of grade pods within the resource. The semi-variograms were used as a
guide to the shape and suitability of the correlograms for modelling and estimation.

The dip of the principal mineralisation varies along the strike length from near vertical
in the deeper sections of the Merlin Mo domains to near horizontal in flexures
towards Mount Dore South for Cu domains. To remove these structural changes
simple unfolding of the variograms and correlograms to a N-S reference plane
dipping 50° towards the east was used for the principal model orientations. The
validity of the variogram model derived from the unfolded composites was verified by
overlaying the variogram models with directional semi-variograms and correlograms.

Down hole variograms with a lag of 2 m used for all cross strike model orientations.
Down hole orientations vary from 60° to 90° towards the west and represent oblique
intersections. Separate assessment of variograms at 60° to 90° was found to provide
some conflicting results as the different orientations tend to be clustered. In most
cases the down hole data from the various orientations were pooled to provide more
robust results. However some small longer range structures were ignored in some
instances where these were found to be only present in the vertical drill holes and
likely to be slight down dip effects from the more dip parallel drilling.




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The majority of the drilling is spaced at 50 m. There is only limited regular 25 m infill
drilling with other closer spaced drilling occurring in only some localities, such as at
crossing drill holes, twin holes, near surface Mount Dore South drilling, 12.5 m infill at
Little Wizard etc. Since many variogram structures appear to be less than the 50 m
general spacing 12.5 m lag was selected for unfolded experimental variograms and
correlograms in order to provide sufficient lags to model the indicated structures.
Longer 25 m lags were used for the directional variogram validation.

Additional variogram parameters used in most cases for the unfolded variogram
calculation include:
•     One or two structure spherical scheme models.
•     Horizontal angle of 45° for modelling and 15° for assessing possible plunges.
•     Vertical angle of 25°.
•     Horizontal distance of 50 m within the unfolded plane.
•     Vertical distance of 6 m across the unfolded plane.

Variogram models used a two structure spherical scheme model in all cases.

The relatively short ranges evident for most variograms and the use of a short lag to
help define these structures, at a quarter of the general drill spacing, make the
variogram model particularly sensitive to the limited clusters of close space drilling
and localised data arrangements.

Initial analysis indicated considerable similarity between the variogram structures for
some element groups, notably:
•     Zn - Pb.
•     Cu - Co - Au - Ag.
•     Mo - Re.
•     S - Density.
•     Fe.

Where possible, variogram model parameters were retained at similar values
between orientations, domains and elements, so as not to produce artefacts in the
estimations. In particular this was done to ensure element relationships or
correlations evident between samples are respected implicitly during estimation and
reflected in the resource estimate.

Evidence for anisotropy between strike and down dip orientations was only noted in a
few cases where the variograms indicated anisotropy was evident, notably for Merlin
Mo-Re.

Closely related elements such as Mo-Re and Pb-Zn used the same variogram model
based on the more significant element, i.e. Mo and Zn respectively, though the other
elements were considered in the modelling process.




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Co, Au and Ag displayed reasonable similarity to the Cu variograms. In most
domains the Cu variogram model is used for Co, Ag and Au with satisfactory results.
Even though there are some small accumulations of Au and Co that do not correlate
with Cu it appears that generally these four elements display sufficient correlation
and similar behaviour for Cu to provide a reasonable indication of the continuity,
despite their lower less significant abundance.

Variogram modelling was undertaken on the basis of the domains derived for each
element group and used for estimation purposes.

Variogram models are summarised in Table 14.14. Mo, Cu and Zn variogram models
for the principal mineralised domains are presented in Figure 14.11 and Figure 14.12
and demonstrate the continuity models typical for each element.

Figure 14.11 Mo & Cu Variogram Models for the Merlin High Grade Mo Domain




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Figure 14.12 Cu & Zn Variogram Models for the Lower Cu & Zn Domains




Density data is not as abundant as the other elements and has a relatively narrow
range in values compared to the expected measurement error. As a result the
density variograms were particularly erratic. Where structure was indicated the
density variograms displayed a similarity to those for S. This correlation has some
geological basis and S was used to indicate the Density variogram model in most
cases.




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Table 14.14       Variogram Models
    Dom ain               Nugget   Minor Axis 55 -> 90   Sem i-major Axis 0 -> 0   Major Axis 0 -> 90
                Element
    Groups                  C0     C1    H1   C2    H2   C1    H1    C2      H2    C1    H1   C2    H2
                  Ag        0.2    0.3   7    0.5   15   0.4   10    0.4     45    0.4   10   0.4   45
                  Au        0.2    0.3   7    0.5   15   0.4   10    0.4     45    0.4   10   0.4   45
                  Co        0.2    0.3   7    0.5   15   0.4   10    0.4     45    0.4   10   0.4   45
Oxide             Cu        0.2    0.3   7    0.5   15   0.4   10    0.4     45    0.4   10   0.4   45
                  Fe        0.1    0.4   7    0.5   20   0.6   15    0.3     50    0.6   15   0.3   50
Copper &
                  Pb        0.2    0.3   10   0.5   25   0.5   20    0.3     40    0.5   20   0.3   40
Poly metallic
Domains           SG        0.1    0.4   7    0.5   20   0.6   15    0.3     50    0.6   15   0.3   50
                  S         0.1    0.4   7    0.5   20   0.6   15    0.3     50    0.6   15   0.3   50
                  Zn        0.2    0.3   10   0.5   25   0.5   20    0.3     40    0.5   20   0.3   40
                  Mo        0.3    0.3   3    0.4   29   0.3   10    0.4    100    0.3   10   0.4   100
                  Re        0.3    0.3   3    0.4   29   0.4   10    0.4    100    0.3   10   0.4   100
                  Ag        0.2    0.4   7    0.4   20   0.6   15    0.2     90    0.6   15   0.2   90
                  Au        0.2    0.4   7    0.4   20   0.6   15    0.2     90    0.6   15   0.2   90
                  Co        0.2    0.4   7    0.4   20   0.6   15    0.2     90    0.6   15   0.2   90
Fresh &
                  Cu        0.2    0.4   7    0.4   20   0.6   15    0.2     90    0.6   15   0.2   90
Trans
                  Fe        0.1    0.4   6    0.5   25   0.5   15    0.4     45    0.5   15   0.4   45
Copper &          Pb        0.2    0.4   5    0.4   15   0.4   20    0.4     85    0.4   20   0.4   85
Poly metallic     SG        0.1    0.4   6    0.5   25   0.5   15    0.4     45    0.5   15   0.4   45
Domains
                  S         0.1    0.4   6    0.5   25   0.5   15    0.4     45    0.5   15   0.4   45
                  Zn        0.2    0.4   5    0.4   15   0.4   20    0.4     85    0.4   20   0.4   85
                  Mo        0.3    0.3   3    0.4   29   0.3   10    0.4    100    0.3   10   0.4   100
                  Re        0.3    0.3   3    0.4   29   0.4   10    0.4    100    0.3   10   0.4   100
                  Ag        0.2    0.8   11    0     0   0.4   20    0.4     75    0.4   20   0.4   75
                  Au        0.2    0.8   11    0     0   0.4   20    0.4     80    0.4   20   0.4   80
                  Co        0.2    0.8   11    0     0   0.4   20    0.4     80    0.4   20   0.4   80
Merlin            Cu        0.2    0.8   11    0     0   0.4   20    0.4     80    0.4   20   0.4   80
High              Fe        0.2    0.3   3    0.5   15   0.4   20    0.4     80    0.4   20   0.4   80
Grade
                  Pb        0.2    0.3   3    0.5   15   0.4   20    0.4     80    0.4   20   0.4   80
Mo
Domain            SG        0.2    0.3   3    0.5   15   0.4   20    0.4     80    0.4   20   0.4   80
                  S         0.2    0.3   3    0.5   15   0.4   20    0.4     80    0.4   20   0.4   80
                  Zn        0.2    0.3   3    0.5   15   0.4   20    0.4     80    0.4   20   0.4   80
                  Mo        0.2    0.4   3    0.4   12   0.3   30    0.5     60    0.3   30   0.5   70
                  Re        0.2    0.4   3    0.4   12   0.3   30    0.5     60    0.3   30   0.5   70

14.13.4 QKNA

Quantitative Kriging Neighbourhood Analysis (QKNA) was undertaken to assist the
search parameter selection. This used three trial blocks to determine the sensitivity
of the model estimates to the selection of the estimation parameters, and in particular
search radius and number of samples. QKNA analysis is method of assessing and
measuring the benefit/disadvantages between different search parameters for
ordinary kriging by assessing four parameters including:
•        Slope of regression.




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•     Negative kriging weights.
•     Reliance on simple kriging mean.
•     Kriging variance.

Mount Dore and Merlin are relatively evenly drilled on 50 m grid centres that equates
to a 50 N-S x 70 m down dip drill hole spacing. Exceptions to this general spacing
include (refer to Figure 9.1):
•     Little Wizard drilling at 12.5 m centres for a small target area.
•     At Merlin and Mount Dore North there are two 25 m lines of drilling with 50 m
      spaced drill holes on a diagonally offset grid.
•     Fanning of drill holes to drill the depth extent of Merlin has resulted in some
      closer spaced drilling data in the upper sequences of mineralisation at Mount
      Dore North.
•     A small area of shallow close spaced RC drilling in Mount Dore South.
•     A few sections of wider spaced lines to the south and north of the drilled areas
      where infilling is not complete or not warranted.

The general spacing would indicate a minimum strike and down dip search range of
>50 m is required for adequate sample selection. QKNA analysis confirms that a
minimum 75 m lateral search is required to produce near optimal results.

Block sizes used for parent cell estimation vary from 5 mE x 5 mRL for underground
targets at Merlin to 10 mE x 10 mRL for open pit estimation. The majority of the
resources at Merlin and Mount Dore dip from between 30° to 80° and average around
50°. This angle is not ideal as it increases the width of the search range required to
ensure that sufficient samples are used in the estimation such that the block is wholly
defined and the data is not subset with respect to the true area of influence needed
for parent cell estimation. For this purpose the minimum required cross strike search
range required for estimation should be 8.5 m and 17 m respectively for the two
parent cell sizes.

14.13.5 Geological Block Model

The block model PFS1006v2.bmf has 1,255,208 blocks created in Maptek Vulcan
Software V8.02 using multiple parent/maximum block sizes for different domains, as
defined in Table 14.15. These parent block sizes correlate to the parent block size
used for grade estimation with Ordinary Kriging.

Domain and category fields supplied for the model are defined in Table 14.16.
Domain values are previously defined in Table 14.4 and Table 14.5. Sub-blocking is
only undertaken for the mineralisation domain boundaries and the top and bottom of
the Kuridala Formation. Other boundaries for oxidation, metallurgical domain,
classification and other geological boundaries were flagged from wireframes and are
not sub-blocked. This is commensurate with the accuracy of the models which are
largely based on 100 m section interpretations.




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Table 14.15           Model Definition
Model Parameter (m)                                                  Easting            Northing    RL
Model Origin                                                         447,000        7,603,787.5     -300
Model Limit                                                          448,500        7,606,437.5     440
Model Extent                                                          1500               2650       740
Parent Block Size (HW & FW w aste, DOM=1,9)                            20                 25        20
Parent Block Size (Kuridala ore & waste, Dom=2,3,4,5)                  10                 25        10
Parent Block Size (Merlin high grade Mo, DOM=6,8)                       5                12.5        5
Parent Block Size (Little Wizard MODOM=6)                              2.5               6.25       2.5
Sub-block Size                                                         2.5               6.25       2.5


Table 14.16           Model Field Values
                 Variable    Default   Type     Description
                 DOM            0      short    Main domain code
                 MODOM          0      short    Mo min domain flag
                 CUDOM          0      short    Cu min domain flag
                 P_DOM          0      short    Poly metallic domain based on Cu equivalent value
                 ROCK           4      short    Rock Code
                 MINL           2      short    Weathering Flag
                 METDOM        -9      short    Metallurgical domain on Sequential Cu
                 cu            -9       float   estimated % value
                 mo            -9       float   estimated % value
                 re            -9       float   estimated ppm value
                 ag            -9       float   estimated ppm value
                 au            -9       float   estimated ppm value
                 co            -9       float   estimated ppm value
                 fe            -9       float   estimated % value
                 pb            -9       float   estimated % value
                 zn            -9       float   estimated % value
                 s             -9       float   estimated % value
                 sg            -9       float   estimated density t/m3 value
                 RESCAT         3      short    Resource classif ication
                 REGION        -9      short    Mineralisation Package
                 Defaults      -9       float   Number of default grades assigned


Additional categorical values, including the main domain code DOM (derived from
the combination of Cu, Mo and polymetallic domains) are defined in Table 14.17.

Note that each DOM is estimated separately for all elements with the exception of:
•      MODOM: Mo domains are estimated independently as 4 separate
       lenses.
•      CUDOM: Cu domains are estimated independently as 6 separate lenses
       with the exception of lenses 3 and 5 that merge at the boundary of North
       and South Mount Dore at around 7,604,900 mN.




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•       P_DOM: polymetallic domains estimated independently as 3 separate
        lenses.
•       MINL: Oxide is estimated separately to transition-fresh material.
•       Validation of the blocking accuracy was undertaken by both.
•       Visually where blocking accuracy was viewed and corrected, particularly
        with regard to enforcing and minimum 2.5 m horizontal width for the thin
        Merlin high grade Mo interpretations.
•       Volume checks between the individual wireframes and block model
        codes indicate minimal block error (refer to Table 14.18).

Table 14.17       Additional Categorical Value Definition
                  Field     Value   Description
                  dom        1      Waste : HW Granite
                  dom        2      Waste : Upper Kuridala FM
                  dom        3      Min : Upper Seq - Copper + Zinc
                  dom        4      Min : Upper Seq - Zinc
Combined
                  dom        5      Waste : Lower Kuridala FM
Domain
                  dom        6      Min : Low er Seq - High grade Mo (Merlin)
                  dom        7      Min : Low er Seq - Copper + Zinc
                  dom        8      Min : Low er Seq - Zinc
                  dom        9      Waste : FW quartzite etc
                  rescat     1      Measured
Resource          rescat     2      Indicated
Classification    rescat     3      Inferred
                  rescat     4      Unclassif ied
                  region     0      Waste
Region            region     10     Upper Mount Dore North mineralisation package
(Mineralisation
Package)          region     20     Low er Mount Dore North mineralisation package, including Merlin
                  region     30     Mount Dore South mineralisation package




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Table 14.18      Wireframe - Block Model Volume Comparison
                                      Triangulation        Block Model       Difference
Dom ain            Wireframe
                                         Mbcm                 Mbcm       Mbcm             %
                   1006c1.00t             7.436               7.432      0.004       0.06%
                   1006c2.00t            12.357               12.356     0.001       0.01%
                   1006c3.00t             3.950               3.934      0.016       0.40%
CUDOM              1006c5.00t            37.492               37.435     0.057       0.15%
                   1006c6.00t             4.113               4.111      0.002       0.04%
                   1006c7.00t             0.333               0.333      0.000       0.02%
                      Total              65.680               65.600     0.080       0.12%
                   1006m4.00t           0.00623              0.00629     0.000       -1.02%
                   1006m5.00t             1.517               1.518      -0.001      -0.08%
MODOM              1006m6.00t             0.772               0.772      0.000       0.03%
                   1006m7.00t             0.294               0.295      -0.001      -0.25%
                      Total               2.589               2.591      -0.002      -0.07%
                   1006p1.00t            53.153               52.975     0.178       0.34%
                   1006p5.00t            46.457               46.453     0.004       0.01%
P_DOM
                   1006p7.00t             2.232               2.232      0.000       0.01%
                      Total             101.842              101.660     0.183       0.18%


Figure 14.13 displays the wireframe model in plan view to compare their extent. Note
that the polymetallic domain includes the Cu domains and the abrupt southern
extend for the polymetallic domain is only due to its extension along the existing
copper interpretations to extend the interpretation. The polymetallic zone becomes
progressively smaller south of 7,605,000 mN.

Figure 14.14 to Figure 14.21 display example cross sections for the different coded
fields.




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Figure 14.13 Plan View of Wireframes
                            Mo (left), Cu (middle) and polymetallic (right) domains




 Figure 14.14 Section 7,605,450 mN:                        Figure 14.16 Section 7,605,450 mN:
          ROCK (Lithology)                                         MINL (Weathering)




 Figure 14.15 Section 7,605,450 mN:                        Figure 14.17 Section 7,605,450 mN:
              METDOM                                              MODOM (Mo Domain)




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    Figure 14.18 Section 7,605,450 mN:               Figure 14.20 Section 7,605,450 mN:
           CUDOM (Cu Domain)                              DOM (Combined Domain)




    Figure 14.19 Section 7,605,450 mN:               Figure 14.21 Section 7,605,450 mN:
           P_DOM (Polymetallic)                           RESCAT (Classification)




14.13.6 Grade Estimation Parameters

Estimation by Ordinary Kriging (OK) was performed for all domains, with grade
estimation for parent cells listed in Table 14.15 and not the sub-cells. The parent cell
size was used to limit the maximum size of the blocks to different sizes depending on
the mineralisation type and anticipated mining method. The two predominant parent
blocks sizes used for mineralised material includes:
•      10 x 25 x 10 m for mineralised waste and upper sequence mineralised
       sequences including all of Mount Dore South. This includes all areas to
       present a potential open pit target based on the current scoping study
       development plan. This block size is considered a suitable size for estimation
       of open pit copper targets and is well supported by the 50 m drilling grid.
       Though some copper lenses to the north do become narrow they are generally
       at least 10 m in width.
•      5 x12.5 x 5 m for all lower sequence mineralisation at Mount Dore North (ie
       north of 7,604,950 mN). This includes the narrow high grade Merlin high grade
       Mo mineralisation that presents a narrow stope target where blocks muck
       wider than 5 m would present an estimation issue with the majority of the
       parent cell lying well outside the constrained lens. Defined high grade copper
       also presents a narrow lens in places requiring a narrow block width. Previous
       block models by QG and Golder have used a larger parent cell size of 12.5 x
       25 x 10 m and 10 x 25 x 10 m respectively for the mineralisation around the
       Merlin high grade zone. This zone comprises disseminated and stringer
       mineralisation of Mo-Re and additional lenses of Cu and/or Zn. The relatively




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      wide parent block sizes compared to the drill hole composites of 2 m has
      resulted in a lack of any moderate grade mineralisation being represented in
      the model and as a result not of this mineralisation, including some of the
      larger copper lenses, have not previously contributed to the stope design as
      neighbouring dilution or potential low grade stope addition. The smaller block
      size of 5 x12.5 x 5 m has been used for the wider lower grade copper and
      polymetallic zone best rectify this issue.

Search Parameters
Search parameters were selected on the basis of the general drill spaced ~ 50 m and
QKNA results (refer to Section 14.13.4). Three search passes were undertaken as
follows:
•     Pass 1 : 75 x 75 x 15 m (strike, down dip and cross strike orientations).
•     Pass 2: 150 x 150 x 30 m.
•     Pass 3: 300 x 300 x 60 m.

All estimates were restricted to a maximum of samples. The blocks were re-
estimated if insufficient drill holes or number of composites were available on the
previous search pass, such that the estimation passes enforced:
•     Pass 1: Minimum 10 composites from a minimum of 4 drill holes.
•     Pass 2: Minimum 10 composites from a minimum of 4 drill holes.
•     Pass 3: Minimum 2 composites.

Some parameters were varied on the basis of the two principal parent block sizes:
•     Open pit 10 x 25 x 10 m blocks or larger:
      -      Discretisation: 2 x 5 x 2.
      -      Maximum of 9 composites per drill hole.
      -      Maximum of 36 composites.
•     Underground 5 x 12.5 x 5 m blocks or smaller:
      -      Discretisation: 2 x 5 x 2.
      -      Maximum of 5 composites per drill hole.
      -      Maximum of 20 composites.

Since a small number of composites remain at <2 m and their removal from the
narrow high grade Mo zones could result in bias. Hence, kriging estimates weighted
by the composite length were employed to improve local estimation by accounting for
shorter intervals. This is applied during kriging but after the calculation of the kriging
weights. The number of sub 2 m samples is minimal and not considered to be
significant (refer to Figure 14.9).

Variables estimated using ordinary kriging used the variogram models presented in
Table 14.14. Copper and polymetallic domains were estimated separately for oxide




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and transition + fresh material using different variograms model parameters. The
estimated variables included:
•     Potential economic elements Cu, Co, Zn, Pb, Mo, Re, Au and Ag with
      assigned default grades to avoid bias introduced by selective sampling of drill
      core.
•     Non-economic elements Fe and S without assigned default grades.
•     Dry bulk density measurements averaged for composite intervals.

Estimation used Golder in-house estimating routines that allow unfolding to individual
surface or solid model. These were used to account for the small variations in dip
that occur throughout the deposit. Most mineralised domains were unfolded to the
middle of each mineralised wireframe (refer to Table 14.4). The main lower
polymetallic domain (P_DOM=5) contains a bifurcation which complicates unfolding.
A simple surface model following the top of the main Cu of Mo mineralisation was
used for unfolding in this case. This surface was also used for unfolding the Kuridala
Formation unmineralised material (DOM=2, 5). Hanging wall granite waste material
was unfolded to the base of the granite. Footwall waste units were unfolded to the
top of the quartzite.

14.13.7 Estimation Results

To ensure completeness of the block estimates and avoid potential issues for
missing grades during mine planning the third and final search passes used large
search radii to ensure most relevant blocks were assigned estimated grades. This
ensured that nearly all mineralised blocks were assigned estimates.

Table 14.20 displays the degree of completeness of the three pass estimation. Most
blocks within the mineralised domains are estimated. Unestimated blocks comprise
mostly extrapolated waste areas. Default values were assigned to all unestimated
blocks, as either 0.001 for all economic elements Cu, Mo, Au, Ag, Co, Re, Zn, Pb or
assigned values defined in Table 14.19 by domain, region and weathering type for
Fe, S and Density.




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Table 14.19      Default Values for Unestimated Blocks
               Description      Region   Dom ain   MINLweathering    Fe        S     Density
                                            1             2          1.72     0.06    2.47
                                            1             3          1.66     0.06    2.50
                                            1             4          1.61     0.07    2.54
                                            2             2          2.25     0.06    2.48
                                            2             3          2.19     0.09    2.60
                                            2             4          2.69     0.29    2.67
                  Waste             0
                                            5             2          2.67     0.10    2.58
                                            5             3          3.04     0.14    2.51
                                            5             4          3.29     0.30    2.59
                                            9             2          2.73     0.07    2.56
                                            9             3          2.57     0.04    2.56
                                            9             4          3.19     0.09    2.59
                                            3             2          3.06     0.99    2.56
                                            3             3          2.85     0.82    2.57
               Mount Dore
                                            3             4          3.45     1.64    2.67
                 North              10
                 Upper                      4             2          2.96     0.09    2.30
                                            4             3          2.52     0.45    2.56
                                            4             4          2.77     0.94    2.65
                                            6             3          2.36     1.70    2.57
                                            6             4          2.51     1.92    2.63
             Mount Dore North               7             3          3.12     0.63    2.55
                  Low er
                                    20      7             4          3.65     1.76    2.67
                    &
                  Merlin                    8             2          2.37     0.01    2.64
                                            8             3          2.94     0.21    2.48
                                            8             4          3.61     0.91    2.67
                                            7             2          3.72     0.07    2.52
                                            7             3          2.98     0.17    2.46
               Mount Dore                   7             4          3.16     1.21    2.62
                                    30
                 South                      8             2          3.64     0.04    2.57
                                            8             3          2.84     0.19    2.53
                                            8             4          3.01     0.63    2.55

Table 14.20      Proportion of Blocks Estimated
                             Dom ain Grades from Cu to Mo Fe   S    Density
                                1           59%          18% 29%     29%
                                2           93%          93% 88%     88%
                                3          100%         100% 100% 100%
                                4           99%          99% 98%     98%
                                5           86%          71% 70%     70%
                                6          100%         100% 100% 100%
                                7          100%         100% 100% 100%
                                8          100%         100% 98%     98%
                                9           58%          45% 29%     29%




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Figure 14.22 to Figure 14.31 demonstrate the block grade estimates for each
element or variable for the same section as the domains displayed in Figure 14.14 to
Figure 14.21.

  Figure 14.22 Section 7,605,450 mN:              Figure 14.24 Section 7,605,450 mN: Zn
                 Cu




  Figure 14.23 Section 7,605,450 mN:                Figure 14.25 Section 7,605,450 mN:
                 Mo                                                Re




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  Figure 14.26 Section 7,605,450 mN:              Figure 14.29 Section 7,605,450 mN: Fe
                 Ag




                                                    Figure 14.30 Section 7,605,450 mN:
  Figure 14.27 Section 7,605,450 mN:
                                                                   Co
                 Au




  Figure 14.28 Section 7,605,450 mN:               Figure 14.31 Section 7,605,450 mN: S
                 Pb




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14.14      Model Validation

Validation of the block model included:
•        Visual inspection of the grade estimates.
•        Global mean and variance comparisons.
•        Review of estimation quality parameters such as number of drill holes, number
         of samples, slope of regression, kriging variance, and average distance of
         samples.
•        SWATH plots. These include comparison of block model and declustered
         composite grade averages for East-West, North-South and Vertical slices.
•        Discrete Gaussian comparisons. This compares the block model distribution
         (grade - volume) to the declustered composite distribution after theoretical
         Gaussian adjustment for the change of support from 2 m composites to the
         two principal block sizes used for estimation.

A comparison of the global mean and variance between the declustered composites
and the volume weighted block model estimates for each combined domain is
provided for Mo and Cu in Table 14.21. The mean grades compare favourably and
indicate no significant bias.

Table 14.21       Global Mean and Variance Comparison
                                                                        Actual        Theoretical
                       Cut Composites           Block Estimates
Element Dom ain                                                        F Factor        F Factor
                   Number   Mean     Var     Number   Mean     Var      Actual    5x12.5x5   10x25x10
              1     15558   0.004   0.0002   83848    0.004   0.000     0.18
              2     7405    0.035   0.0019   130378   0.034   0.000     0.20
              3     4183    0.403   0.2328   107428   0.395   0.037     0.16                   0.25
              4     4731    0.098   0.0163   100344   0.100   0.002     0.11                   0.25
    Cu        5     9279    0.045   0.0031   253135   0.043   0.001     0.19
              6      694    0.328   0.2349   28830    0.338   0.088     0.38       0.45
              7     8291    0.548   0.5401   171526   0.529   0.095     0.18       0.45
              8     3335    0.099   0.0123   146992   0.099   0.002     0.15       0.45
              9     3955    0.019   0.0018   111802   0.015   0.000     0.16
              1     15558   0.001   0.0000   83848    0.001   0.0000
              2     7405    0.002   0.0000   130378   0.002   0.0000    0.14
              3     4183    0.005   0.0004   107428   0.005   0.0000    0.10                   0.32
              4     4731    0.002   0.0001   100344   0.002   0.0000    0.06
    Mo        5     9279    0.002   0.0000   253135   0.002   0.0000    0.13
              6      694    1.299   3.5391   28830    1.322   0.9160    0.26       0.45
              7     8291    0.005   0.0005   171526   0.006   0.0001    0.21       0.41
              8     3335    0.019   0.0028   146992   0.017   0.0007    0.26       0.41
              9     3955    0.001   0.0000   111802   0.001   0.0000    0.50




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The change in variance from composite to block support can be calculated as the
F Factor. This can be compared to the theoretical F Factor calculated from the
variogram model for the desired selective mining unit, or in this case the parent cell
size used for estimation (10 x 25 x 10 m for open pit areas and 5 x 6.25 x 5 m for
underground target areas). Since theoretical selectivity is rarely achieved in practise
due to the information effect and mining errors the theoretical F factors presented in
Table 14.21 are reduced by 20% based on previous experience. The comparison of
the F Factors in Table 14.21 provides some indication as to the ability of the block
model estimate to represent the desired parent blocks that are estimated. The
following can be concluded as far as the grade estimates within each domain:
•     Open pit estimates for Cu display greater smoothing than desired for the
      selective 10 m widths estimated. The parent blocks used for estimation are
      smaller than can effectively be supported by the 50 m spaced drilling resulting
      in the smoothing. These grade estimates effectively reflect a larger selectivity.
•     Underground estimates generally display greater degree of smoothing as the
      target block size is smaller again. For the narrow Merlin zone Cu performs well
      and Mo reasonably well given the greater grade range. For the wider Cu and
      polymetallic underground zones this is reversed with Cu performing poorly and
      Mo performing better. Overall the results indicate significant smoothing of the
      grade estimates for the block sizes selected.

For both target mining sizes the F Factor comparison does not take into account the
hard domain boundaries and searching geometry that have driven the selection of
block sizes used for estimation. The comparison clearly indicates that the drilling
data is not sufficient to effectively estimate the parent block sizes used. As such the
parent block sizes do not reflect the effective selectivity of the resource estimate
which equate to a larger block size. However there is some degree of selectively built
into the hard domain boundaries used in the estimate. The smaller parent block size
used for underground estimation displays a higher F Factor and hence less
smoothing than would have occurred if a larger block size were estimated and that
achieved in the open pit estimates. As a result the resource model will have a greater
chance of reflecting some of the local grade variation within the lower grades areas
surrounding the high grade veins at Merlin, particularly near the hard domain
boundaries, and which could improve the mine stope optimisation and design.

The theoretical F Factors can also be used to adjust the composite grade population
to reflect the block size being estimated. This uses a discrete Gaussian adjustment.
The block model estimates can then be compared to the scaled discrete Gaussian
composite distribution to determine if bias or smoothing is significant across the
distribution. Figure 14.32 to Figure 14.34 presents the discrete Gaussian comparison
for Cu, Zn and Mo broken into three domain groups for the main mineralisation
packages. Results for Cu and Zn are relatively comparably indicating the model
effectively reproduces the range of grades expected. A small degree of smoothing is
indicated, particularly in Mount Dore North which displays greater structural control,
where volume below 0.3% Cu or 0.4% Zn are possibly overestimated slightly while
grades above 0.5% Cu are slightly underestimated.




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For Mo estimates within Merlin there is a higher level of smoothing evident for grades
<1% Mo. These can be attributed to two areas; interburden within multiple high grade
Mo intercepts where smoothing is inevitable and selective mining unlikely to be
achieved, and low grade peripheral areas. Peripheral areas of low grade are typically
thin and there is some small degree of risk than during definition or mining these
areas may incur some loss of reserves. As with any narrow vein mining scenario it is
essential that the mine planning process include a review of the drilling to determine
those areas where smoothing may result in higher risk areas for conversion of
resources to reserves.

Figure 14.32 Discrete Gaussian Comparison for Cu
                               60                                                               3                                         40                                                                            3
                                                         Mt Dore North Upper                                                                                   Mt Dore North Lower (+Merlin)
                                                                                                                                          35
                               50
                                                                                                                                          30
                                                                                                                                                                                                                        2




                                                                                                                                                                                                                            Average Cu Grade
                               40                                                               2
                                                                                                    Average Cu Grade
                                                                                                                                          25

                                                                                                                           Volume (Mm3)
                Volume (Mm3)




                               30                                                                                                         20

                                                                                                                                          15
                               20                                                               1                                                                                                                       1
                                                                                                                                          10
                               10
                                                                                                                                           5

                                0                                                               0                                          0                                                                            0
                                    0    0.25    0.5   0.75    1      1.25   1.5   1.75     2                                                  0       0.25    0.5     0.75      1       1.25     1.5    1.75      2
                                                        Cu Cut-off (%)                                                                                                   Cu Cut-off (%)

                                        Mm3_OK         Mm3_DG             Cu_OK           Cu_DG                                                    Mm3_OK              Mm3_DG               Cu_OK               Cu_DG

                               35                                                               3                                         120                                                                           3
                                                         Mt Dore South                                                                                                        Combined
                               30
                                                                                                                                          100

                               25
                                                                                                    Average Cu Grade




                                                                                                                                                                                                                            Average Cu Grade
                                                                                                2                                          80                                                                           2
               Volume (Mm3)




                                                                                                                           Volume (Mm3)




                               20
                                                                                                                                           60
                               15
                                                                                                1                                          40                                                                           1
                               10

                                5                                                                                                          20


                                0                                                               0                                              0                                                                        0
                                    0    0.25    0.5   0.75     1     1.25   1.5   1.75     2                                                      0    0.25     0.5    0.75         1    1.25     1.5   1.75      2
                                                         Cu Cut-off (%)                                                                                                  Cu Cut-off (%)

                                        Mm3_OK         Mm3_DG             Cu_OK           Cu_DG                                                        Mm3_OK            Mm3_DG                  Cu_OK           Cu_DG




Figure 14.33 Discrete Gaussian Comparison for Zn
                               60                                                               3                                         40                                                                            3
                                                         Mt Dore North Upper                                                                                   Mt Dore North Lower (+Merlin)
                                                                                                                                          35
                               50
                                                                                                                                          30
                               40                                                               2                                                                                                                       2
                                                                                                                                                                                                                            Average Fe grade (%)
                                                                                                    Average Fe grade (%)




                                                                                                                                          25
                                                                                                                           Volume (Mm3)
                Volume (Mm3)




                               30                                                                                                         20

                                                                                                                                          15
                               20                                                               1                                                                                                                       1
                                                                                                                                          10
                               10
                                                                                                                                           5

                                0                                                               0                                          0                                                                            0
                                    0    0.25    0.5   0.75    1      1.25   1.5   1.75     2                                                  0       0.25    0.5     0.75      1       1.25     1.5    1.75      2
                                                        Zn Cut-off (%)                                                                                                   Zn Cut-off (%)

                                        Mm3_OK         Mm3_DG             Zn_OK           Zn_DG                                                    Mm3_OK              Mm3_DG               Zn_OK               Zn_DG

                               35                                                               3                                         120                                                                           3
                                                         Mt Dore South                                                                                                        Combined
                               30
                                                                                                                                          100

                               25
                                                                                                2
                                                                                                    Average Fe grade (%)




                                                                                                                                                                                                                            Average Fe grade (%)




                                                                                                                                           80                                                                           2
               Volume (Mm3)




                                                                                                                           Volume (Mm3)




                               20
                                                                                                                                           60
                               15
                                                                                                1                                          40                                                                           1
                               10

                                5                                                                                                          20


                                0                                                               0                                              0                                                                        0
                                    0    0.25    0.5   0.75     1     1.25   1.5   1.75     2                                                      0    0.25     0.5    0.75         1    1.25     1.5   1.75      2
                                                         Zn Cut-off (%)                                                                                                  Zn Cut-off (%)

                                        Mm3_OK         Mm3_DG             Zn_OK           Zn_DG                                                        Mm3_OK            Mm3_DG                  Zn_OK           Zn_DG




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Figure 14.34 Discrete Gaussian Comparison for Mo
                              3.0                                                                 4                                     1.0                                                                         1
                                                          DOM 6 - Merlin                                                                                             DOM 7 Lower Copper
                                                                                                                                        0.9
                              2.5
                                                                                                                                        0.8
                                                                                                  3
                                                                                                                                        0.7




                                                                                                                                                                                                                         Average Mo Grade
                              2.0




                                                                                                      Average Mo Grade
                                                                                                                                        0.6




                                                                                                                         Volume (Mm3)
                Volume (Mm3)



                              1.5                                                                 2                                     0.5
                                                                                                                                        0.4
                              1.0
                                                                                                                                        0.3
                                                                                                  1
                                                                                                                                        0.2
                              0.5
                                                                                                                                        0.1
                              0.0                                                                 0                                     0.0                                                                         0
                                    0    0.25    0.5    0.75     1     1.25    1.5   1.75     2                                               0       0.25    0.5    0.75      1       1.25    1.5    1.75      2
                                                         Mo Cut-off (%)                                                                                                Mo Cut-off (%)

                                        Mm3_OK          Mm3_DG              Mo_OK           Mo_DG                                                 Mm3_OK             Mm3_DG              Mo_OK               Mo_DG

                              1.0                                                                 1                                       30                                                                         4
                                                       DOM 8 Lower Polymetallic                                                                                             Combined
                              0.9
                                                                                                                                          25
                              0.8
                                                                                                                                                                                                                     3
                              0.7




                                                                                                      Average Mo Grade




                                                                                                                                                                                                                         Average Mo Grade
                                                                                                                                          20
                              0.6
               Volume (Mm3)




                                                                                                                         Volume (Mm3)
                              0.5                                                                                                         15                                                                         2
                              0.4
                                                                                                                                          10
                              0.3
                                                                                                                                                                                                                     1
                              0.2
                                                                                                                                              5
                              0.1
                              0.0                                                                 0                                           0                                                                      0
                                    0    0.25    0.5    0.75     1     1.25    1.5   1.75     2                                                   0    0.25    0.5    0.75         1    1.25    1.5   1.75      2
                                                           Mo Cut-off (%)                                                                                              Mo Cut-off (%)

                                        Mm3_OK          Mm3_DG              Mo_OK           Mo_DG                                                     Mm3_OK          Mm3_DG                  Mo_OK           Mo_DG




14.15      Mineral Resource Classification

The Mount Dore and Merlin resources were previously classified by Mike Job of QG
(2009, 2010), largely on the basis that:
•       The target drill spacing of 50 x 50 m was sufficient to define Indicated
        Mineral Resource under both JORC and NI 43-101 guidelines.
•       The resource interpretations were sufficiently robust.
•       Data quality and quantity are sufficient for Indicated classification.
•       Cu, Mo and Re display the level of continuity required.

Exploration work completed since the last resource estimate has largely
concentrated on the extension of the 50 x 50 m infill drilling pattern at Mount Dore
North. This work along with the subsequent resource analysis and reviews do not
present any reason to change the classification criteria. As a result the basic premise
that 50 m spaced grid drilling is sufficient to define Indicated Resources has been
retained and applied by defining the plan area where 50 x 50 m drilling intersects the
Merlin or lower copper mineralisation, allowing for 25 m of extrapolation. Figure 9.1
presents the plan outline of the Indicated Mineral Resource, which does not directly
correlate the collar locations to the point of intersection with the mineralisation. The
classification was then modified to account for areas of lower continuity including:
•       Hanging wall and footwall lithology units for granite, quartzite and Stavely
        Formation, outside any mineralisation shapes was reset to unclassified.
•       Waste zones outside the mineralisation shapes and within the Kuridala
        Formation were reset to Inferred Mineral Resource given that any
        mineralisation does not display the continuity required for a higher
        classification.




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•       A few blocks that lie above the depletion zone modelled at Mount Dore North
        were reset to Inferred Mineral Resource.

Figure 14.21 demonstrates the application of the classification process with
Unclassified (orange), Inferred (red) and Indicated (blue) classification viewed in
section.

There are a few small areas of more detailed drilling that are not considered suitable
for Measured Mineral Resource classification, including:
•       12.5 m spaced drilling at Little Wizard defines a very small high grade zone.
        The limited extent and the variation in thickness and grade raise some
        concerns that Measured Mineral Resource classification may be optimistic.
•       2 infill lines of staggered 50 m drilling on 25 m offsets are available at Mount
        Dore North. These close down the drill spacing to about 35 m spacing. This is
        not considered sufficient for Measured Mineral Resource classification.
•       An area of close spaced RC drilling at Mount Dore South provides sufficient
        drill spacing, however the previous RC has some untested concerns about
        sample quality that require further investigation before Measured Mineral
        Resource should be considered.

14.16     Mineral Resource Statement

The total Mineral Resource estimated for Mount Dore - Merlin deposits is presented
in two exclusive parts at different cut-off grades.

The Merlin and Little Wizard total Mineral Resource estimate at a 0.3% Mo cut-off is:
                 6.5 Mt at 1.3% Mo and 23 ppm Re Indicated Mineral Resource
                 0.2 Mt at 0.9% Mo and 15 ppm Re Inferred Mineral Resource

The Mount Dore total Mineral Resource estimate at a 0.25% Cu cut-off is:
                 86 Mt at 0.55% Cu Indicated Mineral Resource
                 58 Mt at 0.47% Cu Inferred Mineral Resource

Note this excludes all Cu mineralisation included in the above Merlin + Little Wizard
Mineral Resource reported above a 0.3% Mo cut-off.

Significant additional Zn mineralisation exists outside of the quoted Cu mineralisation
at Mount Dore North. Viable recovery of Zn is not yet demonstrated and stand alone
Zn resources are not included in this resource statement.

Previous resource statements have included some small updates at an upper
bonanza Mo zone at Little Wizard, but otherwise the data used for the previous
greater Mount Dore resource includes drilling completed to 27th July 2010. This
update sees the addition of 93 drill holes for 26,253 m for a total of 469 assayed
resource drill holes for 113,928 m. Drilling over the last twelve months has largely
targeted:




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•         Defining the resource limits to the Merlin mineralisation.
•         Regular 50 x 50 m infill drilling of the Mount Dore North area targeting the
          definition of Merlin.
•         Additional close spaced drilling at Little Wizard.
Table 14.22          Mount Dore - Merlin Mineral Resource Breakdown
                                                      Mo     Re     Cu     Zn   Pb    Ag  Au  Co
Cut-off     Region         Classification       Mt                                                     Density
                                                      %     ppm     %      %    %    ppm ppm ppm
            Lt Wizard      Indicated           15 kt 6.49 83.9 2.29 0.00 0.01 25.0         0.63   21    2.38
                           Indicated            6.5   1.34 23.3 0.33 0.14 0.02       8.3   0.08   81    2.62
Mo %        Merlin
                           Inferred             0.2   0.85 15.1 0.44 0.24 0.02       8.2   0.13   91    2.67
≥0.3
                           Indicated            6.5   1.35 23.4 0.33 0.14 0.02       8.3   0.08   81    2.62
            Merlin Total
                           Inferred             0.2   0.85 15.1 0.44 0.24 0.02       8.2   0.13   91    2.67

Note: The effective date for this mineral resource estimate is 27 June 2010.

The resource was estimated independently by Golder. Technical items of interest
include:
Tenure:
•         All resources are held by Ivanhoe Cloncurry Mines Pty Ltd, a wholly owned
          subsidiary of IVA, giving IVA 100% ownership.
•         The majority of the resource falls within 4 granted mining leases which cover
          the extraction of copper and molybdenum and some other minor elements.
•         Two of the mining leases have expired but have pending renewal applications.
          IVA has no reason to believe that these leases will not be renewed.
•         IVA has applied for rhenium to be added to the mining leases and understand
          this should only be a formality. Otherwise rhenium falls with the encompassing
          EPM held by IVA.

Exploration:
•         The majority of Mount Dore North mineralisation is drilled by modern diamond
          drilling completed by IVA, where data has been collected and collated digitally
          using an integrated commercial data management software, namely acQuire.
•         Mount Dore South is drilled by a mix of RC and diamond drilling by IVA up till
          2008 and several other previous companies over several periods of
          exploration from 1976 through to 2004 when IVA acquired the property.
•         IVA sampling is on regular 2 m intervals with recent changes allowing some
          sub-sampling at the high grade Merlin mineralisation.
•         Density samples are generally undertaken every 10 m using an industry
          standard wax weight method. 10,390 measurements indicate a low variation in
          density values across most lithology, weathering and mineralisation types.




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Quality Control:
•     An Ivanhoe Mines expert from Canada has reviewed the IVA exploration
      procedures and QAQC on a six monthly basis.
•     QAQC assaying by IVA is completed at a high frequency and includes 23% of
      all assays.
•     IVA has completed two gyroscopic down hole survey programs to locate older
      suspect drill holes and some deep holes at Merlin (down to 836 m).
•     IVA has resurveyed the majority of the older drilling and the mining tenements.
•     Independent reviews have been completed by QG and Golder.

Resource Estimation:
•     Polymetallic domaining and interpretations are defined by in the following
      descending order of precedence.
      -      Merlin domains with >0.3% Mo.
      -      Copper mineralisation with >0.25% Cu.
      -      Remaining polymetallic zones (from Zn, low grade Mo or other
             elements) where Cu equivalent >0.25%. Note Cu equivalence is based
             on values presented in a prior scoping study on Merlin and is only used
             to aid interpretation and is not used for reporting.
•     Grades are estimated from cut 2 m composited assays using parent block
      ordinary kriging. Elements estimated include Cu, Co, Mo, Re, Zn, Pb, Au, Ag,
      Fe and S as well as dry bulk density.
•     Unfolding to simple surfaces was used to improve search and sample
      selection and account for the variation in dip which varies from 30° to 80°.
•     Open pit potential parent block size for mineralisation is 10 x 25 x 10 m.
•     Underground block potential parent block size for mineralisation is 5 x 12.5 x
      5 m.
•     Indicated Mineral Resource classification is based on the completion of 50 m
      grid drilling, see Section 14.15.
•     Interpretation used for the resource mineralisation limits most extrapolation to
      a maximum of 25 m.

Resource Issues:
•     At Mount Dore South the resource drilling quality raises some known concerns
      with respect to the quality of the older assaying, selective sampling of older
      drilling and some suspect wet RC sampling processes. These issues have not
      been followed up due to higher priority to define Merlin and remain
      outstanding. As a result close spaced drilling at Mount Dore South is not
      classified as Measured.
•     The Merlin high grade Mo zone is on average 3.9 m in width and defined
      largely on a 50 m spaced square grid. Overall the Merlin vein(s) display good




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      continuity. However, there are no intensively drilled areas at Merlin to
      demonstrate the short range continuity of the vein i.e. if the vein is dislocated
      or pinches and swells. It is unlikely that this can feasibly be done until
      underground development allows the next phase of infill drilling from a more
      direct angle and in a more cost effective manner.
•     Mount Dore is polymetallic with many potentially economic elements present.
      Some of these will not be recoverable or have very low recoveries for some
      processing options such as heap leach. Since more work is required to test
      their economic significance, they are not part of the resource statement which
      includes only Cu, Mo and Re where metallurgical studies have indicated
      potential for reasonable recovery.
•     Open pit resource estimates are based on the assumption of a bulk mining
      scenario where high quality grade control practises are employed.
•     Drill spacing is currently at around 50 m. Though sufficient for Indicated
      Mineral Resource classification the area is structurally complex and there is
      still room for the thin high grade Merlin zone to be offset and folded locally.
      The block model size has been driven by geometry issues and the blocks
      estimated are smaller than can be adequately informed for the general drill
      spacing. Essentially the block estimates include significant smoothing and
      relate to a larger selectivity than the block sizes estimated.
•     No specific close spaced drilling has been completed to date to assist the
      definition of the grade variability and demonstrate continuity of the Merlin vein
      at short range, with the exception of the near surface Little Wizard bonanza
      zone that may not be indicative of the greater Merlin deposit. Such information
      will be required at the next stage of work and may be more practical to derive
      from underground drilling programs.
•     Significant additional Zn mineralisation exists outside of the quoted Cu
      mineralisation at Mount Dore North. Viable recovery of Zn is not yet
      demonstrated and stand alone Zn resources are not included in this resource
      statement.

Development
•     The Prefeasibility Study has been completed with this report documenting the
      resource used as a basis for the current study.
•     IVA has recently completed the purchase of the Osborne mine and mill
      facilities 50 km for the South, see ASX announcement dated 4 October 2010.
      Current development proposals include the transport of ore to Osborne for
      milling and processing. This reduces the required capital cost and permitting
      requirements for establishing a mill at Mount Dore.
•     IVA awarded the contract for underground development of the Merlin deposit
      by means of a decline targeted initially at exploration but eventually for mining.

This Mineral Resource estimate is based upon and accurately reflects data compiled
or supervised by Mr John Horton, Principal Geologist, who is a Member of the
Australasian Institute of Mining and Metallurgy and a full time employee of Golder
Associates Pty Ltd. Mr Horton has sufficient experience that is relevant to the style of




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mineralisation and the type of deposit under consideration and to the activity which
he has undertaken to qualify as a Competent Person as defined in the 2004 edition
of the ‘Australasian Code for the Reporting of Exploration Results, Mineral
Resources and Ore Reserves’ or as a Qualified Person under NI43-101.

Table 14.23 present the Indicated Mineral Resource, broken down by weathering
domains. This indicates:
•       There is a higher proportion of oxide and transitional copper resource at
        Mount Dore South.
•       That Merlin lies below the oxide zone predominantly in fresh rock.
•       The upper metallurgical domain displays some residual copper content based
        on sequential copper analyses and comprises a minor though significant
        proportion of the Mount Dore South mineralisation which may require further
        metallurgical assessment.

Table 14.23      Merlin Mineral Resource Weathering Type
                      Description        Oxide     Trans      Fresh     Oxide   Trans   Fresh
Classification    MINL Dom ain Code        2         3            4      2       3       4
                  Cut-off    Region        Mt        Mt           Mt     %       %       %
    Indicated    Mo >0.3      Merlin                 0.7          5.9    0%     10%     90%
    Inferred     Mo >0.3      Merlin                 0.0          0.2    0%      0%     100%




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15    MINERAL RESERVE ESTIMATES

The Mineral Reserve estimates were prepared after selection of the mining methods
as described in Section 16. The following factors are relevant to the selection of a
minimum mining unit (MMU):
•     Ore body geometry including continuity, width and dip.
•     Anticipated ground conditions during mining.
•     Mining method and equipment selection.

For The Project, the ore body geometry is known to be relatively continuous, with
narrow to moderate widths, and shallow to moderate dip. For the two main mining
methods selected (ie DAF and LHOS methods) the following MMUs apply:
•     DAF mining using mechanised mining with 2-boom jumbos requires a profile of
      4.5 m wide x 5.0 m high, with a variable length of drift. The DAF method is able
      to adapt reasonably closely to a variable orebody geometry and the profile is
      largely unaffected by the anticipated ground conditions.
•     LHOS mining using mechanised mining with uphole drilling from 15 m spaced
      levels and applying a minimum stope excavation width of 3.5 m, before
      unplanned over break. Modelling of narrower widths for the stoping was shown
      to be inconsequential, based on the current resource interpretation. For the
      shallow to moderate dip and variable ground conditions, it is expected to be
      difficult to achieve narrower excavations in practice.

The Datamine Mineable Shape Optimiser (MSO) process was used to apply
mineable shapes to the resource model. The shapes include 0.5 m of dilution on the
footwall and 1.5 m of dilution on the hangingwall of the orebody. The mining
inventory was prepared using the mineralised inventories as follows:
•     Development inventories from the development design wireframes were
      evaluated against the resource block model, with a 98% recovery of the
      Mineral Resource applied.
•     DAF inventories from the MSO inventories inclusive of dilution, with a 97%
      recovery of the Mineral Resource applied.
•     LHOS inventories from the MSO inventories inclusive of dilution, net of
      development, with a 95% recovery of the Mineral Resource applied.

After rounding the results to appropriate significant figures, the Mineral Reserve
estimate is set out in Table 15.1.




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Table 15.1        Mineral Reserve Estimate showing the Split by Mining Method

          Mining Method                   Mt          Mo t         Re kg       Mo %       Re ppm

          Total Proved                           -           -             -          -            -
          DAF - LW, A, B                       2.0     29,000        47,000    1.4           23.3
          DAF - C & D                          0.3      3,000         4,000    1.0           13.6
          DAF - E, F, G & H                    1.8     17,000        36,000    1.0           20.4
          Ore Drive                            0.5      6,000         9,000    1.1           17.2
          Transverse LHOS                      0.2      1,000         1,000    0.8            8.9
          Longitudinal LHOS                    1.9     19,000        31,000    1.0           15.9
          Total Probable                       6.7     75,000      128,000     1.1           19.1
        The effective date for this Mineral Reserve is 10 October 2011.

Significant risks are addressed in Section 25.5 Risks.




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16     MINING METHODS

16.1      Geotechnical

Following an initial high level review and site visit AMC conducted further work to
provide parameters for use in preliminary mine design. (AMC 110031
September 2010).

The work was primarily based on information in the geotechnical drill hole database
and used in the construction of the May 2010 resource model. The data related the
four high grade molybdenum orebodies, being 2A, 2B, 2C & 2D. Information in the
database was reviewed and basic checks made to gain an understanding of the level
and quality of data prior to inclusion in geotechnical analyses.

Each of the logged intervals was assigned parameters to characterize the rock mass.
The rock mass characterization and subsequently defined geotechnical domains
considered the following:
•      Firstly, consideration was given to the rock mass conditions that impinge on
       stoping. For example the granite will not impinge upon the hangingwall of the
       stoping areas and therefore was not considered in the geotechnical domains.
•      Within the rock mass which might impinge on stoping there were only minor
       variations in lithology. Lithology was not considered in subsequent analyses.
•      Physical characteristics such as weathering and spatial variations were also
       considered as part of the geotechnical domains. Lens 2A was spatially
       extensive and was divided into zones, however, Lens 2B, 2C and 2D were
       restricted spatially and didn’t warrant dividing.

The resulting geotechnical domains were hangingwall, resource and footwall units for
the different lenses and where appropriate, these were sub-divided by weathering
and spatial location such as north and south zones.

In general the weathering and fracturing have resulted in poor rock mass conditions
in the hangingwall and resource units. A trend can be seen of improving ground
conditions with an increase in depth and distance from the hangingwall contact of
Lens 2A.

Stope stability was assessed using the drilling data in each domain and the
Matthews Stability Graph method. This resulted in basic stope dimensions for each
lens to be used in on-going mine design work.

In the weathered zone very restricted stope dimensions and extensive support are
required to achieve stable spans. Drift and fill is achievable and more suited to the
weathered rock mass conditions. Drift and fill may also be required in the parts or all
of the unweathered zones because the resource rock mass conditions are poor.

The drift and fill dimensions are based on a 4 to 5m drift height interval are shown in
Table 16.1.




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Table 16.1                     Recommended Drift and Fill Dimensions
                                                         Hangingwall         Sidewall       Backs
                                   Lens                     Strike            Strike        Width        Face


                           Lens 2A & 2D - South
    Weathered




                                                             Unlimited #     Unlimited #     5m          N/A
                              Lens 2A - North
    Unweathered




                             Lens 2A - South

                                                             Unlimited #     Unlimited #     6m          N/A

                              Lens 2A - North


#                  Assumes the installation of development ground support.

The stope dimensions in Table 16.2 are based on a 15 m sublevel interval with an
underhand or overhand sequence for the unweathered rock mass.

Table 16.2                     Recommended Stope Dimensions
                                                             Hangingwall     Footwall                     End Wall
                                    Lens                                                   Crown Width
                                                                Strike        Strike                       Width
                              Lens 2A - South                   25m*            25m           15m*             6m
     Unweathered




                               Lens 2A - North                  30m*            30m           15m*             7m

                                   Lens 2B                       25m            25m           20m*          10m
                                  Lens 2C                        25m            25m           25m*          22m
*                  Assumes the installation of cablebolts.

The strike length of the stopes increases with a steeper resource dip in the northern
zone. The end walls in the resource are currently restricting stope dimensions.
Longitudinal stoping could be achieved in zones where the resource is less than 6 to
7 m wide. If a vertical unsupported footwall is required in the resource unit, strike
length will be restricted to 10 m.

The Lens 2B and 2C dimensions allow for transverse and longitudinal stoping of a
greater dimension than currently used in the mine design.

If an underhand sequence is adopted, the crown values will be replaced with
dimensions based on pastefill exposure and strength requirements. The Project is
based on a zone of molybdenum-rhenium mineralisation that occurs as molybdenite,
hosted in an east-dipping shear and breccia zone, with extremely high-grade portions
occurring as a molybdenite-supported breccia. Given the high value of the ore
material, it was considered likely, at an early stage, that maximising ore recovery was
to be a key attribute of the successful mining system.

The high grade molybdenum orebody (refer to Figure 16.1) comprises of east dipping
tabular lenses dipping from 30° to 60°. The main lens (5) is a large, moderate dipping
and generally narrow lens, with two minor footwall lenses (6 and 7) that are steeper




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dipping and thicker, and a minor Little Wizard (LW) lens (4) that is a very high grade,
up-plunge pod.

The host rock sequence is highly variable and is weathered in the upper levels, and
contains rocks with very poor to fair rock strengths. The footwall rock sequence is
mostly competent. The mine access including decline system and ventilation system
is planned to be constructed mostly within the competent footwall rock sequence.

Figure 16.1      High Grade Molybdenum Wireframe (Longitudinal Section -
                 Aug ’10 Model)



       LW minor lens (4)                              100 mbs

                                                                  Main lens (5)




                                                                                  400m




       Footwall
       minor lenses
       (6 and 7)


                                               700m




16.2      Stoping Methods

The mining methods considered included:
•      Open pit.
•      Caving methods.
•      Unsupported methods.
•      Supported methods.

An open pit for all or part of The Project is considered possible. Evaluation showed
that while technically and financially possible the pit option has a very high stripping
ratio of 83:1, involves a total material movement of 429 Mt, and can only sustain the
targeted mill feed rate with use of substantial ROM stockpiles in some years.




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The caving methods of mining are considered unsuitable due to the orebody being of
insufficient size to consider the block caving method, and the unacceptable risk of
excessive dilution due to the poor ground conditions.

The unsupported methods are considered unsuitable due to the generally poor
ground conditions and the project requirement for high recovery of the high value
ore.

The supported methods of mining identified were:
•     Mechanised development stoping - includes mechanised cut and fill stoping
      and drift and fill stoping, (includes overhand and underhand sequences). Both
      are considered suitable for a variety of orebody conditions and geometry.
•     Blast hole bench stoping - includes benching and Avoca stoping overhand
      methods. Benching is suitable for moderate to steep dipping, narrow orebody
      sections with widths of less than 10 m.
•     Blast hole open stoping - includes Longhole Open Stoping (LHOS), Sublevel
      Open Stoping (SLOS), Vertical Crater Retreat (VCR), Alimak, Raise & Slashing
      variations, up and down holes, overhand and underhand methods. Generally
      not suited to poor ground conditions. SLOS generally not suitable for shallow
      dipping, narrow orebodies.

Given the generally very poor ground conditions within the host rock sequence, it
was resolved that only mining methods comprising small openings and using
consolidated backfill were appropriate for further consideration.

Therefore, the appropriate methods were determined to include:
•     Drift and Fill mining (DAF).
•     Small-scale long hole open stoping (LHOS) with cemented fill.

DAF (refer to Figure 16.2) was chosen as the preferred method for mining in very
poor ground and in flatter sections of the high grade orebody. Small-scale LHOS was
chosen as the preferred method of mining in fair to good ground at moderate to steep
dips. The LHOS may be longitudinal in narrow sections and change to transverse
orientation in wide sections of the orebody. An underhand mining sequence (from
top-down) was chosen as the default sequence.




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Figure 16.2      High Underhand DAF Mining in a Narrow Orebody




A Strategy Optimisation (SO) study using comparative assessment techniques was
performed to determine suitable mining parameters, such as Cut-off Value (COV),
mining system selection and production rate. The SO optimised the value of The
Project, taking into account the constraints imposed by key information for resource,
geotechnical conditions, mining productivities, metallurgical performance, cost and
financial assumptions. Based on the assumptions and criteria applied, underhand
small-scale LHOS mining was shown to provide the highest value if this mining
method could be implemented on a mine wide basis, with small scale LHOS used
where geotechnical conditions permit.

16.3      Mine Design

The underground mine was designed to suit modern trackless mining, using the DAF
and LHOS mining methods. The PFS mine design was prepared for the purpose of
enabling further decisions to be made and included optionality in the design, as
follows:
•      Case 1: DAF mining method for all lenses and all mining blocks.
•      Case 2: DAF mining in main lens, with predominantly transverse LHOS mining
       in footwall lenses in mining blocks C & D.
•      Case 3: DAF mining in south main lens only, with transverse LHOS mining in
       footwall lenses in mining blocks C & D, and longitudinal LHOS in north main
       lens in mining blocks E, F, G & H.

Case 2 was selected as the basis for mineral reserve estimation and further work. A
common decline access and ventilation infrastructure development system will suit




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the three mine design cases. Figure 16.3 shows the Case 2 mine design with the
common decline and ventilation infrastructure. The mine was divided into nine mining
blocks (Little Wizard plus blocks A-H) for scheduling purposes.

Based on the objective of targeting the maximum recovery of economic ore, a break
even cut-off was applied in preparing the PFS mine designs and schedules. The
application of an optimal cut-off should be reviewed following further detailed studies.

The cut-offs applied to the PFS mine designs and schedules +were:
•      DAF ore mining, ore handling and ore processing at A$ 200 /t.
•      LHOS ore mining, ore handling and ore processing at A$ 150 /t.

Figure 16.3      Case 2 Mine Design Layout




16.4      Mining Schedule

The mining schedule was prepared assuming a top-down mining system is employed
in all mining areas, as follows:
•      DAF mining assuming 40 m/mth average advance (including backfill) rate per
       available ore heading, based on a single level, centre-out sequence.
•      Longitudinal LHOS assuming 420 t/d stoping rate (including backfill, excluding
       development) based on a single level, 2 stope end-retreat sequence.
•      Transverse LHOS assuming 600 t/d stoping rate (including backfill, excluding
       development), based on a single level, 2 stope centre-out sequence.

The mining schedule was prepared using the Earthworks Production Scheduler
(EPS) software with input information prepared in Mine2-4D and Excel.




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The mine designs were evaluated in EPS on 15 m vertical increments to align with
the sublevel interval. All scheduling was based on activities accumulated within each
15 m increment, based the north and south mining areas.

The following assumptions apply to the schedule:
•                    Commencement of the surface decline on 1 October 2010 (completed).
•                    The schedule was prepared on a quarterly (3 monthly) basis.
•                    There are no external constraints applied to the schedule, and all activities are
                     able to start when available to do so.
•                    Where multiple lenses occur on a level, DAF mining is assumed to be able to
                     operate concurrently on adjacent lenses, although in practice this may involve
                     multiple levels.
•                    Decline development is scheduled to occur “as soon as possible” (ASAP).
•                    Production is scheduled to occur ASAP.
•                    Waste and ore drives are scheduled to occur “as late as possible” (ALAP).

The resultant production schedule is shown in Figure 16.4. Ore is sourced from 60%
DAF mining, 8% development, and 32% LHOS. All mining blocks are opened up
quickly and production reaches 150 kt/a in 2012, 300 kt/a in 2013 before reaching full
production of about 500 kt/a in 2014. The peak rate of 500 kt/a or above is
maintained for about ten years and is generally able to be maintained whilst the high
productivity footwall lens (6 & 7) stoping blocks are producing and then falls away
rapidly thereafter.

Figure 16.4                     Annual Production Schedule


                                                         Production schedule
                     700.0                                                                  3.00

                     600.0                                                                  2.50
                     500.0
                                                                                            2.00
    Ore mined (kt)




                     400.0
                                                                                                   % Mo




                                                                                            1.50          Ore mined
                     300.0
                                                                                            1.00          Molybdenum head grade
                     200.0

                     100.0                                                                  0.50

                       0.0                                                                  0.00
                             2012
                                    2014
                                           2016
                                                  2018
                                                         2020
                                                                2022
                                                                       2024
                                                                              2026
                                                                                     2028




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The average estimated numbers of heavy mobile equipment for life of mine are
shown in Table 16.3.

Table 16.3       Average Numbers of Heavy Mobile Equipment
                                                              Average
                            Type
                                                              Number
                            Development jumbo                     4
                            Longhole drill                        2
                            Loader - 14 t capacity                2
                            Loader - 17 t capacity                2
                            Truck - 55 t capacity                 3
                            Explosives chargeup                   2
                            Shotcreter                            2
                            Agitator truck                        3
                            Integrated tool carrier               2
                            Scissor lift                          1
                            Grader                                1
                            Backhoe                               1




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17     RECOVERY METHODS

17.1      Process Plant Design Criteria

The process design criteria are primarily a compilation of data from the results of
bench test work programs, vendor and independent consultants and information
supplied by IVA. Jacobs provided additional engineering design criteria from their
past experience on similar applications. Where test work results were unavailable,
assumed data was used.

The majority of the concentrate treatment area design criteria were sourced from the
experience of independent consultants as only limited test work has been completed
to date. Given that insufficient concentrate sample was available to undertake
extensive roasting test work within the project schedule, the process design criteria
for each of the concentrate treatment areas relies on the experience of those who
have operated these processes.

Jacobs database was referenced where no other data were available. The following
is a summary of the process plant design criteria.

17.1.1    General

Annual ore throughput, operating schedules and plant availabilities were agreed with
IVA and reflect best practice for each of the operating areas. The annual ore
throughput is limited by the mining rate.

17.1.2    Ore Blending

Head grades of the ore reflect the expected average ore grade. As the mine plan
evolves and variability is better understood, the design ore grade and its implications
on the plant design should be reviewed.

Ore mineralogy was extrapolated, using the most recent head grades, from the
mineralogical assessment of the four major lithologies (AMMTEC, 2009).

The run-of-mine (ROM) ore physical properties were based on similar ore types.

17.1.3    Crushing

Metallurgical design parameters for the crushing properties of the ore were taken
from the crushing test work report (AMMTEC, 2009). Samples of each of the four
major lithologies were subjected to a series of tests to determine the unconfined
compression strength (UCS) and Bond abrasion index (Ai), along with the Bond rod
and ball mill work indices. The crushing parameters are a summary of a Bruno
crushing model prepared by Metso (2009).




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17.1.4    Grinding

As with crushing, the grinding metallurgical parameters were taken from the 2009
AMMTEC crushing report. A conservative Bond ball mill index was used for the
design of the mill to ensure this area does not become a process bottleneck.

17.1.5    Flotation

The flotation design criteria are the culmination of flotation test work findings from the
2010 laboratory program undertaken at Metcon (2011, 3 reports). A scale up factor of
2 between the laboratory and plant residence times was used for the study, however
this will be reviewed in the next phase to ensure sufficient residence time is available
to achieve the desired metal recoveries.

A conservative concentrate grade of 30% molybdenum was used for the design. This
ensures sufficient capacity in the flotation cells and maximum volume of concentrate
to be roasted.

17.1.6    Services & Tailings

Thickening test work had not been undertaken at the time of writing the Metcon
(2011) reports. Similar settling characteristics to the existing Osborne material were
assumed, giving an underflow density of 70% as advised by Metago Environmental
Engineers, who are designing the tailings storage facility.

Subsequent to the Metcon report, tailings settling testwork has been undertaken
which indicates that the PFS assumptions are valid and will not require a material
change to the design. The final results will be incorporated into the detailed design of
the plant. Existing Osborne services will be used where possible.

17.1.7    Concentrate Handling

The settling and filtration characteristics of the molybdenum concentrate were based
on testing by Pocock Industrial. Drying properties of the molybdenum concentrate
were assumed based on the experience of EHP Consulting Inc. These assumptions
will be reviewed as further test data becomes available.

It has been assumed that the copper concentrate produced will be directed to the
existing copper concentrate thickener, with this thickener having adequate capacity
to handle the additional small volume of material.

17.1.8    Roaster & Off Gas Handling

The roaster design criteria were provided by EHP Consulting Inc based on their
operating experience. Only limited roasting test work has been undertaken on small
samples.

The design allows for a much higher temperature in the lower hearths of the roaster
to promote maximum rhenium volatilisation and recovery to the off gas. The off gas
handling system has been designed with the aim of maximising rhenium recovery.




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Bearing this in mind, wet scrubbing of the off-gas directly from the roaster was
selected to minimise the recirculation and potential rhenium losses expected from a
dry dust collection system that is more typically seen in this application. The solids
collected in the wet scrubbing system are likely to be fine and hard to settle, hence a
conservative approach has been taken to thickener design.

A wet sulphuric acid production plant has been selected for the duty of SO 2 removal
prior to release of gas to the atmosphere. Without a sulphuric acid plant a lime
scrubber would be required.

17.1.9    Rhenium Recovery

The rhenium recovery design criteria, including solvent extraction, ion exchange,
rhenium evaporation and crystallisation, product drying and packaging, were
generated by EHP Consulting Inc based on their operating experience. No test work
has been completed in this area.

The ion exchange design criteria may be considered conservative as the new
generation resins which are much more selective against molybdenum were not
selected, opting instead for a resin with more history on loading capacity in
production facilities.

17.1.10 Calcine Leach & Molybdenum Recovery

The calcine leach and molybdenum recovery design criteria were provided by EHP
Consulting Inc based on operating experience. Only limited leach test work has been
undertaken on small calcined samples to confirm complete molybdenum dissolution
can be achieved. The final molybdenum trioxide product will be packed in 1 tonne
bags for sale.

17.2      Process Plant Design

Jacobs E & C Australia Pty Ltd (Jacobs) was engaged to determine process plant
requirements, together with capital and operating costs of the process plants to an
accuracy of -20% +30% for a treatment rate of 500,000 t/a of ore using IVA specified
average head grades of 1.2% molybdenum, 0.2% copper and 18 g/t rhenium, where
the molybdenum is present as molybdenite, and copper as chalcopyrite.

The Jacobs process plant engineering study (Report 11402-00-G0712 Rev P1) was
completed in December 2010 and progressively updated in 2011. The PFS report
and this Technical Report are based on the Rev P3 update of February 2011.

The capital cost estimate for the process plant was based on process flow sheets
and the attendant plant requirements to produce saleable products of molybdenum,
copper with silver credits, and rhenium as ammonium perhennate.

The operating cost estimate was based on labour, power, consumables and
maintenance to operate the aforementioned process plants.

The flow sheet for the process plant engineering study is divided into three sections:




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•     Crushing (Figure 17.1).
•     Concentrate Production (Figure 17.2).
•     Concentrate Treatment (Figure 17.3).

The crushing facilities are located at The Project mine site and consist of a Run of
Mine (ROM) ore pad, three stages of crushing, and a fine ore stockpile.

The concentrate production facilities are located at Osborne and consist of a fine ore
stockpile, ball mill, rougher, scavenger and cleaner flotation to produce a bulk
sulphide concentrate. This is followed by concentrate regrind and copper depression
flotation stages to produce both molybdenum and copper concentrates. The
molybdenum concentrate is thickened, filtered and dried.

The concentrate treatment facility is to be located adjacent to the concentrate
production facility at Osborne. The facility consists of a concentrate roaster and off-
gas scrubbing system including a sulphuric acid production plant, together with
hydrometallurgical plants to purify the molybdenum and rhenium to saleable grade
products. Note that in Figure 17.3 the strip solution from the SX advances to IX as
shown, and the raffinate is neutralised and disposed of (not shown). The strip
solution from the IX advances to crystallization as shown, and the barren solution
advances to the calcine leach or recycles to the quench tower (not shown).

The processing plants are supported by utility and ancillary facilities. Utilities include,
water treatment facilities, compressed air, along with reagent mixing and dosing
systems. Power will be provided from the Osborne power station. Site ancillaries for
the Osborne site include an administration building, maintenance workshops, a
canteen and change house, laboratory, store, and domestic waste treatment
facilities.

Key process plant parameters are outlined in Table 17.1.

Table 17.1       Key Process Plant Parameters
                  Key Design Parameters                     Value          Unit
      Total ROM ore feed                                   500,000           dt/a
      Molybdenum grade, ROM ore                              1.1            wt%
      Molybdenum grade, concentrate                          30             wt%
      Molybdenum production                                 5,030             t/a
      Overall molybdenum recovery                            84.5             %
      Rhenium production                                    7.209            dt/a
      Overall rhenium recovery                               80.9             %


A number of option studies were undertaken, all based on modifying the existing
Osborne plant to produce molybdenum concentrate and allowing for spare capacity
to treat copper/gold ore which may be mined. As the modification required to the
existing processing plant would result in reducing the capacity of the copper/gold
circuit, IVA directed that a new and separate molybdenum concentrate circuit was to
be constructed at Osborne, so as not to impact on the capacity of the existing




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copper/gold circuit. This involved the following changes to the Options Study
facilities:
At Merlin
•     ROM pad, crushing facilities and ore handling facilities.
•     Crushed ore transported to Osborne.

At Osborne
•     Crushed ore delivery and stockpiled on existing ROM pad.
•     Ore reclaim and conveying to new ball mill.
•     New flotation, concentrate filtering, storage and load-out.
•     Tailings dewatering.
•     Concentrate treatment facility and associated services and utilities.

The crushing facilities located at the Merlin mine site consist of a ROM ore pad, three
stages of crushing, and a fine ore stockpile.

The concentrate production facility located at Osborne consist of a ball mill, rougher,
scavenger and cleaner flotation to produce a bulk sulphide concentrate, followed by
concentrate regrind and copper depression flotation stages to produce both
molybdenum and copper concentrates. The molybdenum concentrate is thickened,
filtered and dried ready for transfer to the concentrate treatment plant.

The concentrate treatment facility consists of a concentrate roaster and off-gas
scrubbing system incorporating a sulphuric acid production plant, along with
hydrometallurgical plants to purify the molybdenum and rhenium to saleable grade
products.

The process plant is not innovative and there are several examples of similar plants
in commercial operation:
•     The ore preparation circuit is designed to treat 500,000 t/a dry ore. There is no
      design margin applied to the mill design capacity. Run-of-Mine (ROM) ore is
      transported 50 km by truck to the concentrate production plant.
•     Approximately 18,000 dt/a of molybdenum concentrate is produced in the
      concentrate production plant. The concentrate contains 30% molybdenum.
•     The concentrate treatment facility includes roasting of the sulphide concentrate
      to produce a molybdenum calcine.
•     The roaster is designed to treat a total of 18,000 dt/a in a single roaster.
•     The downstream calcine leach is designed to treat 15,000 dt/a.
•     The average power load for the crusher at the Merlin mine site is 600 kW, and
      for the Osborne concentrator (including the ball mill) and roaster plant are
      2,500 kW and 1,500 kW respectively.
•     The process plant water requirement is 72 m 3/h.




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•      The annual requirement for process plant materials is approximately 7,000 t.

The key ROM feed grades used for the study are given in Table 17.2.

Table 17.2        ROM Feed Grades
       Ore          Mo %        Cu %               Re g/t        Fe %        Ag g/t    Feldspar %       Quartz %

     Merlin          1.2             0.2            18           1.19          9             60.7          26.1


For the average life of mine feed grade, the estimated overall molybdenum and
rhenium recoveries are given in Table 17.3. This estimate will be refined in further
studies based on further mine schedule information and variability testing.

Table 17.3        Metal Recoveries
                                                            Molybdenum (%)             Rhenium (%)
          Overall (from ROM feeds)                               84.5                         80.9


Production estimates based on average feed grade are outlined in Table 17.4.

Table 17.4        Plant Production Estimates
                                      Mo Trioxide               Contained Mo          APR            Contained Re
              Product
                                          t/a                        t/a               t/a                t/a
        Annual Production                  7,605                    5,069             10.48              7.28
    Note: The feed grade in Table 17.2 used by Jacobs and the plant production estimates in Table 17.4 vary
    marginally to those calculated in the Financial Analysis.




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Figure 17.1      Crushing Circuit Schematic




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Figure 17.2      Concentrate Production Plant Schematic




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Figure 17.3      Concentrate Treatment Plant Schematic




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18    PROJECT INFRASTRUCTURE

Significant new plant and infrastructure will be established at the brownfield Merlin
mine site to support the underground mine. Ore from the Merlin mine site will be
crushed on site and hauled to Osborne for treatment in a new molybdenum / rhenium
treatment facility. The run of mine (ROM) pad for Merlin mine site will be located in
the area north of decline boxcut and the existing haul road between the Mount Dore
Copper Heap Leach Project and Merlin mine site. The design of these facilities will
form part of the Feasibility Study and should include:
•     Surface infrastructure and facilities:
      -      Administration office complex.
      -      Heavy vehicle workshop.
      -      Hydrocarbon storage facility.
      -      Light vehicle workshop.
      -      Wash down bay.
      -      Refuelling bay.
      -      Settling ponds.
      -      Box-cut services pad.
•     Underground infrastructure and facilities:
      -      Electrical distribution.
      -      Primary dewatering system.
      -      Primary ventilation system.
      -      Refuelling bay.
      -      Service bay.
      -      Magazine.
      -      Communications.

Power supply to the underground workings will come from an overhead feeder which
will terminate adjacent to the surface portal.

Studies undertaken for PFS purposes indicate a maximum demand of 8.0 MVA at an
11 kV supply voltage.

The ventilation circuit for the mine will consist of two major intake airways, the Main
Decline and Fresh Air Raise, and two major exhaust airways, the North Return Air
Raise and South Return Air Raise. The mine will be divided into two separate,
ventilation districts, North and South, although there will be some limited interaction
between the two districts along connected foot wall drives and shared intake airways.
The PFS work indicated a total mine airflow requirement of 460 m 3/s.




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The water supply for the mine is assumed to be provided by a bore field adjacent to
the Merlin mine site. Estimates prepared as part of the PFS suggested that the mine
demand for process water will be 13 l/s.

Water pumped from the mine will be treated in surface settling ponds and be
available for reuse or transfer to evaporation ponds for disposal.

18.1      Paste Plant

Besides the mining operations and associated infrastructure listed above the Merlin
mine site will include a paste fill backfill plant. This plant will be a new dry paste fill
plant or the relocated existing Osborne paste plant if it is not required for ongoing
mining operations at Osborne. There is a sufficient supply of dry tailings at Osborne
and at nearby old Starra tailings dams for making backfill at the Merlin mine site. The
Merlin mine site paste backfill plant will have a rated capacity of 144 m 3/h processing
approximately 250 t/h of tailings. Tailings could be backhauled from Osborne in the
trucks used to haul ore to Osborne. The feasibility study will include further work on
the selection of a backfill option.

The paste plant will be located on top of the proposed mine workings at about mid
distance along orebody strike in order to enable gravity delivery of paste fill to most
of the mining areas. This will position the plant on the eastern side of the quartzite
ridge, above the underground production areas. The final plant location will be
determined in future studies.

Paste fill will be delivered underground by gravity through one of the two inclined
175 m long cased 200 mm NB surface boreholes and a series of internal boreholes
and level distribution pipelines. Little Wizard and upper south areas will have their
own borehole which will receive paste fill by agitator trucks.

The paste production system is simple. Harvested dry tailings will be fed to a dry
paste backfill plant similar to the one at Osborne using a front-end loader. Tailings
will be repulped and mixed with cement. These dry system paste plants are
essentially “off the shelf”, have short delivery times and will require up to six months
to commission. The paste backfill plant at Osborne has been in use for three years
only and is in very good condition. If future operations at Osborne will not require
paste backfill then it may be more cost effective to relocate the plant to the Merlin
mine site.

18.2      Infrastructure

Infrastructure is required to support The Project, both at the Mount Dore/Merlin mine
site area and at the Osborne site. The infrastructure components include power,
water supply, aerodrome, accommodation villages and communications.

18.2.1    Electric Power

•      Power by contract generation, approx 10 MW. This will be a gas-fired power
       station supplying power at 11 kV, with the gas trucked from Osborne.




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•      Power supply to the underground workings will come from an overhead feeder
       which will terminate adjacent to the surface portal.
•      Power quality equipment (eg: harmonic filters and power factor correction) will
       be provided at the mine site instead of back at the power station. This will
       provide flexibility of supplying future loads by the same power station. Provision
       will be made for expanding the power station to enable supply to other future
       mines if required.
•      The existing power station at Osborne will be refurbished/upgraded to provide
       24 MW.

18.2.2    Other Infrastructure

Other infrastructure required at Mount Dore/Merlin mine site includes:
•      Water supply from nearby boreholes.
•      The existing aerodrome will be used for small charter aircraft commuter flights
       primarily from Mount Isa and for medical evacuation where necessary.
•      A microwave dish will be installed to provide a link to the Osborne site.

The infrastructure facilities to support the Osborne site will include:
•      Water supply will be from the existing borefields at the Longsight sandstone
       aquifer.
•      The existing Osborne aerodrome will be used for fly in fly out commuter
       services and for medical evacuation.
•      Accommodation is provided by a 320 room village 4 km south of the Osborne
       mine. An additional 132 man contractors camp is also available.
•      Mining personnel will commute daily to the Merlin mine site at Mount Dore.
•      The Osborne site is serviced by a microwave link that connects into the
       national grid. A mobile radio installation is being constructed to allow for mobile
       phone usage around the Osborne site.

18.3      Tailings and Effluent Management

Metago Environmental Engineers (Australia) Pty Ltd (Metago) prepared a report on
tailings and effluent management for the PFS.

At Osborne, high density tailings (HDT) were deposited on Tailings Storage Facility 2
(TSF2) from 2002 up to cessation of the Barrick Osborne operations in 2010. This
facility has approximately 100,000 m 3 of capacity remaining for tailings storage.

IVA proposes to process copper ores mined from Kulthor, Lucky Luke and other
orebodies at a rate of 1.5 Mt/a for 9 years. It is proposed that management of the
tailings continue to be at high density in light of the cost and water savings previously
demonstrated at Osborne. Assuming a density of 1.8 t/m 3 for HDT, the capacity
required for the copper tailings is 5 Mm 3 which means that additional tailings storage
capacity is required for the copper tailings.




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Processing of molybdenum ore from The Project is proposed to be at a rate of 0.5
Mt/a for 16 years and, given the benefits in cost and water savings, it is proposed
that these tailings will also be managed as a high density slurry. At a density of 1.8
t/m 3 the volume of molybdenum tailings that need to be stored is 4.5 Mm 3. A new
storage facility for these tailings will also be required.

Three options were considered:
•      Option 1 - A new HDT facility located on immediately west of TSF1.
•      Option 2 - A new HDT facility located between TSF1 and TSF2.
•      Option 3 - An expansion of TSF2 to the south and west, together with the
       relocation of the current return water dam.

The concept-level study assessed the long term deposition layout and embankment
development to contain the HDT for the copper and molybdenum scenarios
proposed. The conceptual design includes:
•      New containment bunds for each scenario, sized and positioned to contain the
       required tailings volumes.
•      Stormwater, seepage and collection channels where deemed necessary to
       allow free movement of seepage and runoff water from the facility as well as to
       keep ‘clean’ and ‘dirty’ runoff separated.
•      Spillways and spillway locations based on bund alignment.
•      A contributing catchment assessment of the storage requirements for the
       proposed new reclaim dam.
•      Pipework including the length of the tailings delivery pipelines and the reclaim
       water pipeline from the reclaim pond to the plant.
•      A complete set of conceptual level drawings for each option.
•      A schedule of quantities and a cost estimate for all potential options.

From these assessments, Table 18.1 shows the cost (A$) per m 3 of tailings material
deposited for the potential TSF’s Options.

Table 18.1         TSF Options Comparison based on A$ per m3 Tailings
      TSF Option      TSF Footprint    HTD Capacity        Total Cost   Cost/Capacity Ratio
                          (ha)             (m 3)             (A$M)            (A$/m3 )
    Option 1A               98             5,073,000        8.2                 1.61
    Option 1B               124            7,366,000        8.6                 1.16
    Option 1C               89             4,736,000        7.9                 1.67
    Option 2                139            4,615,000        8.2                 1.77
    Option 3                125            5,147,000        7.5                 1.45


Further studies will address:
•      Total tailings capacity for each option.
•      Practicality and costs, including relocation of existing sites and facilities.




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With respect to gypsum and effluent management facilities at a potential smelter in
Townsville, Metago developed a concept level design for the gypsum evaporation
facility to assess the feasibility of evaporating effluent at Townsville. In addition, a
comparison of evaporation facility requirements at Townsville compared with
Osborne was made. The assessment of the gypsum storage facility included:
•      Sizing of a three year storage of gypsum and ammonia leach in a contained
       storage pad.
•      Preparation of a schedule of quantities for the storage pad.

A similar assessment was made of an effluent evaporation facility. While it was
concluded that the gypsum storage costs will be similar for either Townsville or
Osborne locations, the higher evaporation potential at Osborne results in a smaller
and hence lower-cost facility.

18.4      Personnel

The implementation of construction of the process plants, infrastructure and haul
road will be undertaken by the Owner’s team supervising the engineering company
under an EPCM contract or other contract defined during the Feasibility Study stage.

Once constructed, The Project will be managed by a General Manager Operations
responsible for managing two separate operations:
•      Mining copper/gold sulphide ores at Osborne and processing the ore through
       the existing 2 Mt/a Osborne processing plant.
•      Mining the Merlin mine site molybdenum/rhenium ore and processing the ore
       through a new 500,000 t/a flotation plant and roasting plant at Osborne.

Both copper/gold and molybdenum/rhenium operations will use shared services for
geology, mining, processing, maintenance, administration and health safety and
environment. Operating costs for these shared services are proportioned 50:50
between the two operations.

The mining workforce for the initial decline development at the Merlin mine site is
being provided by a contractor. The PFS assumes that the development and
operation of the mine will be by IVA. The peak workforce for the mine is estimated at
180 employees and contractors with an average of around 160 personnel. The
processing and maintenance workforce is estimated at 85 employees for concentrate
production and 51 employees for concentrate treatment.




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Table 18.2         Operating Workforce and Numbers of Personnel
                                                                         Additional.
                                                                 Stand                 50% of
                                                                 Alone   to Copper     Shared
                                                                           Plant
        The Project Management *
           Management                                              3         1           2
        Mining - Merlin Mine Site
           Management, supervision, technical                     10         8           9
           Operators, trades, clerical                            170       170         170
           SUB TOTAL                                              180       178         179
        Processing – Concentrate Production
           Management, supervision, technical                      5         1           2
           Operators, cleric al                                   42         12         15
           Maintenance mgmt, supervis ion technic al              14         0           7
           Maintenance operators ,trades                          24         4          10
           SUB TOTAL                                              85         17         34
        Processing – Concentrate Treatment
           Management, supervision, technical                      2         2           2
           Operators, cleric al                                   37         37         37
           Maintenance mgmt, supervis ion technic al               5         5           5
           Maintenance operators ,trades                           7         7           7
           SUB TOTAL                                              51         51         51
        Adm inistration *
           Clerical, w arehouse, power , air port , logistic s    25         5          15
           Village contractors, maintenance.                      15         9          12
           SUB TOTAL                                              40         14         27
        TOTAL EMPLOYEES                                           359       261         293


The “Stand alone” workforce numbers assume that only the Merlin ore body is mined
and processed through the modified Osborne treatment plant.

The “Additional to Copper Plant” workforce numbers assumes that a standalone
molybdenum rhenium flotation plant and roaster and the existing Osborne copper
gold plant are operated at the same time and requiring additional personnel.

The “50% shared” workforce assumes both the Osborne copper gold plant and
molybdenum rhenium processing plants are operating. The total workforce numbers
are shared between the two operations. The costs associated with this shared
workforce are used for the operating cost analysis.




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19     MARKET STUDIES AND CONTRACTS

19.1      Marketing Studies

The Project is unusual with respect to the grades of molybdenum and rhenium and
has a number of attributes which make it unusual in a geological sense. This will
strongly impact on the marketing strategy that is to be adopted. These attributes are:
•      The Project’s average mine grade is 1.1 % molybdenum and 19 g/t rhenium
       compared to the Henderson Mine (USA) grade of 0.2% molybdenum which has
       been considered the world’s highest grade molybdenum mine.
•      The rhenium in the Merlin deposit represents the first direct source of rhenium
       in the world and is not a typical by-product rhenium produced by convoluted
       molybdenum by-product streams via primary copper flotation.
•      The Project production represents approximately 3.5% of the world output of
       molybdenum and approximately 15% of the world production of rhenium.
•      The Project rhenium can be marketed as “secure supply”.

19.2      Molybdenum

As part of this economic evaluation, IVA commissioned a report on the outlook for
molybdenum and rhenium markets from Roskill (2011). The following discussion
regarding the molybdenum market has been sourced from that report.

While the market for molybdenum is transparent and well understood, the rhenium
market is opaque, shadowy and small. These factors will affect how the products are
marketed.

Molybdenum is used primarily as an alloying agent in steels. Molybdenum enhances
wear resistance, toughness and corrosion resistance in steels. It is also used as a
catalyst for some chemical reactions. In the sulphide form of molybdenite it is used
as a lubricant and in the oxide form used for pigments, tool steels, super alloys, cast
iron, welding alloys, high strength low alloy steels and stainless steels.

Molybdenum concentrates are either processed to the oxide by companies with their
own roasting plants or sold to toll treatment roasting operators. They are normally
sold on long term contracts and occasionally on spot. Spot sales rise when there is a
high demand for the product. The molybdenum price has been low up until 2003
when the price increased substantially then dropped in the GFC. The price is now
recovering to around the US$ 15-18 /lb range.

Between August 2008 and March 2009 molybdenum prices, responding to the global
financial crisis and downturn in world economies, fell from US$34 /lb to US$8 /lb. In
2010 prices recovered to US$16 /lb Mo. Until the financial crisis, prices had risen in
response to supply limitations and growing demand, principally in China. In the next
five years the molybdenum market will resume the growth in the years to mid-2008,
with consumption expected to grow by 5% per year for the period 2011 to 2015.




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In 2011, a recovery in global economies saw production of mined molybdenum reach
a record high of 225 kt, an increase of 12% year-on-year, having declined by around
6% in 2009. The increase in production in 2010 took place across producing
countries, with output rising by 10% in China, and 26% in the USA. Prior to this,
global output had increased strongly. For the period 2000 to 2008, production
increased by an annual average rate of 5% per year.
Molybdenum production is concentrated in a relatively few countries, with China, the
USA and Chile accounting for almost 80% of total world production in 2010. Chinese
molybdenum production was the largest in the world from 2002 to 2004 but fell
sharply in 2005 because of government-enforced mines closures. It then more than
doubled between 2005 and 2008, to regain its leading position. In 2010 Chinese
output was 80 kt, rising by 10% over 2009.
The ten largest producing companies accounted for 65% of world molybdenum mine
production in 2010. Freeport McMoran in the USA and Codelco in Chile were the
largest producers, together accounting for 23% in 2010.

Consumption

Consumption of molybdenum is also estimated to have risen by around 12% in 2010,
as improving economic sentiment has seen demand for all commodities rise.
Demand at the world level has increased faster than that of mine supply. For the
period 2000 to 2008, when global consumption was at its largest, demand rose by an
average of 5.4% per year.
China’s domestic economy continued to grow during the global recession, with the
result that it became the world’s largest consumer of molybdenum. In 2010, Chinese
consumption was 60 kt, accounting for almost 30% of the world total.
Molybdenum is consumed mainly as the oxide or ferromolybdenum in stainless and
low alloy steels. Estimated worldwide demand for molybdenum by application in
2010 is shown in Table 19.1.
Table 19.1          Estimated 2010 World Molybdenum Demand by Application, 2010
Application                                                       Kt       %
Stainless steel                                                   50       24
Full alloy steel                                                  33       16
Tool and high speed steel                                         23       11
High strength low alloy (HSLA) steel                              21       10
Carbon steel                                                      19        9
Catalysts                                                         17        8
Molybdenum metal and alloys                                       12        6
HPA/Superalloys                                                   10        5
Cast iron                                                         12        6
Lubricants                                                        4         2
Pigments/corrosion inhibitors                                     4         2
Other chemical                                                    2         1
Total                                                             208      100
Source: IMOA, Roskill




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Prices

Between August 2008 and March 2009 the price of molybdenum oxide fell from
US$34.50 /lbMo to a low of US$8 /lbMo, and that of ferromolybdenum fell from
US$81.50 /kgMo to US$20.50 /kgMo, levels not seen since early 2002.

Cutbacks by major producers, and demand in China remaining strong, resulted in
tightening of supplies and a weak rally of molybdenum prices, which had doubled by
July 2009 to peak at US$18 /lbMo for oxide and US$40 /kgMo for ferromolybdenum.

During 2010, molybdenum prices remained stable, without much volatility. Oxide
prices in Europe averaged US$16 /lbMo, with ferromolybdenum prices averaging
US$40 /kgMo. In the first seven months of 2011, prices rose slightly, averaging
US$16.67 /lbMo and US$40.96 /kgMo, for oxide and ferromolybdenum respectively.

Market Outlook

With the decline in prices in late 2008, many new mining projects were slowed or
suspended. With more than thirty molybdenum mine projects with combined capacity
of about 136 ktpa at some level of feasibility assessment, longer term supply appears
to be assured.
Over the next five years, around 140 kt of new roasting capacity is expected to come
online, of which around 60 kt will be located in China. This will raise world capacity to
more than 400 ktpa, which is more than sufficient on a global basis.
Demand for molybdenum is expected to grow at above 5% per year up to 2016,
driven primarily by 7% per year growth in China. Production of stainless, carbon and
alloy steel is expected to increase strongly throughout the forecast period.

The outlook for molybdenum prices up to 2016 is one of some volatility. In 2011 and
2012 prices are likely to firm towards US$18 /lb as demand increases at a faster
pace than new supply. As the market moves into deficit after 2013, prices will rise to
around US$30 /lb.

Table 19.2       Forecast Molybdenum Oxide Price
                                         Nom inal Price    Real Price
                             Year
                                           (US$/lb)       (2011 US$/lb)
                             2010             15.91           16.37
                             2011             16.50           16.50
                             2012             17.45           17.03
                             2013             19.25           18.30
                             2014             23.10           21.41
                             2015             25.35           22.88
                             2016             27.00           23.77
                        Source: Roskill forecasts




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19.3      Rhenium

The following discussion regarding the rhenium market has been sourced from the
Roskill report (Roskill, 2011) on the outlook for the molybdenum and rhenium
markets.
Rhenium has the unique property that, when alloyed, it improves mechanical
strength and reduces plastic deformation at elevated temperatures and as such is
used in a number of high temperature applications including:
•      Aerospace applications such as turbine blades and other static parts.
•      Rocket thruster elements.
•      Reforming catalysts for the petroleum industry.
•      Small high temperature applications such as crucibles, electrical contacts,
       electromagnets, electron tubes, heating elements, ionization gauges, mass
       spectrographs, semiconductors, temperature controls, thermocouples and
       vacuum tubes.
•      Medical applications.

Over 80% of rhenium is used in the aerospace industry in single crystal nickel
superalloy turbine blades operating in the high temperature zone of the engine.

Rhenium addition to the alloys allows the turbines to be operated at higher
temperatures without creeping. It therefore is a major contributor to improved fuel
efficiencies of jet engines over the past ten years and the use of rhenium in jet
turbines is expected to increase substantially in future. Rhenium is sold in the form of
ammonium perrhenate (APR), which has the chemical formula NH4ReO4.

Rhenium prices have settled down after the volatility and high prices seen in 2008.
The credit crisis and associated recession in the major economies (China excepted)
burst the rhenium ‘bubble’ and prices plummeted. Prices fell back to US$4,000 /kg
after averaging just over US$6,000 /kg in 2009, almost half of the average spot price
in 2008.

In Roskill’s rhenium report published in mid-2010 (Roskill, 2010), rhenium prices
were forecast to average between US$4,000 and US$4,250 /kg in 2010 and between
US$4,750 and US$5,500 /kg in 2011 as the market recovered from the steep drop in
2009.As shown in Table 19.3, the average price for rhenium metal in 2010 was
US$4,300 /kg, which was close to the figure forecast. However, in the first 6 months
of 2011, rhenium prices have averaged US$4,250 /kg, significantly below the figure
forecast in Roskill’s 2010 report. Even allowing for a slight recovery in rhenium prices
in the second half of 2011, the 2011 average is unlikely to top US$4,500 /kg, which is
still 10 to15% lower than the original Roskill forecast.

It appears that the rhenium market has found a new equilibrium level at between
US$4,000 and US$5,000 /kg. This is still significantly higher than the contract prices
recorded in Chile in the 2010 report, although as contracts are renewed the
disconnect between spot and contract prices has narrowed considerably.




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Compared to the forecasts made in the 2010 report, Roskill does not expect to see
any significant changes to the forecast supply/demand balance in the remainder of
the forecast period. On the demand side, the catalysts market continues to exhibit
steady growth, as predicted. The aerospace sector is cyclical and a flattening out in
demand was expected in 2012, while high oil prices may also have a dampening
effect on demand for air travel with a knock on effect for the superalloy industry.
Moves by some consumers, such as GE, to reduce or even remove rhenium from
superalloys appear to have had a limited effect on demand to date.

On the supply side, there would appear to be sufficient refining capacity on-line and
in the pipeline to cope with the forecast growth in demand to 2016. However, the
quantities of rhenium in concentrates may continue to be a constraining factor. A
significant future source of rhenium units could come from increased recycling in the
superalloy industry. However, the organisation of recycling in the industry is still
somewhat haphazard and with rhenium prices at much lower levels than 2 years ago
the incentive is slightly less pressing.

It appears that the Roskill forecasts made in 2010 were a little on the bullish side and
given that prices in 2011 are likely to be 15% or more lower than forecast in the 2010
report, we have adjusted our forecasts for 2012 to 2016 accordingly. As we expect
the forward supply/demand curve to be similar to the 2010 forecast, the future price
trends for rhenium are forecast to be similar, albeit at a slightly lower level.

Table 19.3          Forecast Average Rhenium Metal Prices 2010 to 2016

                                                                  Actual Price and Revised Forecast
       Year                 Original Forecast Price (US $/kg)
                                                                              (US$/kg)

                             High                      Low

       2010                  4,000                    4,250                     4,300

       2011                  4,750                    5,500                  4,100-4,600

       2012                  4,500                    5,200                  3,750-4,250

       2013                  5,000                    5,800                  4,250-4,750

       2014                  6,250                    7,000                  4,750-5,250

       2015                  6,500                    7,500                  5,250-5,750

       2016                                                                  5,500-5,900
Source: Roskill forecasts


The marketing strategy is to sell the molybdenum trioxide (MoO3) in long term supply
contracts and on the spot market, while the bulk of the APR would be sold in a long
term supply agreement with aerospace engine manufacturers. Production peaks will
be sold on the spot market.

The PFS includes an option to sell concentrate to a toll roasting outfit and not build a
roaster and rhenium recovery plant.




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The base case for the PFS is to produce all of the molybdenum rhenium concentrate
at 30% Mo grade. Provided the ore can be mined selectively, it will be possible to
produce two grades of molybdenum rhenium concentrates.
•      A high grade product of >50% Mo grade representing 60% of the molybdenum
       production.
•      A medium grade product of 30% Mo grade representing 40% of the
       molybdenum production.

These two sulphide concentrate products have different cost structures for sale or toll
treatment. The following assumptions were made:

Concentrate Grade > 50% Molybdenum
•      Processing Cost US$ 0.65 - 0.75 /lb Mo (assume US$ 0.70 /lb for this study).
•      Metal Return         99% Molybdenum payable.
•      Rhenium              ~10% metal return.

Concentrate Grade ~30% Molybdenum
•      Processing Cost US$ 1.50 /lb Molybdenum.
•      Metal Return         93% Molybdenum payable.
•      Rhenium              ~10% metal return.

19.4      Contracts

AMC is aware of the following contracts and understands that they have been
awarded or will be awarded on normal commercial terms.

19.4.1    Mining

The box cut and exploration decline was started under contract in the fourth quarter
of 2010. The scope of work is to access the Merlin orebody at the end of the second
quarter of 2012. Development to the orebody and extraction of ore will then
commence around the third quarter of 2012. This timetable sets the time to
implement construction and commissioning of the crushing and flotation plants to be
available by the third quarter of 2012.

The decline was adjacent to the Little Wizard orebody in late June 2011 and this now
permits an access to be driven to start trial mining of Little Wizard. Ore recovered
from this orebody will be stockpiled until the concentrator is available to process the
ore. Work is being undertaken by an experienced mining contractor, managed by the
IVA Operations staff.

19.4.2    Haul Road Access to Osborne

A 53 km long haul road is planned to join the Merlin mine site with the Osborne
treatment plant. A redesign is under way to reduce the cost and improve construction
efficiency. No contract has yet been awarded. Construction time is approximately




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10 months depending on the timing of the wet season at the end of the year. It is
proposed that road construction will be implemented by one civil contractor and
managed by an IVA team reporting to the IVA Project Group.

19.4.3    Power Station

The existing power station at Osborne includes five diesel/gas Wartsila 3.85 MW low
speed and one 100% gas Jenbacher 3.35 MW medium speed generators. The
Wartsila engines require major overhauls and conversion to 99% gas usage. IVA
expected to let contracts in July 2011 for upgrades to these engines with a
completion date of September 2012. IVA expects that the work will be managed by
Wartsila and assisted by IVA staff.

19.4.4    Process Plants

Jacobs has prepared an implementation plan for the process plants. The plan
excludes other areas of the project, such as mining, infrastructure and tailings.

The process plant facilities are based on two separate locations of the facilities:
•     Merlin mine site for the Crushing Plant.
•     Existing Osborne facility for the Concentrator Plant and Concentrate Treatment
      Plant.

The Merlin mine site will be a greenfield site and Osborne will be a new plant on an
existing facility which is effectively greenfield except for tie-ins to existing utility
systems.

Due to the separate locations of facilities the implementation strategy for each will
vary.
•     Engineering will generally be performed in a single location.
•     Procurement will be performed using world-wide suppliers to ensure that best
      quality, price and delivery is achieved. The Procurement function will be co-
      located with Engineering such that design can be progressed efficiently.
•     Separate construction contracts will be awarded within each facility. The split of
      contracts will be generally based on engineering discipline, schedule and trade
      competence. A typical split is:
      -      Earthworks.
      -      Civil works.
      -      Structural / mechanical / piping installation.
      -      Electrical & instrumentation installation.




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19.4.5    Feasibility Study

AMC and other parties have been contracted to undertake work which will form part
of the Feasibility Study design of development and mining operations for The Project.
This work will include the detailed mine design and life of mine plans.

19.4.6    On-going Engineering Work

Jacobs is in the process of obtaining tender prices for major and minor equipment
and starting detailed design of the Mo/Re flotation process plant. This work is being
progressed so that critical path equipment can be purchased to maintain the
schedule to start processing the Project ore in the third quarter of 2013.




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20     ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY
       IMPACT

The current Plan of Operations identifies a total rehabilitation liability of $4.9 M, of
which $3.7 M is for rehabilitation of Legacy Sites associated with previous operations
which Ivanhoe has a legal responsibility to resolve. Further assessments of surface
and ground water quality are being undertaken along with acid producing potential of
waste rock and ore which will further define these liabilities and the potential controls.
Whilst these Legacy Sites are separate from The Project they are part of the overall
Cloncurry Project.

20.1      The Project Operation:

The Project is a simple underground mining operation with relatively low volume
production of 500,000 t/a ore. Waste rock will largely be disposed on surface in
relatively low volume waste rock dumps located to fit within the existing topography.
The development of The Project relies on the dewatering of the adjacent Mount Dore
aquifer which has been partly dewatered during past operations and recovered to
pre-existing levels within 6 years and water quality once dewatering ceased.

20.2      Waste Rock Characterisation

The decline is being excavated in the Staveley Formation, which is separated by a
40 m thick inert quartzite unit from the mineralised Kuridala formation containing the
Merlin deposit.

For the exploration decline sixteen samples of the strata, representative of the waste
rock from the decline development had been analysed for acid producing potential
and reported in the current Plan of Operations. The samples are all classed as ‘non
acid forming’ and geochemically benign.

A second geochemical assessment of the waste rock and ore from the mine areas
has confirmed this earlier observation on the decline material and indicates that a
small proportion of the waste rock generated from the development of the deposit is
PAF (potentially acid forming). It is expected that only low volumes of potentially acid
forming waste rock will be encountered close to the footwall of the ore body and this
will be managed by preferential handling and disposal into one of the existing open
pit voids.

During the mining phase progressive sampling of decline development waste will be
undertaken to confirm the material designation from the previous assessments.

Ore will be managed on an elevated ROM pad and crushed prior to transport to the
Osborne site for processing. The ROM pad and crusher will drain to a pond designed
to hold a 1:100 year 2 month wet season. Due to the turnover of ore material it is not
expected that significant oxidation or liberation of metals or salts will occur. The catch
dam is a conservative option to contain any deleterious contaminants in all except
extreme events. Regular water sampling will also provide early warning of any
contaminant development.




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20.3      Groundwater

Apart from the Mt Dore aquifer there are no significant ground water resources within
a radius of at least 10 km. The aquifers associated with the mineralised areas are
generally sparse and difficulty has been experienced in obtaining sufficient
groundwater for exploratory drilling programs. As the ore bodies on Starra were
developed outflows of water increased with depth and it is apparent that the orebody
is the aquifer while the surrounding sediments of the Stavely formation are tight with
little to no groundwater. The Merlin Mine is bounded to the east by the Mt Dore
Granite and to the west by the Stavely formation both of which appear to act as
obstacles to any significant groundwater movement. The mineralised aquifers have
water quality which is generally moderately to highly saline with elevations of arsenic,
boron, cobalt, copper, fluoride, iron, manganese, molybdenum, uranium, zinc and
sulphates. There is little opportunity for significant impact on groundwater systems.

20.4      Surface Water

The area is drained by ephemeral streams that flow only during wet season rain
events and for a short time after. There are naturally significant metal levels in
surface soil and stream sediments adjacent to mineralised zones. This equates to
significant metal elevations in surface water in undisturbed areas of arsenic, cobalt,
copper, iron, lead, manganese, nickel and zinc. Within the context of the mineralised
surface water quality and the proposed management of mineralised material at
Merlin it is unlikely that there will be any significant impacts on surface water quality.

20.5      Conclusion

The above summaries indicate the proposed Merlin operation has a low probability of
causing any significant environmental liabilities. It is also recognised that The Project
contains a number of Legacy Sites for which the environmental liability has been
summarised in the current Plan of Operations at approximately $3.7 million. Further
detailed assessments may increase the cost of the known liability. Active
rehabilitation and control of the Legacy Sites will be implemented from 2011, thus
progressively reducing liability.

20.6      Permits

Permits currently held for The Project include:
•      Environmental Authority MIN100459006.
•      Plan of Operations May 2011 authorising the Merlin mine site exploration
       decline, commencement of dewatering from the Mount Dore aquifer and
       construction of the access road to Osborne.
•      Water Licence number 69411J for 260 megalitres/year.
•      Water Licence number 604203 for dewatering the Mount Dore aquifer.
•      Riverine Protection Permit 491006 for access across the Mort River.
•      Riverine Protection Permit 490992 for alternate access across the Mort River.




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•      Quarry Permits SP20110851, 20110852, 20110854, 20110855, 20110856 for
       borrow pits to construct access road.
•      Add purposes of processing and roasting to Mining Lease 90040 to facilitate
       the construction and operation of the roaster and concentrator on the Osborne
       mining lease. Accepted by the regulator subject to the following Environmental
       Authority amendment.

Further Permitting Required:
•      Environmental licence ERA16 to support the development of Quarry Permits.
•      Amendment to the Merlin Environmental Authority MIN100459006 to authorise
       the full mine development.
•      Amendment to the Osborne Environmental Authority MIN100459006 to
       authorise construction of the roaster, concentrator and tailings storage facility.
•      Amended Merlin Plan of Operations to authorise the full mine development.
•      Amended Osborne Plan of Operations to authorise construction and operation
       of the roaster, concentrator and tailings storage facility.
This permitting work is being managed by IVA and approvals are expected by
November 2011.

20.7      Social/Community Relations Requirement

IVA is the owner of Starcross Pastoral Holding which is the owner of the Pastoral
Leases that cover The Project area. IVA has entered into a sub-leasing agreement
with a pastoralist in relation to this ground.

Social and Community relations are an ongoing priority for IVA. Engagement occurs
regularly and as required based on the stage of project. IVA conducts an annual
Community Relations Day each year where mining and exploration plans for the
following year are discussed. This day also includes representatives from the Local
Authority, Queensland Department of Employment, Economic Development and
Innovation (DEEDI) and Department of Environment and Resource Management
(DERM) (the regulators).

Merlin mine site area is covered by a Native Title Claim held by the Yulluna People
(QUD 189/10) with whom IVA has a Cultural Heritage Management Agreement. The
existing mining leases were granted pre-native title and are subject to no
agreements. Any new applications for mining leases will trigger a Right to Negotiate
process.




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21     CAPIT AL AND OPERATING COSTS

21.1      Capital Costs

The capital cost estimate was compiled by IVA and is presented here in summary
format. The capital cost estimate is a collation of a number of specialist consultants’
data for mining, process plants, associated infrastructure and owners costs. IVA’s
compilation of costs was reviewed by AMC for the purpose of this report.

21.1.1    Summary Costs

The total project cost is A$ 518.6 M which includes a contingency of A$ 86.4 M as
shown in Table 21.1. The initial capital to commissioning is A$ 337 M.

Table 21.1         Summary of Capital Cost Estimate
                                                              Total Capital   Capital to First
           AREA                                                  A$M          Production A$M
           Mining *                                                 209.4             60.4
           Processing - Crushing & Flotation                         60.4             60.4
           Processing – Roasting                                     85.8             85.8
           Infrastructure                                            60.6             59.6
           Ow ners Cost                                              13.5             13.5
           Commissioning                                              2.5              1.2
           Contingency                                               86.4             56.2
           Total                                                    518.6            337.0
                                     * The mining costs include Indirects

21.1.2    Direct Costs

Direct costs include mining, process plant and infrastructure requirements. The costs
were obtained from the estimates prepared by AMC for mining, by Jacobs for
processing and by others for infrastructure. The contingency estimates prepared by
third parties are excluded from direct costs. Engineering costs for the process plant
are excluded from direct costs.

The direct costs are summarised in Table 21.2.




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Table 21.2       Summary of Direct Costs

                                                                  Total
                DIRECT COSTS
                                                                  A$M

                Mining
                  Development (AMC)                                 194.3
                  Aquif er de-watering                                2.5
                  Diamond drilling                                    9.6
                  Decline Development (AMC)                           3.0
                Total Mining                                          209
                Processing
                  Crushing (Jacobs)                                   5.8
                  Flotation (Jacobs)                                 37.0
                  Tailings                                            7.0
                   Roaster (Jacobs)                                  68.6
                Total Processing                                      118
                Infrastructure
                  Mining ( Decline development)                      30.0
                  Pow er - Osborne (50%)                              5.2
                  Pow er - Mount Dore (Contract)                      1.5
                  Haul Road (50%)                                    16.5
                  Accommodation for mining & roaster employees        7.4
                Total Infrastructure                                 60.6
                Total Directs                                         388


21.1.3    Indirect Costs

The indirect costs are summarised in Table 21.3.




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Table 21.3        Summary of Indirect Costs

                                                                             Total
                INDIRECT COSTS
                                                                             A$M

                Owners Cost
                   Decline & Dew atering Management                              1.8
                   Project Mgmt Team, Insurance, Travel                          2.5
                   Consultants                                                   2.4
                   Feasibility Study incl. detailed engineering                  4.0
                   Haul Road (50%)                                               0.7
                   Permitting & Govt. Construction Taxes                         2.1
                Total Ow ners Costs                                             13.5
                Jacobs Engineering
                   Crushing                                                      1.5
                   Flotation                                                     9.2
                   Roasting                                                     17.2
                Total Jacobs Engineering                                        27.9
                   Commissioning                                                 2.5
                Total Indirects                                                 43.8
                Note: Mining indirect costs are included in the AMC mining direct costs.

21.1.4    Contingency

Contingency for mining was estimated by AMC at 15%. This is increased to 17% to
include the diamond drilling program and aquifer dewatering.

A contingency of 20% is applied to the total cost of Direct and Indirect costs for
processing, infrastructure and owners costs. AMC considers this level of contingency
appropriate for a PFS quality estimate and is consistent with the contingencies
applied by IVA’s engineers for mining and process plant estimates.

The total contingency is A$ 86.4 M as shown in Table 21.4.

Table 21.4        Contingency Allowance
                                                                                     Total
         Contingency
                                                                                     A$M
         Total Direct & Indirect Costs - Mining                                        209.4
         Contingency at 17%                                                              35.5
         Total Direct & Indirect Costs – Processing & Infrastructure                   254.5
         Contingency at 20%                                                              50.9
         Total Contingency                                                               86.4



21.1.5    Exclusions

Excluded from the capital cost estimate are:
•     Working capital.




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•      Sustaining costs.

21.1.6    Total Project Capital Cost

The total project cost is A$518.6 M as set out in Table 21.1.

Mining capital includes all development work, ventilation, dewatering for the life of
mine.

Processing - Crushing and Flotation includes a three stage plant located at the Merlin
mine site and the 500,000 t/a grinding and flotation facilities at Osborne including
EPCM costs to construct and install.

Processing - Roaster includes the molybdenum multi-heart roaster, gas cleaning and
recovery of molybdenum and rhenium products for sale.

Infrastructure - includes 50% of the Osborne to Merlin mine site haul and access
road, 50% of the re-build and conversion of the power station to 99% gas, additional
accommodation units for mining and processing employees.

Owners costs includes the Owners project management team, consultants, detailed
mine design and plant engineering design to improve the accuracy of the capital cost
estimate.

Commissioning includes specialists in roasting and the cost of the first few weeks of
plant commissioning until first concentrate production.

Contingency of 20% has been allowed to account for the level of accuracy of the
engineering design and study.

21.1.7    Cost to First Production

The total capital costs to first production are estimated to be A$ 337.0 as shown in
Table 21.1.

Mining capital includes the decline development and ventilation required to access
the Merlin orebody to provide ore to commission and operate the flotation processing
plant to first concentrate production. This cost also includes access to the Little
Wizard orebody which will provide the first molybdenum ore. This ore will be
stockpiled until the processing plant is available for commissioning.

21.2      Operating Costs

21.2.1    Introduction

The operating cost estimate was compiled by IVA and is presented in summary
format. The operating cost estimate is compiled from information provided by AMC
for mining, from Jacobs for the process plant together with other data from IVA. AMC
has reviewed the reasonableness of the compilation by IVA for the purpose of this
report.




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21.2.2     Employee Costs

Employee cost estimates were prepared as follows:
•     Management salaries based on existing cost structure and numbers at
      Osborne.
•     Mining - AMC produced operating costs for mining assuming all the work was
      being undertaken by the IVA employees. AMC did not include escalation,
      capital cost, royalties, taxes, employee on-costs, mine water, air-commute and
      accommodation costs.
•     Processing - Jacobs produced the labour costs for the crushing, flotation and
      roasting process plants. Excluded were salary on-costs.
•     Administration cost of salaries is based on the number of employees in the
      areas of commercial, warehouse, IT, human resources, safety, environmental
      and indigenous relations that were employed when Osborne was operating at
      full capacity.
•     An accommodation cost of $50 per person per day was included.

Employee costs were derived from the study to evaluate site options. Employee
costs were calculated on the basis of:
•     A stand-alone operation where only the Project ore was processed through the
      existing modified Osborne flotation circuit.
•     The number of additional employees required to operate a new molybdenum
      rhenium flotation plant while the existing copper gold plant is being operated.
•     The number of employees required to operate both copper gold and
      molybdenum rhenium process plants using a shared workforce which includes
      management, supervision, maintenance and administration.
•     The combined operating costs, less the costs already included in the mining
      and processing.
Employee numbers are summarised in Table 21.5.

Table 21.5         Employee Numbers
                                                     Stand
                                                                  Additional
                                                     alone                      Shared &
                              Area                                 to Cu/Au
                                                     Mo/Re                     Proportioned
                                                                     Plant
                                                     Plant
         The Project Management                        3             1               2
         Mining - Merlin Mine Site                   180           178             179
         Processing - Crushing & Flotation            85            17              34
         Processing - Concentrate Treatment           51            51              51
         Administration                               40            14              27
         Total Num ber                               359           261             293




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21.2.3    Combined Operating Costs

Overall operating costs for the project processing 6.72 Mt of ore over the mine life
are given in Table 21.6.

Table 21.6         Combined Operating Costs
                                                                                 Life of Mine
                               Operating Cost Area                A$ per t Ore
                                                                                     A$M
         Mining                                                        176.00         1,182.1
         Ore Haulage to Osborne                                           5.64           37.9
         Concentrator (50% cost shared w ith Cu/Au Ops)                 31.85           213.9
         Concentrate Treatment (Osborne)                                28.73           193.0
         Site Administration                                            11.99            80.6
         Realisation                                                      6.15           41.3
         Rehabilitation                                                   2.68           18.0
         Average Cash Operating Cost per t Ore                         263.05         1766.8
         Royalty                                                        18.72           125.7
         Average Cash Cost per t Ore                                   281.77         1892.5
         Depreciation                                                   77.21           518.6
         Total Average Cash Cost per t Ore                             358.98         2411.1


IVA is liable to pay a production royalty to the Queensland State Government. The
royalty is calculated from the value of the products sold, less the costs of
transportation to the market and refining costs, times 2.7%. The Queensland
Government does not prescribe a royalty rate for rhenium and there is some potential
for rhenium to be liable for royalty at 2.5% or some other rate determined by the
Queensland Government.

Amendments to the Queensland royalty rate will see copper and other base metals
move to a variable rate based on quoted market prices. Royalty rate for molybdenum
concentrates is 2.7%.

21.2.4    Unit Operating Cost

Operating costs per lb of molybdenum produced are given in Table 21.7.




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Table 21.7         Operating Costs per lb of Molybdenum
                                          Area                             A$ per lb Mo
                 Mining                                                           8.44
                 Ore Haulage to Osborne                                           0.27
                 Concentrator (50% cost shared w ith Cu/Au Ops)                   1.53
                 Concentrate Treatment (Osborne)                                  1.38
                 Site Administration                                              0.57
                 Realisation                                                      0.29
                 Rehabilitation                                                   0.13
                 Average Cash Operating Cost per lb Mo                           12.61
                 Royalty                                                          0.90
                 Average Cash Cost per lb Mo                                     13.51
                 Depreciation                                                     3.70
                 Total Average Cash Cost per lb Mo                               17.21
     * Excludes by-product credits

Table 21.8 gives the Average Cash Operating cost per lb Mo and the Average Cash
Cost per lb Mo including by-product credits. By-product credits for rhenium, copper
and silver amount to the equivalent of US$ 5.13 per lb to the average cash cost of
molybdenum.

Table 21.8         Average Cash Cost for Molybdenum
                                                                  By-products
       Average Operating in US$ per lb Mo              Mo                                 Total
                                                                  (Re, Cu, Ag)
    Average Cash Operating Cost per lb Mo            10.57             (4.53)               6.04
    Royalty                                           0.62              0.13                0.75
    Average Cash Cost per lb Mo                      11.19             (4.41)               6.78




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22     ECONOMIC ANALYSIS

22.1      Introduction

The financial model was developed in-house by IVA. AMC reviewed the model logic,
consistency of input assumptions and integrity of the calculations.

All costs are constant in 2010 Australian dollars (A$), as established at the start of
the PFS, with no provision for inflation escalation. The annual cash flow projections
were estimated over the life of the mine based on capital expenditures, production
costs, corporate costs and sales revenue. The financial indicators examined for each
option of the project include after-tax cash flow, net present value (NPV), internal rate
of return (IRR) and payback period.

This section incorporates a number of changes to the project schedules, metal prices
and exchange rates that have been adopted by IVA since the draft PFS report was
completed. AMC is satisfied that these changes reflect the most likely scenario for
project development. It will be necessary, prior to initiating a final Feasibility Study, to
examine the effects of the changed assumptions on the mining and processing
schedules and costs.

The following sections identify the assumptions which form the basis for the
economic analysis.

22.2      Metal Sale Prices

Sale prices of metals are based on a Roskill independent report on the outlook for
molybdenum and rhenium markets which was commissioned by IVA (Roskill 2011).
These prices are non escalated real prices. Sensitivity analysis, which has been
prepared for metal sale prices above and below Roskill’s forecasts, demonstrates
financial returns at a wide range of molybdenum and rhenium prices.

22.3      Exchange Rates

Figure 22.1 shows the historical exchange rates for A$/US$ rates. The Australian
dollar is currently (September 2011) at a high against the US$ due to the weakness
in the US economy and the strength of the minerals sector in Australia driven by the
strong demand for iron ore and coal exported to China.

The financial model for The Project uses A$/US$ 1.00 for the first three years (ie
2011-2013) and 0.83 thereafter. (The implied average A$/US$ (weighted by revenue)
over the Life of Mine is approximately 0.84.)

.




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Figure 22.1                         A$ / US$ Exchange Rate


                                           Daily AUDUSD Exchange Rate
                           1.20

                           1.00
    AUDUSD Exchange Rate




                                                                      5-year
                                   Average: 0.74
                           0.80

                           0.60

                           0.40

                           0.20

                           0.00




22.4                          Taxes

Australian corporate tax is payable by IVA at the company level, not the project level.
The assumption is that IVA and any new legal entities created for the purpose of The
Project will be consolidated for taxation purposes. The following assumptions have
been applied when calculating corporate tax payable.
•                          30% corporate tax rate (for the purpose of the economic model, this is applied
                           to the Project cash flows only (ie the impact of other IVA company cash flows
                           has been ignored)).
•                          Tax payable yearly.

Project assets are depreciated over their useful life, according to Australian Taxation
legislation. A number of other factors will impact the quantum and timing of tax
payable within the Ivanhoe Australia Consolidation group.
•                          As at 31 December 2010, the IVA group has carried forward tax losses of
                           approximately A$204 M. The IVA group’s tax losses were not included in the
                           economic analysis of The Project and as such represent potential upside to the
                           economic evaluation of The Project.
•                          IVA will generate further tax losses pre-production from The Project and
                           continued exploration activities during 2011. (Project tax losses have been
                           factored into the calculation of tax payable in the economic model. No forecast
                           of future tax losses relating to other IVA projects or corporate activities is made
                           in the economic model).




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22.5      Goods and Services Tax

A Goods and Services Tax (GST) at a rate of 10% is levied by the Federal
Government on purchases by individuals and corporations on non-exempt goods and
services. Business can claim back GST on most business inputs. It is assumed that
most of the product sales will be to overseas customers, so no GST is applicable.
For sales to customers in Australia, GST will be applicable, but would have minimal
impact on the timing of the project cash flows.

22.6      Carbon Trading Scheme

The Australian Federal Government announced its Policy for a Carbon Tax on 10
July 2011.

The majority of the work for the Project PFS was undertaken prior to this
announcement, and hence the impact of a carbon tax has not been factored into the
economic model or the PFS.

22.7      Royalties

22.7.1    Queensland State Royalty

Queensland State Mineral legislation imposes a royalty on the sale of minerals. The
royalty rates applicable to sales of molybdenum and rhenium are fixed at 2.7% and
2.5% respectively. Effective from 1 January 2011, the royalty rate applicable to
copper sales is variable between 2.5% and 5.0%, varying in 0.02% increments, and
depending on average metal prices.

The financial analysis of The Project has applied royalty rates of 2.7% to sales of
molybdenum, 2.5% to sales of rhenium and 4.7% to copper sales. AMC notes that
based on the US$2.00/lb Cu price assumption used in modelling, the royalty rate
applicable to copper sales is around 3.0% and thus the analysis is conservative.

The Queensland Department of Minerals and Energy offers a discount on the royalty
rates applicable to sales of molybdenum and other prescribed metals as stated in the
Minerals Resources Regulation 2003 where, as a result of further processing
undertaken in Australia, the product for sale reaches prescribed grades. A 20%
discount may be applicable to sales of molybdenum from The Project as a result of
the roasting which is planned to take place in Australia. (Note that no such discount
has been included in the economic analysis of The Project and as such represents
potential upside to the economic evaluation of The Project.).

22.7.2    Previous Owner Royalty

There is no previous owner royalty or commitments to third parties based on The
Project mineral sales.




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22.8      Native Title Compensation

IVA currently pays compensation on a yearly basis to the local communities. For
Exploration tenements and Mineral Leases, the claimants are paid a fixed amount,
which is CPI increased for subsequent years. This amount, although currently small
and immaterial is included in administration costs for The Project economic model.

If IVA changes the size or shape of the mineral lease or applies for more surface
rights than originally granted, then there is a need to negotiate new agreements. If
this were the case then the current compensation calculation may change.

22.9      Other Royalties/Agreements

Burke River Bore field compensation is a flat fee, plus Consumer Price index (CPI)
increases, and is included in the administration costs estimated for The Project.

22.10     Revenue Deductions

Shipping and logistical cost for molybdenum trioxide and APR are based on the
products being roasted on site and transported to Townsville. The cost of roasting
the molybdenum and the flue capture of the rhenium are incorporated into the
processing costs, while transport is a revenue deduction for the purpose of
determining Queensland State royalties payable. No storage costs have been
estimated for the rhenium if it is decided to store and sell in future periods. (Note that
rhenium volumes are low relative to other products and given the density of the
metal, will require little space for storage).

22.11     Reclamation

A total amount of A$18 million has been allocated for reclamation and rehabilitation
from 2017 onwards.

22.12     Project Financing

No assumptions have been made for project financing in the economic model.

22.13     Assumptions

Table 22.1 gives the key assumptions used in the financial model.




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Table 22.1       Key Assumptions
                      Commodity Prices
                      Molybdenum price                      US$/lb          Profiled
                                                                             price
                      Rhenium price                         US$/kg          Profiled
                                                                             price
                      Copper price                          US$/lb           4.00
                      Gold price                            US$/oz           1,350
                      Silver price                          US$/oz           20.00
                      Exchange rates
                      AUD:USD (2011)                        A$:US$           1.06
                      AUD:USD (2012 – 2013)                 A$:US$           1.00
                      AUD:USD (2014 onw ards)               A$:US$           0.83
                      Inflation
                      Price inflation                           % pa           0
                      A$ cost inflation                         % pa           0
                      US$ cost inflation                        % pa           0



Roskill forecasts molybdenum oxide prices to increase from an average price of
US$16.50 /lb for 2011 to US$27.00 /lb (US$23.77 /lb in real terms) in 2016. Roskill’s
forecast for rhenium metal is US$4,350 /kg in 2011 increasing to US$5,700 /kg
(US$5,018 /kg in real terms) in 2016.

The economic model is based on Roskill forecast real prices for molybdenum and
rhenium from 2011 to 2016 inclusive. IVA has held 2016 prices constant thereafter
and has selected an AUDUSD exchange rate of 1.00 for 2012 - 2013 and 0.83
thereafter. Further information regarding the outlook for the molybdenum and
rhenium markets is provided in Section 19.1 of this report.

22.14      Cash Operating Costs

Table 22.2 shows the average operating cost over the mine life in US$/lb of
molybdenum and the value of the by-products.

Table 22.2       Average Cash Operating Costs
                                                                       By-Products (Re,
                                                    Mo                                      Total
                                                                           Cu, Ag)
 Average Cash Cost                        US$/lb      10.57                        (4.53)      6.04
 Royalty                                  US$/lb         0.62                       0.13       0.75
 Average Cash Cost                        US$/lb      11.19                        (4.41)      6.78




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22.15                         Revenue

Figure 22.2 shows the sales revenue from molybdenum and by-products over time.

Figure 22.2                               Value of Sales Revenues


                                                                Sales revenue by metal
                         450.0
                         400.0
                         350.0
   Sales revenue (A$m)




                         300.0                                                                                                                                                       Silver
                         250.0                                                                                                                                                       Gold
                         200.0                                                                                                                                                       Copper
                         150.0
                                                                                                                                                                                     Rhenium
                         100.0
                                                                                                                                                                                     Molybdenum
                            50.0
                             0.0
                                    2012
                                           2013
                                                  2014
                                                         2015
                                                                2016
                                                                       2017
                                                                              2018
                                                                                     2019
                                                                                            2020
                                                                                                   2021
                                                                                                          2022
                                                                                                                 2023
                                                                                                                            2024
                                                                                                                                                      2025
                                                                                                                                                             2026
                                                                                                                                                                    2027
Figure 22.3 shows the production of molybdenum, including by-product revenues                                                                                              2028
and operating costs over time.

Figure 22.3                               Production vs. Unit Operating Cost


                                           Production vs. unit operating cost*
                            14.0                                                                          35.00
                                                                                                                    Unit operating cost (US$/lb Mo)




                            12.0                                                                          30.00
        Production (Mlbs)




                            10.0                                                                          25.00
                             8.0                                                                          20.00
                                                                                                                                                                    Molybdenum production
                             6.0                                                                          15.00
                                                                                                                                                                    Unit operating cost
                             4.0                                                                          10.00
                             2.0                                                                          5.00
                                                                                                                                                                              * operating costs
                             0.0                                                                          0.00                                                                are net of by-
                                                                                                                                                                              product credits &
                                   2012
                                           2014
                                                  2016
                                                         2018
                                                                 2020
                                                                         2022
                                                                                2024
                                                                                        2026
                                                                                                   2028




                                                                                                                                                                              royalties




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22.16                             Cash Flow

Figure 22.4 shows the after-tax cash flow.

Figure 22.4                              After Tax Cash Flow


                                                             After-tax cash flow*
                               2000.0


                               1500.0
   After-tax cash flow (A$M)




                                                                                                             After-tax cash flow
                               1000.0

                                                                                                             Cumulative after-tax cash
                                500.0                                                                        flow

                                  0.0
                                        2011

                                               2013

                                                      2015

                                                             2017

                                                                    2019

                                                                           2021

                                                                                  2023

                                                                                         2025

                                                                                                2027


                                                                                                                    * pre-financing
                               -500.0


22.17                             NPV, IRR and Payback Period

The NPV Sensitivity to Discount Rate for The Project is shown in Table 22.3.

Table 22.3                               NPV Sensitivity to Discount Rate
                                                              Discount Rate (%)            NPV (A$M)
                                                                           5                           935
                                                                           8                           690
                                                                           10                          562
                                                                           12                          457
                                                                           15                          332


The Internal Rate of Return (IRR) is 31.8%.
The Payback Period (from the date of first production) is approximately four years.

22.18                             Net Present Value (NPV) Sensitivity

Figure 22.5 shows the NPV sensitivity at +/- 10%




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Figure 22.5       NPV Sensitivity at +/-10%


                         NPV sensitivities at +/- 10%*
             AUDUSD exchange rate (±10%)

                  Molybdenum price (±10%)

              Molybdenum recovery (±10%)

                 Molybdenum grade (±10%)

                     Operating costs (±10%)

                      Rhenium price (±10%)

                  Rhenium recovery (±10%)

                     Rhenium grade (±10%)

                        Capital costs (±10%)

 -200.0     -150.0    -100.0      -50.0        0.0    50.0       100.0      150.0       200.0      250.0
                                          Change in NPV (A$M)
                                               -10%   +10%               * Based on NPV of A$690M




Table 22.4 shows the NPV at various metal prices.

Table 22.4        NPV at Various Metal Prices
                                                               Rhenium Price (US$/kg)
          Molybdenum price (US$/lb)                          4,300          5,000          7,500
                               25.00           A$M            804            834            944
                               22.00           A$M            616            647            756
                               20.00           A$M            492            522            632
                               18.00           A$M            366            397            507
                               16.00           A$M            241            272            382



22.19      After-tax Cash Flow (ATCF) Sensitivity

Figure 22.6 shows the ATCF sensitivity at +/- 10%.




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Figure 22.6       ATCF Sensitivity at +/-10%


            After-tax cash flow sensitivities at +/- 10%*
            AUDUSD exchange rate (±10%)

                 Molybdenum price (±10%)

             Molybdenum recovery (±10%)

                Molybdenum grade (±10%)

                    Operating costs (±10%)

                     Rhenium price (±10%)

                 Rhenium recovery (±10%)

                    Rhenium grade (±10%)

                         Capital costs (±10%)

 -300.0         -200.0        -100.0            0.0      100.0          200.0         300.0           400.0
                                            Change in ATCF              * Based on after-tax cash flow
                                                  -10%   +10%           of A$1,563M


Table 22.5 shows the After Tax Cash Flow at various metal prices.

Table 22.5        ATCF at Various Metal Prices
                                                                 Rhenium Price (US$/kg)
          Molybdenum price (US$/lb)                          4,300          5,000             7,500
                               25.00             A$M         1,732          1,790             1,996
                               22.00             A$M         1,388          1,446             1,652
                               20.00             A$M         1,159          1,217             1,422
                               18.00             A$M            928             986           1,193
                               16.00             A$M            696             754            963




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23     ADJACENT PROPERTIES

The Merlin Mo-Re resources lie within the Mount Dore North lower Cu-Zn sulphide
mineralisation. This is below the Mount Dore upper oxide mineralisation that is
subject to a separate preliminary economic analysis based on open pit mining and
heap leach extraction of copper.

The Merlin and Mount Dore deposits are located in the Kuridala formation and is
overlain uncomformably by largely barren granite. The deposit lies below the
southernmost contact with this granite body. Exploration has identified anomalous
copper mineralisation to the North of Mount Dore extending along the margin of the
thrusted granite for over 8 km. Lady Ella is a copper deposit previously mined by
Selwyn in 1990’s. It lies in the same structural and stratigraphic position as Mount
Dore but resides on the northernmost contact of the same granite body.

Further to the north lies Mount Elliot which was previously mined and is another
focus for IVA’s exploration and definition work.

Other nearby deposits mined by Selwyn in the 1990’s reside on the Starra line of
mineralisation 1 km to the West. These bear no direct relationship to the Mount Dore
mineralisation but their close proximity was the cause of the remnant infrastructure
including the existing ex-mine camp, mill footings and tailings facilities.




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24    OTHER RELEVANT DATA AND INFORMATION

24.1 Pre-feasibility Study Optimisation Opportunities

The PFS has identified a number of optimisation opportunities that will be
investigated during any feasibility study (FS) of the Merlin project.

The key opportunities are briefly outlined below.

Mining cost:
•     The PFS has assumed 15 m sublevels in the underground mine. As a
      consequence of further evaluation following completion of the PFS these have
      now been modified to 20 m sublevels which will reduce the mining cost. Any FS
      will be completed on this basis.
•     The PFS focused attention on traditional mining methods. It has been identified
      that innovative mechanical mining may have potential at Merlin and thus will be
      included in any FS scope. A mix of long hole open stoping and drift and fill
      stoping will also be considered as part of the optimisation of mining methods.
•     Sensitivity analysis which has been undertaken on the Merlin project PFS
      indicates that a 10% reduction in mining costs is likely to improve NPV and
      after-tax cash flow by approximately A$47 M and A$84 M respectively.

Molybdenum concentrate grade:
•     In this report, for design purposes, it has been assumed that all of the Mo/Re
      concentrate grade will be 30% Mo, however testwork has shown that two
      concentrate grades at 50% Mo and 30% Mo can be produced. Further work will
      be completed during the early stages of the FS to optimise the product
      streams.

Capital cost:
•     The molybdenum/rhenium sulphide concentrate produced will be processed
      through a       purpose-built, conventional technology,        roasting and
      hydrometallurgical plant to produce saleable products including molybdenum
      trioxide and ammonium perrhenate (APR). The sale of the Mo/Re concentrate,
      the optimisation of the process design to minimise capital cost as well as the
      potential to produce other Mo products including ferro-molybdenum (FeMo),
      ammonium di-molybdenate and sulphuric acid, will be investigated in the early
      stages of the FS. This has the potential to reduce the capital cost of the
      processing plant.
•     This PFS assumed that the Mo/Re roaster and hydrometallurgical plant is
      located at Osborne. Alternative locations in other countries will be evaluated in
      the early stages of the FS to potentially reduce the capital expenditure and
      optimise the return on investment prior to commencing the FS.
      Sensitivity analysis which has been undertaken on the Merlin project PFS
      indicates that a 10% reduction in overall capital costs is likely to improve NPV
      and after-tax cash flow by approximately A$31 M and A$35 M respectively.




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Life extension:
•     Exploration within the existing tenements where there have been high grade
      molybdenum finds has the potential to extend the life of the Mo/Re project.




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25     INTERPRETATION AND CONCLUSIONS

25.1      Geology

This section is unchanged from the Technical Report prepared by Golder in
October 2010 (Golder, 2010) and reproduced here with some exclusions of Merlin
specific information which is not relevant to the Mount Dore copper oxide resource.

25.1.1    Merlin - Little Wizard Mo-Re Deposit

The Merlin Zone Mo-Re deposit lies within the northern half of the Mount Dore
project area, Cloncurry district, Queensland, Australia. The high-grade hydrothermal
Mo-Re mineralisation is a recent discovery that has been subject to a substantial
exploration program by Ivanhoe Australia Limited. To June 30, 2010, 143 holes
intersect the Merlin and Little Wizard domains, up from 76 used in the last Merlin
estimate QG (2009). Though QG did update the small Little Wizard deposit in
February 2010, significant drilling has been completed at Little Wizard and Merlin
that is material to the Mineral Resource.

Molybdenum-rhenium (Mo-Re) and copper (Cu) mineralisation at Mount Dore is
hosted within a tectonised sequence of metashale, metasiltstone, schist and phyllites
belonging to the Proterozoic Kuridala Formation in the Eastern Fold Belt of the
Mount Isa Inlier.

Fracture-controlled and breccia-matrix molybdenite mineralisation is hosted within K-
feldspar-altered and albitised black metashales and siltstones, which lie above and
below the foliated schist and phyllite. The footwall structure at the base of the foliated
phyllite and schist appears to have the strongest Mo and is inferred to have
developed good open structures due to competency contrast. This basal contact also
appears to have acted as a significant barrier for the Mo-rich fluids, resulting in
ponding of the metals in the favourable structures just below the contact. The
mineralised matrix-breccias contain sub-rounded clasts of K-feldspar and clay-
altered siltstone with very minor clay, with molybdenite partially to completely
replacing the breccia matrix.

Exploration of the Merlin deposit has primarily been diamond drilling with lesser RC
drilling. IVA has implemented a rigorous set of protocols for logging, sampling,
assaying and quality control. These protocols have been effective in ensuring data
quality is not compromised throughout the exploration programs.

The spatial distribution of Mo-Re grades in the Merlin zone displays an abrupt
boundary often marked by the occurrence of massive molybdenite over short
intervals. Such sharp boundaries require hard domain boundaries to control grade
estimation and smooth of high and low grades.

The high grade domain boundary was wireframed to constrain the high grade Mo
mineralisation as a continuous narrow planar body with some lower splay structures.
The broad orientation of the mineralised domains is parallel with the lithological units
and structures. The available drilling remains relatively wide requiring significant
interpretation that the modelled vein is continuous without small scale disruption.




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25.1.2    Resource Estimate

A block model was constructed to estimate grade and for mine planning purposes.
The principal block size used for the ordinary kriged estimates was varied from 10 x
25 x 10 m block size for the upper Mount Dore copper zones deemed suitable for
open pit assessment and a smaller 5 x 12.5 x 5 m block size for lower and Merlin
underground mining assessment.

Classification of resources was based on the target drill spacing of 50 m for Indicated
Mineral Resource, which is now largely complete for the Merlin deposit.

The Merlin and Little Wizard total Mineral Resource estimate at a 0.3% Mo cut-off is:
                 6.5 Mt at 1.3% Mo and 23 ppm Re Indicated Mineral Resource
                 0.2 Mt at 0.9% Mo and 15 ppm Re Inferred Mineral Resource

The Mount Dore total Mineral Resource estimate at a 0.25% Cu cut-off is:
                 86 Mt at 0.55% Cu Indicated Mineral Resource
                 58 Mt at 0.47% Cu Inferred Mineral Resource

Significant additional Zn mineralisation exists outside of the quoted Cu mineralisation
at Mount Dore North. Viable recovery of Zn is not yet demonstrated and stand alone
Zn resources are not included in this resource statement.

25.2      Geotechnical, Mining and Mineral Reserve

A decline-access underground mine is most suitable. The drift and fill (DAF) and
longhole open stoping (LHOS) underground mining methods were selected as most
suitable. The mining schedule was prepared assuming a top-down mining system is
employed in all mining areas, as follows:
•      DAF mining assuming 40 m/mth average advance (including backfill) rate per
       available ore heading, based on a single level, centre-out sequence.
•      Longitudinal LHOS assuming 420 t/d stoping rate (including backfill, excluding
       development) based on a single level, 2 stope end-retreat sequence.
•      Transverse LHOS assuming 600 t/d stoping rate (including backfill, excluding
       development), based on a single level, 2 stope centre-out sequence.

The resultant production schedule shows that ore is sourced from 60% DAF mining,
8% development, and 32% LHOS. All mining blocks are opened up quickly and
production reaches 150 kt/a in 2012, 300 kt/a in 2013 before reaching full production
of about 500 kt/a in 2014. The peak rate of 500 kt/a or above is maintained for about
ten years and then falls away rapidly thereafter.

After rounding, The Project mineral reserve estimate (including Little Wizard) is:
           6.7 Mt at 1.1% Mo and 19.1 ppm Re Probable Mineral Reserve




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25.3      Processing

The required plant design includes:

At Merlin mine site
•      Run of Mine (ROM) pad, crushing facilities and ore handling facilities.
•      Fine crushed ore transport to Osborne.

At Osborne
•      Crushed ore delivery and stockpiling on existing ROM pad.
•      Ore reclaim and conveying to a new ball mill.
•      New flotation, concentrate filtering, storage and load-out.
•      New concentrate treatment plant.

Suitable flowsheets have been developed. Key process plant parameters are
outlined in Table 25.1.
Table 25.1        Key Process Plant Parameters
                   Key Design Parameters                    Value        Unit

         Total ROM ore feed                                500,000         dt/a
         Molybdenum grade, ROM ore                           1.2          wt%
         Molybdenum grade, concentrate                       30           wt%
         Molybdenum production                              5,069           t/a
         Overall molybdenum recovery                         84.5           %
         Rhenium production                                 7.277          dt/a
         Overall rhenium recovery                            80.9           %



25.4      Infrastructure

Infrastructure requirements for the Mount Dore/Merlin mine site area and for the
Osborne site include power, water supply, aerodrome, accommodation villages,
communications and tailings storages. Suitable arrangements have been identified
for each of these items and can be further developed in a Feasibility Study.

25.5      Risks

Significant risks and uncertainties that have been identified include:
•      Delays in roaster approval through the environmental impact statement
       process.
•      Delays in mine development or production.
•      Failure or poor performance of the proposed underhand mining methods.
•      Inability to find a suitable roaster location.
•      Management of mine water.




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•     Inability to secure buyer or off take for molybdenum.
•     Mineral grade incorrectly modelled, reducing mineral reserve.
•     Capital cost or operating cost underestimated.
•     Adverse movements in foreign exchange rates.
•     Uranium content in ore results in reduced product marketability.
•     Inability to secure sufficient water supply and/or water quality.
•     Decrease in molybdenum or rhenium prices.
•     Insufficient human resources to manage multiple projects.
•     The effect of cement from shotcrete on mill recovery.
•     The effect of poor ground conditions on mining.

Further studies are required to resolve these uncertainties.




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26     RECOMMENDATIONS

AMC recommends that, subject to a decision by IVA to proceed, the next phase of
work should be data acquisition and studies leading to completion of a Feasibility
Study. AMC notes that underground development is already under way.

From AMC’s own data, the cost of a Feasibility Study for an underground mine
(including processing plant) at this scale can range from 2% to 12% (average 6%) of
the project capital cost. Given the work already completed and the existing shared
facilities, AMC expects that the cost will be at the lower end of this range. On the
estimated capital cost of A$518.6 M for this project AMC estimates the Feasibility
Study cost, including fieldwork, consultants, owners’ costs and permitting, to be
around $10 M. AMC considers this estimate to be consistent with an element of the
Indirect Costs listed in Table 21.3 although the itemisation of costs is different.

The following recommendations are made for some of the work areas for a
Feasibility Study. The recommendations are not intended to set out a Scope of Work
for a Feasibility Study. The costs of the recommended work would be included in the
Feasibility Study cost.

26.1      Geological

This section is unchanged from the Technical Report prepared by Golder in October
2010 (Golder, 2010) and is here reproduced in its entirety for the convenience of the
reader.

Resource definition work is a process of building upon existing knowledge and at the
same time rebuilding from a new starting position. This is a progressive process and
there usually remains some area where more work is required. Excluding the next
logical step of infilling the current resource drilling to bring the resource to a higher
level of confidence, for exploration and resource definition the following areas are
identified as requiring further work:
•      Continue the drill hole relogging process such that the lithological model
       can be more tightly defined.
•      Undertake a more detailed audit of the drilling undertaken by previous
       companies and the early IVA drilling pre 2007.
•      Review the occurrence and grade of the wet RC samples at Mount Dore
       South and determine if corrective action is required.
•      Make better use of the orientated core results to assist the interpretation
       of the narrow mineralisation shapes.
•      Further assess the apparent change between the northern and southern
       areas.
•      Improve the understanding of the weathering profile on potential metal
       recovery.

Additional exploration program drilling should be targeted at The Project source
which is the focus of the current development plan. At this stage no Measured




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Mineral Resources are defined and the current 50 m spaced drilling is at the lower
end of the Indicated Mineral Resource confidence range. Additional drilling is
required to define:
•      The close range variability and continuity of the Merlin veins both for grade
       estimation and for mine planning purposes. This should be initiated with a
       small program of targeted surface drill holes followed by both exploration
       development and sampling and close spaced underground drilling in a few key
       areas.
•      The general drill spacing should be closed down to 25 m. Given that
       underground development has been commenced, infill drilling would be most
       effective from underground both in terms of cost and possibly intersection
       angle.

26.2      Geotechnical

Work in subsequent studies should include:
•      Further assessment of major geological structures, discontinuity orientations
       and weathering profile.
•      Stress modelling of mine openings and backfill.
•      Development of the ground control management plan.
•      Guide on geotechnical requirements, systems and monitoring program for the
       implementation of the mine.
•      An outline of further geotechnical work required for the implementation phase
       of the project.
•      Further assessment of geotechnical risks to the project.

In order to produce a study to feasibility level the following information is required, to
an appropriate standard; this would be gained from a targeted geotechnical drilling
programme:
•      Geotechnical and geological logging data and down hole surveys from the
       geological drilling programmes with documentation of quality control process.
•      A focus on gathering structural logging information form oriented core.
•      All core photos for both geological and geotechnical drilling programmes.
•      Geotechnical laboratory test data.
•      Geological interpretations of lithology / alteration / faults / shears as
       wireframes.
•      Any hydrogeological studies or information.
•      Site specific in-situ stress information.

Work on these matters is ongoing.




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26.3      Mining

Further key areas of work to be completed as a basis for further mining studies
include:
•      Geotechnical work as described above.
•      Revision of metallurgical recovery parameters and associated cost and
       financial assumptions affecting the mining value model.
•      Completion of appropriate levels of open pit and associated studies, to confirm
       that underground mining remains the most attractive option.
•      Completion of suitable test work to optimise the paste fill.
•      Revision of optimisation studies using PFS mining and metallurgical
       information to determine the appropriate cut-offs to apply to the mine design
       and schedules.

The mining component of the Merlin Feasibility Study should include the following
mining components:
•      Advanced geotechnical review, assessment and modelling.
•      Mine design and scheduling.
•      Mine services and infrastructure planning.
•      Mine ventilation planning.
•      Backfill planning and preliminary system design.
•      Mine equipment and organisational planning.
•      Operational review and benchmarking.
•      Re-estimation of Mineral Reserves.

26.4      Processing

A review of the process highlighted several areas for refinement of the process.
These are discussed below, and should be pursued in the next phase of the project.

26.4.1    Concentrate Production

•      Recent test work has revealed that the level of graphite in the ore appears to
       be a good indicator as to whether a high grade concentrate can be achieved.
       Graphite levels below 0.1% consistently gave saleable grade concentrate in the
       variability flotation tests. There does not appear to be a strong correlation to the
       total carbon and concentrate grade. The high levels of graphite appear to be
       isolated to the black shale ore lithology. A review of the black shale lithology
       should address whether ore can be produced with little or no black shale
       content, in which case the concentrate grade could be much higher than the
       estimated 30% molybdenum used in this study. This may open the opportunity
       to either sell the concentrate prior to roasting, or reduce the requirement for the
       calcine leach circuit post roasting.




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•     Regrind testwork was successful in allowing depression of the copper
      sulphides from the molybdenum concentrate. A study should address whether
      the ferric chloride leach stage can be deleted.
•     A study should address whether, instead of cleaning the copper concentrate to
      increase the grade, the lower grade concentrate can be fed directly into the
      existing copper flotation circuit for upgrading and blending to final concentrate.
      There may be a need to oxidise the residual sodium hydrosulphide (NaHS)
      prior to the transfer. If so, this would be done with the addition of peroxide to a
      mixing tank.
•     A review of the mechanical flotation cells is required to ensure that additional
      retention time is allowed for to ensure expected recoveries are achieved in
      periods of higher grade feed.
•     Allowance should be made for lime addition to the mill feed belt. This will allow
      for pH control of the rougher feed slurry. Sometimes this is required to improve
      froth characteristics. A higher pH may also be required if the feed has a lot of
      copper as this will help copper recovery. pH control may also be required if any
      of the ore is ‘acidic’ to avoid corrosion issues.

26.4.2    Concentrate Treatment

•     The Ion Exchange (IX) feed will be oxidised with peroxide. This is to oxidise
      any selenium which might be present and would otherwise precipitate on the
      resin and foul it. It is unlikely that there will be any selenium present and the
      requirement for this step should be reviewed in the next phase of work.
•     The molybdenum leach is fast and may not need the 5 tanks specified in this
      study. A review is required.
•     Collecting the molybdenum leach residue as a filter cake rather than repulping
      should be considered. If repulping is required, using the neutralised raffinate
      would be more suitable than fresh water.
•     Anhydrous ammonia should be considered to replace aqua ammonia due to
      the large quantity used in an effort to reduce operating costs.

26.4.3    Crushing Facility

•     A study should be undertaken to compare a new plant against a contracted
      crushing facility. This would reduce the capital cost by approximately A$8.5 M.

26.4.4    Location of the Concentrate Treatment Facility

Alternate locations to site the molybdenum rhenium roasting and hydrometallurgical
plant to produce molybdenum trioxide and ammonium perrhenate for sale from the
sulphide concentrates produced at Osborne were investigated.

Three potential locations were investigated:
1.    In the heavy industrial designated zone in Townsville, North Queensland
      adjacent to the Port of Townsville.




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2.    An overseas location to take advantage of lower capital and operating costs.
      Joint ventures with Korean companies located in Korea were identified along
      with other potential locations in Asia.
3.    At the Osborne mine site.

In the PFS, the location of the roaster was selected to be at the Osborne site, but the
option for an overseas site is still possible. The final decision should be made
following further investigation and studies during the Feasibility Study.

26.4.5    Sulphur Dioxide Roaster Gas Cleaning

The initial roaster flowsheet developed by Jacobs was based on scrubbing the SO2
from the exit gas stream using a lime scrubber. This produced a gypsum product and
an emission level of 320 t/a SO2. To comply with perceived maximum SO2 emission
levels for Queensland (there is no legislated emission level) it was questionable
whether the lime scrubber would meet DERM requirements. Alternative scrubbers
were evaluated. These include:
•     Lime/caustic soda scrubber, caustic soda or sodium carbonate neutralising
      SO2 gas scrubber.
•     Haldor Topsoe wet sulphuric acid (WSA) plant.

A conventional sulphuric acid plant was rejected early on in the evaluation process
because of the high capital cost (> A$50 M) and inefficiency due to the low
concentration of SO2 (2%) in the gas stream.

The final plant configuration adopted for the PFS was for a Haldor Topsoe wet
sulphuric acid (WSA) plant.

The final decision over these two alternatives should be made in the
Feasibility Study.

26.4.6    Sell or Toll Treat Molybdenum Concentrates

The option studied was to export the sulphide concentrate and sell or toll treat the
concentrates at an existing facility. This option avoids the construction of a roaster
plant in a remote location in Australia.

The capital cost reduction for not constructing a roasting plant is approximately
A$108 M.

A financial analysis compared the base case with a molybdenum roaster at Osborne
with the expected return of metal and refining cost based on the value of the rhenium
in the concentrate. The analysis showed a substantial reduction in NPV if the
molybdenum concentrate is sold or toll treated instead of constructing a purpose built
roaster.

It was concluded that a purpose built roaster and rhenium recovery plant had a
significantly higher NPV compared to selling concentrate and was therefore essential




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to extracting the full value of the mineral resource. This should be confirmed in the
Feasibility Study.




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27    REFERENCES

Aker Solutions, 2010. Ivanhoe Australia, Osborne upgrade project, Pre-feasibility
study, Options study report, Report Number 11402-00-G0713 Rev P1, October 2010.

Aker Solutions, 2010. Osborne Upgrade Project Pre feasibility study Ivanhoe
Australia Hazard identification report, Report Number 11402-00-K0704 Rev P1,
October 2010.

Aker Solutions, 2010. Osborne Upgrade Project Pre feasibility study Ivanhoe
Australia Hazard identification report, Report Number 11402-00-K0811 Rev P1,
September 2010.

Aker Solutions, 2011. Ivanhoe Australia Osborne Upgrade Project – Process Design
Criteria – Concentrator Plant. Report Number 11402-02-B8701 Rev P1, March 2011.

Aker Solutions, 2011. Ivanhoe Australia Osborne Upgrade Project – Process Design
Criteria – Crushing Plant. Report Number 11402-01-B8701 Rev P1, March 2011.

Aker Solutions, 2011. Ivanhoe Australia Osborne Upgrade Project – Process Design
Criteria – Roaster Plant. Report Number 11402-03-B8701 Rev P1, March 2011.

Aker Solutions, 2011. Ivanhoe Australia Osborne Upgrade Project – Summary,
11402-00-J0007RevP1, March 2011.

Aker Solutions, 2011. Osborne Upgrade Project Process Plant Engineering Study.
Section 3 Process Plant. Report Number 11402-00-G0712 Rev P3, February 2011.

Aker Solutions, 2011. Osborne Upgrade Project, EPC Schedule - Overall Project By
Area, 11402-00-J0002 Rev P1, March 2011.

Aker Solutions, 2011. Process Design Basis. Report number 11402-00-B0702 Rev
P4, February 2011.

AMC Consultants Pty Ltd, 2010. Merlin PFS Geotechnical Assessment, Ivanhoe
Australia Ltd, Project Number 110031, September 2010.

AMC Consultants Pty Ltd, 2010. Merlin Underground Mine Pre-feasibility Mining
Study. Ivanhoe Australia Ltd. Project Number 110031, November 2010.

AMMTEC, 2009. Burnie Research Laboratories, 2009. Flotation Testing of Merlin
Molybdenite Ores – T 0457-1& T 0457-2.

EHP Consulting, 2009. Roaster Design Criteria, IAL-Roast-Pressure Leach Study,
IAL-EHP-March 2009-A, 15 June 2009.

Golder Associates, 2010. NI43-101 Technical Report. Mount Dore – Merlin Deposit,
NWQld, Australia. Report number 107631016-011-Rev0, Submitted 19 Oct 2010.




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Golder, 2010. Mount Dore - Merlin Deposit, NWQld, Australia, Technical Report
compiled under NI43-101 by J Horton, 27 June 2010.

Hillier, K., 2009. Preliminary Assessment of Bulk Density Measurements. IVA internal
file note dated December 2009.

Jacobs, 2011. Merlin Molybdenum / Rhenium Pre-Feasibility Study. Report number
11402-00-G0721 Rev P1, March 2011.

JORC Code 2004. Australasian Code for Reporting of Identified Mineral Resources
and Ore Reserves. Report of the Joint Committee of the Australasian Institute of
Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of
Australia (JORC), issued December 2004.

Lodewyk, M., 2010. Report to the Mining Registrar, Mount Isa, S.274. Mineral
Resources Act, Maintenance of Posts, Mining Leases 2566, 2688 to 2694, 2733,
2745 and 2746. M H Lodewyk Pty Ltd, registered surveyor report.

Metago Environmental Engineers (Australia) Pty Ltd, 2011. Tailings and Effluent
Management Evaluations for the Ivanhoe Osborne Project Pre-Feasibility Study.
Project No. 349-001 Report No. 1/11, January 2011.

Metcon Research, 2011. Merlin Project Ivanhoe Merlin Feasibility Phase I (Flotation
Tests on Composite No. 1), Q757-01-028, February 2011.

Metcon Research, 2011. Merlin Project Ivanhoe Merlin Feasibility Phase II (Flotation
Tests on Composite Nos. 1, 2 and 3), Q757-01-028, February 2011.

Metcon Research, 2011. Merlin Project Ivanhoe Merlin Feasibility Phase III (Flotation
Tests on Variable Composites), Q757-01-028, February 2011.

Metso, 2009, Process Design Criteria, June 2009

Parsons Brinckerhoff Australia Pty Limited 2010. Osborne Mine Re-powering Options
Study. Report Number 2158728A_REPT_001_Rev_B, 16 December 2010

Pocock Industrial, Inc, 2011. Sample Characterization & PSA, Flocculant Screening,
Gravity Sedimentation and Pulp Rheology Studies conducted for: KD Engineering &
METCON Research, Aker Solutions Project #11402, January 2011.

Roskill, 2010. Rhenium: Market Outlook to 2015, Eighth Edition, 2010.

Roskill, 2011. Ivanhoe Australia Limited, Study of the Market for Molybdenum, 29
July 2011.

Sketchley, D., 2008a. QAQC Review of SGS Geochem, Townsville, Queensland,
Australia for Ivanhoe Cloncurry Mines Pty. Ltd. Cloncurry, Queensland, Australia,
Ivanhoe Mines report, dated 7 Feb 2008.




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Sketchley, D., 2008b. Cloncurry Technical Note – SG QAQC Issue, Ivanhoe
Australia Limited memorandum, dated 5 July 2008.

Sketchley, D, 2008c. Sample Prep, Analysis & QAQC Procedures, Ivanhoe Australia
Limited report, dated 9 November 2008.

Sketchley, D., 2009. Technical Report on Assay QAQC work at Merlin Zone,
Cloncurry District, Queensland, Australia for Ivanhoe Australia Limited, dated 18 May
2009.

Sketchley, D., 2010a. Sample Prep, Analysis & QAQC Procedures, Ivanhoe Australia
Limited report, dated 31 January 2010.

Sketchley, D., 2010b. Cloncurry QAQC Site Review, Internal Ivanhoe Australia
Limited memorandum, dated 13 February 2010.

Sketchley, D, 2010c. Review of QAQC Work on Updated Merlin Assay Data. Internal
memorandum for Ivanhoe Australia Limited, dated 19 February 2010.

SRK, 2010. Independent technical report on the Merlin project, Qld, SRK consulting
(SRK) report for Ivanhoe Australia Ltd, reference IVA001 NI43-303 report on Merlin
Rev4, dated 5 March 2010. Available at www.ivanhoeaustralia.com/i/pdf/NI43-
101_Merlin_Scoping_Study_Final.pdf




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28    DATE AND SIGNATURE PAGE

Title: NI43-101 Technical Report Merlin Project Prefeasibility Study

Effective Date of Technical Report: 10 October 2011

Effective Date of Mineral Resource Estimate: 4th August 2010 (as released on the
ASX)

Effective Date of the Mineral Reserve Estimate: 10 October 2011

Qualified Person: Peter L McCarthy, BSc (Eng), MGeosc, FAusIMM (CP), SME,
employed by AMC Consultants Pty Ltd as a Principal Consultant was responsible for
1,2,3,4,5, 13, 15, 16, 17 (except 17.2), 18, 19, 20, 21 (except processing
component), 22, 24, 25 (except 25.1), 26 (except 26.1, 26.4), 27




Peter L McCarthy
Principal Consultant

Date: 10 October 2011




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                            CERTIFICATE of AUTHOR – Peter L McCarthy

I, Peter L. McCarth y, BSc (Eng)(Min), MGeosc, FAusIMM (CP) do hereby certify that:

1.    I am Principal Mining Engineer and Director of AMC Consultants Pty Ltd., Level 19, 114 William
      Street, Melbourne, Victoria, 3000, Australia.

2.    I am one of the authors of NI 43-101 Technical Report titled “Merlin Project Prefeasibility Study,
      NW Queensland”, Australia (The Technical Report), Dated, 10 October, 2011
3.    I graduated with a Bachelor of Science (Engineering) in Mining Engineering from the University
      of New South Wales in 1976 and a Master of Geoscience (Mineral Economics) from Macquarie
      University in 1986.
4.    I am a Fellow and Chartered Professional of The Australasian Institute of Mining and Metallurgy.
5.    I have worked as a mining engineer for a total of 35 years since my graduation from university
      and a total of 42 years in the mining industry.
6.    My relevant work e xperience for the purpose of The Technical Report is:
      •      25 years as consulting mining engineer specialising in mine planning, feasibility studies
             and project evaluation for AMC Consultants Pty Ltd.
      •      5 years as a lecturer and senior lecturer in Mining Engineering for the Ballarat College of
             Ad vanced Education.
      •      2 years as project engineer and mine manager for mine development projects.
      •      9 years as a mining engineer and trainee mining engineer for Conzinc Riotinto of Australia
             Limited.
7.    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
      fulfil the requirements to be a Qualified Person for the purposes of NI 43-101.
8.    I am responsible for the preparation of Sections 1, 2, 3, 4, 5, 13, 15, 16, 17 (except 17.2), 18, 19,
      20, 21 (except processing component), 22, 24, 25 (except 25.1), 26 (e xcept 26.1, 26.4) and 27 of
      The Technical Report.
9.    I visited the site on 15 and 16 June 2011.
10.   I have not had any prior involvement with the property that is the subject of The Technical
      Report.
11.   I am independent of the issuer applying all of the tests in section 1.4 of National Instrument 43-
      101.
12.   I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been
      prepared in compliance with that instrument and form.
13.   As of the date of this certificate, to the best of my information, knowledge and belief, the
      Technical Report contains all scientific and technical information that is required to be disclosed
      to make the Technical Report not misleading.

Dated this 27 Day of October 2011




Original signed and sealed by:
Peter L McCarthy BSc (Eng)(Min), MGeosc, FAusIMM (CP).




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                               CERTIFICATE of AUTHOR – John Horton

I, John Horton, of Brisbane, Australia, do hereby certify that as the author of “Merlin Project
Prefeasibility Study, NW Queensland, Australia”, dated, 10 October, 2011, make the following
statements:

1)    I am employed as a Principal Geologist with Golder Associates Pty. Ltd. 601 Coronation Drive,
      Toowong, Brisbane, Queensland, 4066, Australia.

2)    I graduated with a degree in B Sc (Hons) in geology from the University of Queensland in 1986. In
      addition, I have obtained postgraduate diploma degrees in computing from the University of
      Queensland in 1987 and a postgraduate certificate in geostatistics from Edith Cowan University in
      2006.
3)    I am a fellow in good standing of the Australian Institute and Mining and Metallurgy (AusIMM
      #107320) and a member of the Australian Institute of Geoscientists (AIG #1844).

4)    I have practiced my profession continuously since graduation.

5)    I have read the definition of “qualified person” set out in National Instrument 43-101 (N.I. 43-101)
      and certify that, by reason of my education, affiliation with a professional association (as defined in
      N.I. 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person”
      for the purpose of N.I. 43-101.

6)    My relevant experience with respect to Mount Dore and Merlin Deposits includes over 25 years in
      exploration, mining geology and grade estimation of copper in Australia and Asia. This includes
      associated molybdenum mineralisation as well as molybdenum dominant work in Mexico. Over all
      of this period this experience includes a number of copper deposits within the same province as
      Mount Dore with similar geological settings.

7)    I am responsible for the preparation of all sections of the original technical report titled “2010
      Mineral Resource Estimate, Mount Dore - Merlin Deposit, NWQld, Australia” and dated
      19 Oct 2010 which has been used as the basis for some section of this study titled “Merlin Project
      Prefeasibility Study, NW Queensland, Australia” and dated 10 October-2011 (the “Technical
      Report”) relating to the Mount Dore - Merlin property. I visited the Mount Dore – Merlin property on
      17 Feb 2010 for 6 days and subsequently on the 22 April 2011 for 6 days.

8)    I have not had prior involvement with the property that is the subject of the Technical Report other
      than the previous technical report dated 19 Oct 2010.

9)    As of the date of this Certificate, to my knowledge, information and belief, the sections of this
      Technical Report for which I am responsible contains all scientific and technical information that is
      required to be disclosed to make the technical report not misleading.

10)   I am independent of the Issuer as defined by Section 1.4 of the Instrument. I have read National
      Instrument 43-101 and the sections for which I am responsible in this Technical Report have been
      prepared in compliance with National Instrument 43-101 and Form 43-101F1.

Dated this 27 Day of October 2011




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                               CERTIFICATE of AUTHOR – Tom Hunter

I, Tom Hunter, BSc (Met), SEC, Grad Dipl Earth Sciences (Min Econ), FAUSIMM (CP) do hereby certify
that:

1.    I am Associate Director, Business Development, Mining and Minerals of Jacobs E&C Australia
      Pty Ltd of 600 St Kilda Road, Melbourne, 3000 Victoria, Australia.
2.    I am one of the authors of Nl 43-101 Technical Report titled "Merlin Project Prefeasibility Stud y
      NW Queensland, Australia” (The Technical Report), Dated, 10 October 2011 and was involved
      with sections 17.2, 21 (processing component) and 26.4.
3.    I graduated with a Bachelor of Science (Metallurgy) from the University of Queensland in 1970
      and a Bachelor of Economics from James Cook University of North Queensland in 1973 and also
      received a graduate diploma in Earth Sciences (Mineral Economics) from Macquarie University
      in 1982.
4.    I am a Fellow and Chartered Professional of The Australasian Institute of Mining and Metallurgy.
5.    I have worked as a metallurgist for a total of 40 years since my graduation from university and a
      total of 42 years in the mining industry.
6.    My relevant work e xperience for the purpose of The Technical Report is:
      •      30 years as a metallurgist or manager specialising in metals production, project and
             business development.
      •      10 years as a manager involved with engineering and business development.
7.    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 Nl 43-101) and past relevant work
      experience, I fulfil the requirements to be a Qualified Person for the purposes of Nl 43-101.
8.    I am responsible for the preparation of Sections 17.2, 21 (processing component), 26.4 of The
      Technical Report.
9.    I have not visited the site.
10.   I have not had any prior involvement with the property that is the subject of The Technical
      Report.
11.   I am independent of the issuer applying all of the tests in section 1.4 of National Instrument
      43-101.
12.   I have read National Instrument 43-101 and Form 43-101 F1, and my sections of the Technical
      Report have been prepared in compliance with that instrument and form.
13.   As of the date of this certificate, to the best of my information, knowledge and belief, my sections
      of the Technical Report contains all scientific and technical information that is required to be
      disclosed to make the Technical Report not misleading.

Dated this 27th Day of October 2011




Original signed and sealed by:
Tom Hunter, BSc (Met), SEC, Grad Dipl Earth Sciences (Min Econ), FAUSIMM (CP).




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