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Review of Metallurgical Testwork undertaken on Batman Deposit Gold Ore Samples from the Mt Todd Project Feasibility Study for Vista Gold Australia Pty Ltd
Table of Contents 1.1 Introduction 8 1.2 Testwork Conclusions 8 1.3 Review of Previous Metallurgical Testwork 11 1.4 Sample Selection, 2011 Test Work Campaign 20 1.5 Head Assays 23 1.6 Mineralogy 31 1.7 Comminution Testwork 34 1.8 Leach Optimisation 55 1.9 Effect of HPGR on Gold Extraction 65 1.10 Cyanide Variability Testwork 67 1.11 Bulk Leaches 76 1.12 Ancillary Testwork 77 1.13 Agitator Testwork 79 1.14 Materials Handling 80 1.15 Cyanide Destruction 80 1.16 Thickener Testwork 82 1.17 Flotation Testwork 83 1.18 Magnetic Separation Testwork 84 1.19 Recovery and Reagent Consumption 85 1.20 Further Recommended Testwork 89 Date: January 2013 Page 3 of 92
List of Tables Table 1 - Average gold and leach extractions from leach test work only 9 Table 2 - Abrasion Index Results 12 Table 3 - Bond Mill Work Indices 12 Table 4 - Whole ore leach tests, HPGR vs. Conventional Crush 13 Table 5 - Grind optimisation and pre-conditioning tests 15 Table 6 - NaCN optimisation results, maintain / decay, Composite 2 16 Table 7 - NaCN optimisation results, maintain / decay, Composite No. 1 16 Table 8 - Cyanide optimisation leach tests 17 Table 9 - CIP test results 18 Table 10 - Drill hole VB08-041 leach test results 19 Table 11 - Mine Plan Summary 21 Table 12 - Comminution samples by origin 22 Table 13 - Main Ore Type Composite Head Assay 24 Table 14 - Variability composite head assay 26 Table 15 - Mt Todd Mineral Abundance 32 Table 16 - Gold in Batman samples 33 Table 17 - Average particle and grain sizes 34 Table 18 - SMC Test results 35 Table 19 - CWI Results 36 Table 20 - UCS Results 37 Table 21 - BWi Results (106 µm closing screen) 38 Table 22 - BWI at different closing screen sizes for the master composite sample 39 Table 23 - BWI at different closing screen sizes for various variability samples 40 Table 24 - Bond Work Index (200 µm closing screen), HPGR Product vs. Conventionally Crushed feed, Polysius 41 Table 25 - LABMILL Testwork Results 43 Table 26 - Bond Work Index, MHT-001 & MHT-004 HPGR Product, AMMTEC 44 Table 27 - Bond Work Index, 2010 HPGR Sample HPGR Product, AMMTEC 44 Table 28 - Grind establishment tests, conventionally crushed verses HPGR product 45 Table 29 - ATWAL abrasion test results 47 Table 30 - ATWAL wear rate categories 47 Table 31 - Summary of MAGRO semi-industrial scale test results 48 Table 32 - Initial grind optimisation tests 50 Table 33 - Final grind optimisation tests 52 Table 34 - Comparative NPV for Batman ores at various grind sizes and leach times 54 Table 35 - Comminution circuit design criteria 54 Date: January 2013 Page 4 of 92
Table 36 - Pre-aeration Tests 56 Table 37 - Sparging Tests 57 Table 38 - Cyanide Optimisation Tests 58 Table 39 - Pulp density optimisation tests 59 Table 40 - Pre-conditioning and Lead Nitrate Tests 62 Table 41 - Lead nitrate and pre-conditioning option comparison 63 Table 42 - Cyanide maintenance tests 64 Table 43 - HPGR verses conventionally crushed CIL tests, MHT-001 and MHT-004 65 Table 44 - HPGR verses conventionally crushed CIL tests, 2010 HPGR composite 66 Table 45 - Leach Variability Tests, 125 µm grind size 69 Table 46 - Leach variability tests, 90 µm grind size 74 Table 47 - Bulk Leach Tests 76 Table 48 - Fleming Constants, Testwork 77 Table 49 - Fleming Constants, Design 77 Table 50 - Oxygen uptake rate (master composite) 78 Table 51 Slurry Apparent Viscosity 78 Table 52 - Leach Residue Solution Analysis 80 Table 53 - Air SO2 Test Results, Sample WH4270 81 Table 54 - Air SO2 test results, sample WH4271 81 Table 55 - Thickener Testwork Results 82 Table 56 - Rougher Flotation Testwork Results 83 Table 57 - Magnetic Separation, 100% passing 3.35 mm, master composite 84 Table 58 - Magnetic Separation, 100% passing 3.35 mm, Var-79 84 Table 59 - Leach results at optimised conditions (24 h), main composites, 90 µm 85 Table 60 - Leach results at optimised conditions (24 h), variability composites 86 Table 61 - Leach Extraction Results 87 Table 62 - Leach extraction results, by ore type 87 Table 63 - Design Recoveries 88 Table 64 - Reagent consumptions from leach testwork only 88 Table 65 - Reagent consumptions from leach testwork only, by ore type 88 Date: January 2013 Page 5 of 92
List of Figures Figure 1 - Plan View of 2011 Metallurgical Drill Holes 20 Figure 2 - Grade comparison expected verses versus actual assays - variability composites 25 Figure 3 - Cyanide soluble copper relationship 31 Figure 4 - Closing screen size versus average BWi for master composite sample 39 Figure 5 - Closing screen size versus average BWi for variability samples 41 Figure 6 - MHT-001 Feed and Product Size Distributions 42 Figure 7 - MHT-004 Feed and Product Size Distributions 42 Figure 8 - Master composite leach kinetics at various grind P80 sizes, initial grind optimisation 50 Figure 9 - Initial Grind Optimisation NPV Analysis 51 Figure 10 - Master composite leach kinetics at various grind P80 sizes, final grind optimisation 53 Figure 11 - Final Grind Optimisation NPV Analysis 53 Figure 12 - The effect of pre-aeration on leach kinetics 56 Figure 13 - The effect of dissolved oxygen levels on leach kinetics 58 Figure 14 - The effect of pulp density on leach kinetics 60 Figure 15 - The effect pre-conditioning and lead nitrate conditioning on gold leach kinetics 62 Figure 16 - The effect pre-conditioning and lead nitrate conditioning on copper leach kinetics 63 Figure 17 - Grade extraction relationship for gold, 125 µm grind size 73 Figure 18 - Grade extraction relationship for copper, 125 µm grind size 73 Figure 19 - Grade extraction relationship for gold, 90 µm grind size 75 Figure 20 - Grade extraction relationship for copper, 90 µm grind size 75 Figure 21 - Copper extraction verses depth, drill holes VB11-009 and VB11-010 76 Figure 22 - Oxygen uptake rate (master composite) 78 Figure 23 - Viscosity versus Shear Rate (master composite) 79 Date: January 2013 Page 6 of 92
Appendices Appendix 1 Testwork Sample Selection Appendix 2 ALS Ammtec Metallurgical Testwork Report Appendix 3 Mt Todd Gold ECS Grind Optimisation Report Date: January 2013 Page 7 of 92
1.1 Introduction A metallurgical test work program was developed by Vista Gold on samples representing the Batman deposit to support the Mt Todd Project Feasibility Study (FS) during 2011-2012. The aim of this program was to re-validate the findings of previous testwork programs, and to expand on the understanding of the metallurgical variability of the Batman orebody. The program was completed by ALS Ammtec in Perth, Western Australia, during the second half of 2011 and first quarter of 2012. The test work program was managed by Vista Gold and comprised: comminution characteristics of the ore, including ball mill work index tests, SMC tests, crusher work index tests and unconfined compressive strength tests for selected comminution and variability composites grind establishment cyanide leaches on the master composite sample cyanidation concentration leaches on the master composite sample investigating the optimum sparging conditions for the master composite sample investigating the benefit of pre-aeration, pre-conditioning and lead nitrate addition for the master composite sample investigating the benefit of HPGR for gold extraction gathering of engineering data, including oxygen uptake, carbon kinetics and slurry viscosity cyanide detoxification tests by the continuous air/so2/cu2+ process leach performance of 99 variability samples at a grind size of 125 µm P80 and of 30 variability samples at a grind size of 90 µm P80 identifying the level of magnetic sulfide mineralisation within the master composite sample rougher flotation performance mineralogical assessment on selected variability samples. In addition, several other tests were performed as below. ALS Ammtec prepared samples for dispatch to: SPX Flow Technology to establish the optimum leach density and agitation requirements Outotec for thickening testwork ThyssenKrupp Polysius for high pressure grinding and ball mill grindability testwork TUNRA to establish the materials handling characteristics. Test work was completed on composite and variability samples prepared by Vista Gold from eight drill holes that intersected the ore beneath the existing pit in the Batman deposit and the samples are considered representative of the Batman ore. 1.2 Testwork Conclusions Conclusions drawn from the comminution and metallurgical test work programs on the composites and variability samples of the remaining ore in the Batman deposit are detailed below: Date: January 2013 Page 8 of 92
The testwork performed is considered sufficient to demonstrate the flowsheet and provide the process plant design criteria for the FS design The Batman ore is considered free-milling and non preg-robbing and amenable to gold extraction by conventional cyanidation processes The Batman ore is classified as competent to very competent The results of the crusher work index ( CWI ) tests on selected Batman variability samples ranged from 4.7 kwh/t to 8.4 kwh/t, with an average of 6.3 kwh/t and 75th percentile value of 8.4 kwh/t The results of the BWi tests on the Batman comminution samples indicated little variability between the samples with an average of 26.1 kwh/t and a 75th percentile value of 27.7 kwh/t at a closing screen size of 106 µm, with all samples displaying above average hardness relative to the global database There are two main ore types in the Batman deposit (oxide and sulphide ore) that can be defined and classified by their different leach performance and reagent consumption rates. The two main ore types have been categorised by depth, with the oxide ore defined as material above the base of oxidation while the sulfide ore is below the base of oxidation. The base of oxidation, on average, is at a depth of 40 m below surface The average gold and copper leach extractions at the grind size P80 of 90 µm, and CIL reagent consumptions for the Batman oxide and sulfide ore samples are shown in Table 1. It should be noted that actual plant extractions are typically marginally lower than achieved in the laboratory. These figures exclude the residual free cyanide required at the end of the leach circuit to inhibit copper loading onto the carbon, as well as the cyanide required for the cold cyanide wash in the elution circuit. Table 1 - Average gold and leach extractions from leach test work only Ore Type Gold Leach Extraction % Au Copper Leach Extraction % Cu Sodium Cyanide Consumption kg/t NaCN 1 Lime Consumption kg/t (60% CaO) 2 Oxide Ore 84.4 42.4 0.77 0.64 Sulfide Ore 81.7 6.6 0.59 0.54 Weighted Average 81.7 6.9 0.59 0.54 1 In addition to the cyanide consumption numbers shown in Table 1, 0.16 kg/t NaCN for sulfide ore and 0.56 kg/t NaCN for oxide ore was added to ensure sufficient free cyanide exists in the system to prevent copper loading onto carbon. To accommodate the cold cyanide wash in the elution circuit, an additional 0.02 kg/t was added. This resulted in a total cyanide consumption used in operating costs of 0.77 kg/t. 2 In addition to the lime consumption numbers shown in Table 1, 0.84 kg/t lime for sulfide and 3.10 kg/t lime for oxide ore is required for the cyanide detoxification circuit. This resulted in a total weighted average lime consumption of 1.40 kg/t lime at 60%w/w solids. This equates to a total consumption of 0.91 kg/t at 92% CaO, which is the quality of the lime sourced for the Mt Todd project. Date: January 2013 Page 9 of 92
There is high variability in cyanide soluble copper between the oxide and sulfide ore samples. This is attributed to the difference in copper mineralogy: oxide material contains secondary copper minerals (such as chalcocite and bornite) while sulfide material contains primary copper minerals (such as chalcopyrite). The leach kinetics for secondary copper minerals in cyanide solution are typically faster than for primary copper minerals The current mine schedule indicates that the oxide ore represents only 0.66% of the total tonnage in the Batman reserve. The oxide ore will not be processed separately but blended with sulfide ore, with the expected performance of the blended feed approximately the weighted average shown in Table 1. For the Batman samples tested, pre-conditioning with lime for 4 hours prior to cyanidation decreased the cyanide consumption by up to 30%, but also increased the lime consumption by 60%. There was no improvement in gold extraction observed with pre-conditioning. This is contrary to the findings of 2010 feasibility testwork, which showed a significant increase in gold extraction as well as a reduction in cyanide consumption (see Table 5). Consequently, the preconditioning step that existed in the 2010 prefeasibility study flowsheet, has been removed There appears to be no benefit to leach kinetics from elevated dissolved oxygen levels achieved by the sparging of oxygen in place of air, therefore air sparging only is recommended for the CIL circuit The addition of lead nitrate [Pb(NO3)2] to the leach at a rate of 100 g/t appears to increase the gold leach kinetics for the Batman ore, as well as reducing NaCN consumption. Pb(NO3)2 addition also has the added benefit of suppressing the leaching of copper, reducing the formation of copper cyanide complexes There was no observed improvement in leach kinetics or overall gold extraction on samples crushed by HPGR and ground to a P80 of 90 µm, compared to samples that were conventionally crushed Dynamic thickening testwork indicates underflow densities greater than 65%w/w solids with acceptable overflow clarity could be achieved with settling rates between 0.2 t/m2/h and 1.64 t/m2/h at a P80 grind size of 125 µm. The grind optimisation testwork has indicated an optimum P80 of 90 µm and further thickening testwork is required to confirm the settling rates for thickener sizingat this finer grind size The Batman ore contains magnetic pyrrhotite, which contains up to 40% of the gold mineralization. This is to be considered further in the design of the tramp metal management system HPGR testwork and modeling indicates a circuit specific energy of 4.1 kwh/t,confirming the high ore competency indicated by the SMC DWi and Axb values measured for this ore HPGR circuit modeling indicates a relatively high circulating load, due to: - The relatively coarse HPGR product (reflecting the high ore competency) - The fine wet screen separation size proposed for the operation (required to minimize the milling inefficiencies related to the very high BWi values) BWi testwork completed on HPGR products has indicated a reduction in the measured ball mill work index for HPGR crushed material compared to conventionally crushed material. This observation is supported by the shorter grind times required by the HPGR product in the laboratory mill to achieve a P80 of 90 µm Preliminary rougher flotation tests achieved poor copper recoveries ranging between 40% and 45%, which was not deemed sufficient to justify the inclusion of a flotation circuit in the flowsheet. Flotation was not pursued further as a potential processing path for the Batman ores Date: January 2013 Page 10 of 92
TUNRA materials handling tests indicated that the Batman sulfide ore should be moderately easy handling material with a low bulk strength. The testwork indicates no special considerations are required for the chute/bin design for handling the ore Cyanide detoxification tests using the air/so2/cu2+ process at a laboratory scale demonstrated that the weak acid dissociable cyanide (CNWAD) in the CIL tailings slurry can be reduced to <10 ppm CNWAD. 1.3 Review of Previous Metallurgical Testwork In this section, testwork undertaken during previous study phases is reviewed. Subsequent Sections 1.4 to 1.19 cover sample selection, assays and mineralogy, and testwork procedures and results related to the current testwork program. 1.3.1 Introduction Two test work campaigns have previously been completed on Mt Todd Batman orebody. Both campaigns were completed by Resource Development Incorporated (RDi) in its laboratory located in Colorado, USA. The results of these campaigns formed the basis of both the initial feasibility study completed by RDi in 2009, and the prefeasibility study (PFS) completed in 2010. 1.3.2 Background/Timeline A summary of the previous metallurgical testwork programs and reviews completed by Vista Gold follows: RDi completed a study in 2006 using historical metallurgical testwork results (completed by others) and proposed a conceptual flowsheet consisting of crushing, grinding, flotation to produce a copper concentrate and a float tailings CIL circuit to recover gold RDi later received some core samples from an historic drill program to perform new testwork to validate the conceptual flowsheet The core samples were classified as transitional ore with the predominant sulfide species being pyrite and secondary copper mineralisation (such as chalcocite and bornite). It should be noted that the current mine plan indicates that the transitional material represents only a small fraction the remaining resource the preliminary results from this RDi testwork program validated the conceptual proposed flowsheet and indicated an overall gold recovery of about 82%, with 90% of the copper reporting to a cleaner concentrate grading 19% Cu at a primary grind of 200 mesh Vista then completed two drilling programs in 2007 and 2008 focusing on the deeper ore beneath the existing Batman pit. This deeper ore was considered to be more representative of the remaining resource than that sampled for the 2006 testwork The testwork performed on the 2007/2008 drill core failed to duplicate the results of the previous 2006 testwork program on the historic core samples. Subsequent investigations indicated a significant change in mineralogy of the deposit with increasing depth. The copper mineralogy changed from secondary to primary copper mineralisation (ie. from chalcocite to chalcopyrite) and pyrrhotite replaced pyrite as the dominant sulfide species Date: January 2013 Page 11 of 92
The testwork program failed to produce an acceptable copper concentrate from the drill core samples representing the deeper, primary ore. Given the low copper prices at that time the testwork was undertaken, Vista Gold decided to remove the copper flotation circuit from the proposed flowsheet and shift the focus to optimising gold recovery. All further testwork focused on optimising a whole ore leach process flowsheet. 1.3.3 Samples The following composites were prepared for RDi s test work program from samples representing deeper Batman ore from the 2007/2008 drilling program: Composite 1-1200 kg composite sample made up from 2007 drill core. The composite included samples from five drill holes selected to be representative of a cross section of the deposit. The head assay was 1.3 g/t Au, 0.92% S and 447 ppm Cu. The sequential copper analysis indicated 80.4% of the copper in the sample was primary copper. The dominant sulfide in the sample was pyrrhotite Composite 2-140 kg composite sample made up from 2008 drill core with a head assay of 0.89 g/t Au and 450 ppm Cu. The sequential copper analysis indicated 80.3% of the copper in the sample was primary copper. Similar to Composite 1, the dominant sulfide in the sample was pyrrhotite Drill hole 41 sample sourced from the oxide and transitional zones (depth of 0 65 m), with a head assay of 1.78 g/t, 1.42% S and 448 ppm Cu. The sample was split into four main interval samples labeled 3A, 3B, 3C and 3D. Composite 3A, representing the 0-20 m interval had copper oxides as the dominant copper species. Composite 3B representing the 20-30 m interval was dominated by secondary copper minerals. 1.3.4 Comminution Testwork Abrasion Index Abrasion tests were performed by Phillips Enterprises LLC, with the results shown in Table 2. Table 2 - Abrasion Index Results Composite Abrasion Index, Ai 1 0.1432, 0.1368 2 0.1301 Ball Mill Work Indices Bond ball mill work index test work was completed on the composites and on a single drill hole at varying depth. The results show little variation and no discernible trend with depth as summarized in Table 3. Table 3 - Bond Mill Work Indices Composite BWi, kwh/t Closing Screen, µm 1 26.3 150 2 28.5 150 Range of other results 25.5-31.6 150 Date: January 2013 Page 12 of 92
HPGR Testwork An HPGR (High Pressure Grinding Rolls) pilot test work campaign was completed by Polysius in Germany on a 100 kg sub-sample of Composite 1. The results were used to assess the difference in power requirement between a SAG/ball mill circuit, a 3-stage crush/ball mill circuit and a HPGR/ball mill circuit to generate a 150 µm P80 product. The assessment concluded that an HPGR/ball mill circuit has a significantly lower specific energy requirement (a difference of 9.48 kwh/t at a 150 µm P80 grind) than the SAG/ball option. On this basis, RDi selected an HPGR/ball mill circuit for the prefeasibility process flowsheet. Material from the HPGR pilot test work was returned to RDi and subjected to flotation and cyanide leaching test work. Cyanide leaching was completed on flotation tailings and whole ore cyanide leaching to assess if the inclusion of the HPGR would provide a benefit for copper and gold recovery. The test results showed no benefit for gold recovery in the leaching of flotation tailings or for whole ore leaching. Whole ore leach tests were performed on material crushed by an HPGR and compared to material that was conventionally crushed at various grind times. Leach conditions targeted 40%w/w solids, 0.2%w/v NaCN, 24 hours residence time, oxygen sparging and ph 11. The results are summarised in Table 4. Table 4 - Whole ore leach tests, HPGR vs. Conventional Crush Parameter Conventional Crush HPGR Crush Grind time, min 30 45 60 30 45 60 Grind size, P 80 µm 198 58 38 59 36 <36 Gold extraction, % Au 75.3 61.4 38.8 76.7 32.8 34.8 NaCN Consumption, kg/t 1.25 2.64 2.87 2.38 2.62 2.73 Lime Consumption, kg/t 3.15 3.17 3.07 2.05 1.95 1.82 The results indicated that a finer grind size can be achieved within the same grind time for HPGR crushed material compared to conventionally crushed material, with gold extraction decreasing at finer grind sizes. A possible reasoning for this behavior may be that at finer grind sizes, the liberation of cyanide consuming sulfide minerals increases consuming cyanide and reducing gold extraction. This outcome conflicts with the test work performed by ALS Ammtec in 2011 (see Section 1.7.9), which indicated an increase in gold extraction with finer grind sizes when combining oxygen sparging with the same leach conditions. The RDi test work was completed using bottle rolls, whereas the 2011 ALS Ammtec test work was completed in an agitated vat. 1.3.5 Gravity Testwork Gravity recovery test work at P80 grind sizes of 212, 150 and 106 µm was completed on the Composite 1 sample. The results indicated limited free gold at the grind sizes tested. Any free gold present did not concentrate to a significant extent. Gravity concentration was not pursued further. Date: January 2013 Page 13 of 92
1.3.6 Flotation Testwork (to confirm flowsheet developed in 2006) Flotation test work was completed on Composite 1. Rougher flotation tests resulted in 80% gold recovery to concentrate at a 75 µm grind size. Desliming and regrinding of rougher tailings followed by a scavenger flotation stage improved the overall gold float recovery to approximately 85%. The best copper concentrate produced contained 6% Cu, compared to 19% Cu concentrate produced in the 2006 test work program. Analysis of the concentrate indicated the dominant sulfide mineral was pyrrhotite and not pyrite. Further mineralogical examinations concluded that the dominant sulfide minerals in the majority of the remaining resource consisted of pyrrhotite and chalcopyrite (a primary copper mineral), whereas the sample used in the 2006 testwork program was predominately pyrite and secondary copper minerals. The production of a saleable copper concentrate was not as viable as indicated in previous test work. Also, the processing challenges associated with cyanide soluble copper encountered in previous operations at Mt Todd would be a lesser problem when treating the deeper primary ores. The flotation of copper was therefore deemed unnecessary, and a whole ore leach process flowsheet was pursued with a greater focus on gold recovery. 1.3.7 Leach Testwork Grind Optimisation Tests Grind optimisation cyanide leach tests were performed at P80 grind sizes of 212, 106, 75 and 53 µm, at an initial cyanide concentration of 0.1% NaCN and leach ph 11. Timed samples were taken at 6, 24, 48 and 72 hours. These tests were repeated with 0, 2 and 4 hours of pre-conditioning. Pre-conditioning was referred to as pre-aeration during this testwork, however no air or oxygen was added to the slurry at any point, with preaeration defined as the addition of lime to the slurry for ph adjustment prior to the addition of cyanide. This is referred to as pre-conditioning hereafter in this report. The results are shown in Table 5. The grind optimisation test work in Table 5 indicated that: approximately 80% of the cyanide soluble gold reports to solution within 24 hours gold extraction increases with fineness of grind. This contradicts the results observed during the HPGR testwork completed by RDi in Section 1.3.4, but is consistent with the grind optimization work completed by Ammtec in 2011 (see Section 1.7.9). As it was difficult to suspend solids at a P80 grind size of 212 µm, the results of this grind size were discounted. Grind sizes finer than 106 µm were not considered as it was concluded that the results indicated little improvement in gold extraction. This conclusion is however in conflict with the actual results, which indicated a difference of almost 6% recovery between the 106 µm and 53 µm grinds. The optimum primary grind size selected for the 2010 feasibility study was 150 µm. The selection criteria and analysis undertaken to derive at this primary grind size is however unclear. Pre-conditioning involved the addition of lime for ph adjustment, followed by agitating the slurry for 0, 2 or 4 hours before cyanide leaching. Results showed a significant improvement in gold recovery and a reduction in cyanide consumption with increased pre-conditioning times and a 4-hour pre-conditioning stage was included in the flowsheet. Date: January 2013 Page 14 of 92
Pre-conditioning tests performed during the 2011 test work program, on samples representing material beneath the existing Batman pit, did not replicate these results. Refer to Section 1.8.5. Table 5 - Grind optimisation and pre-conditioning tests Pre-conditioning Grind P 80 Calculated Head grade NaCN Lime Gold Extraction, % Copper Extraction, % Hours µm Au g/t Cu ppm kg/t kg/t 24 h 48 h 72 h 72 h 0 200 0.85 440 1.73 4.50 75.9 77.1 78.3 18.7 0 106 1.07 447 1.85 4.42 72.6 76.5 77.7 21.7 0 75 1.08 473 1.67 4.44 78.5 79.9 81.2 21.0 0 53 1.11 486 1.97 4.39 79.7 83.7 85.1 21.0 2 200 0.92 440 0.72 4.49 75.3 78.0 79.4 20.5 2 106 0.87 438 0.72 4.71 78.7 80.0 81.4 20.6 2 75 0.93 477 0.77 4.70 75.5 78.4 82.9 22.8 2 53 1.06 455 0.90 4.90 85 86.4 86.4 23.9 4 200 0.92 448 0.60 4.71 72.4 75.3 76.4 19.7 4 106 0.90 435 0.59 4.80 80.7 83.7 85.1 23.7 4 75 0.88 427 0.60 4.73 81.2 82.6 82.4 25.0 4 53 0.90 489 0.60 4.92 86.4 87.9 89.4 25.6 Cyanide Optimisation Tests Initial tests were performed on Composite 2 to examine the difference between maintaining cyanide concentration and allowing it to decay at varying grind sizes. The test conditions were: 4 hours of pre-conditioning initial ph of 11 40% w/w solids no air or oxygen sparging initial sodium cyanide concentration of 0.10%w/v, maintained at 0.10% w/v or allowed to decay P80 grind sizes of 75 µm, 106 µm and 150 µm. The outcome of this work is summarised in Table 6. Tests were also completed on Composite 1 at a grind size of 150 µm. These results are shown in Table 7. Date: January 2013 Page 15 of 92
Table 6 - NaCN optimisation results, maintain / decay, Composite 2 Pre-conditioning Grind Calculated Head grade NaCN Lime Gold Extraction, % Copper Extraction, % Maintain / Decay Hours µm Au g/t Cu ppm kg/t kg/t 24 h 48 h 48 h 4 150 1.29 498 0.43 4.83 78.2 81.3 18.5 Maintain 4 150 1.25 513 0.42 4.01 78.5 81.0 16.6 Maintain 4 150 1.03 475 0.30 4.03 76.0 77.4 16.2 Decay 4 150 1.19 505 0.31 3.96 77.8 79.2 15.3 Decay 4 106 1.10 475 0.48 5.02 78.9 79.8 17.5 Maintain 4 106 1.11 446 0.37 5.01 78.9 80.2 17.4 Maintain 4 106 1.05 486 0.36 5.14 79.2 79.1 17.3 Decay 4 106 1..11 448 0.42 5.57 76.8 79.0 17.8 Decay 4 75 1.08 542 0.42 6.46 78.6 81.2 17.3 Maintain 4 75 1.23 542 0.47 6.51 85.9 87.2 17.4 Maintain 4 75 0.97 490 0.43 6.52 81.5 84.2 17.5 Decay 4 75 0.97 445 0.42 6.49 79.1 83.4 19.0 Decay Table 7 - NaCN optimisation results, maintain / decay, Composite No. 1 Pre-conditioning Grind Calculated Head grade CN Lime Gold Extraction, % Copper Extraction, % Maintain / Decay Hours µm Au g/t Cu ppm kg/t kg/t 24 h 30 h 24 h 30 h 4 150 1.14 522 0.60 9.42 82.6-13.5 - Decay 4 150 1.1 496 0.89 9.43 79.1-13.2 - Maintain 4 150 1.03 457 0.54 9.42-80.4-14.3 Decay 4 150 0.93 484 0.60 9.51 79.3 82.2 12.8 14.5 Maintain The test work results presented in Table 6 and Table 7 indicate that: composite 1 exhibits better leach kinetics compared to Composite 2 at a P80 grind size of 150 µm, but the NaCN and lime consumptions are higher for Composite 1 almost 80% of the soluble gold reported to solution within 24 and 48 hours, with either the decay and maintain mode at all grind sizes tested maintaining a constant NaCN concentration over the course of leaching is not beneficial. Date: January 2013 Page 16 of 92
A second series of leach tests was completed at a range of initial NaCN concentrations. The test conditions were: 1 kg sub-samples of Composite 1 P80 grind size of 150 µm 40% w/w solids (assumed Colorado tap water) initial ph of 11 adjusted with lime 4 hours of pre-aeration initial cyanide concentration varying between 0.05%, 0.075% and 0.10% w/v, maintained at that level or allowed to decay no air or oxygen sparging 36 hours leaching with solution assays taken at 6, 24, 30 and 36 hours (solutions were not assayed at all times for each sample). The results are summarised in Table 8. Table 8 - Cyanide optimisation leach tests NaCN concentration Calculated Head grade NaCN Lime Gold Extraction, % Maintain / Decay % Au g/t Cu ppm kg/t kg/t 24 h 30 h 36 h 0.1 0.79 428 0.54 9.48 - - 82.2 Decay 0.1 0.85 420 0.60 9.41 82.8-84.0 Maintain 0.075 1.07 475 0.47 7.98 - - 86.5 Decay 0.075 0.83 474 0.50 7.97 82.1 83.5 - Maintain 0.05 0.9 472 0.39 7.97-79.7 - Decay 0.05 0.96 478 0.45 7.97 81.1 82.8 - Maintain The test results presented in Table 8 indicate that increasing the initial cyanide concentration from 0.05% to 0.1% had little impact on gold extraction. It was recommended that a NaCN concentration of 0.05% be maintained during leaching or an initial NaCN concentration of 0.075% be used. However, despite this recommendation, all further leach testwork conducted by RDi used an initial NaCN concentration of 0.1%. Date: January 2013 Page 17 of 92
Carbon-in-Pulp (CIP) Tests Four CIP tests were performed under the following conditions: 1 kg sub-samples of Composite 1 grind size of 150 µm 40% w/w solids initial ph of 11 adjusted with lime 4 hours of pre-aeration initial cyanide concentration of 0.1% w/v, run in decay mode no air or oxygen sparging 30 hours leaching with solutions assayed at 30 hours. 20 g/l of carbon was added after 24 hours. The results of the CIP tests are summarised in Table 9. Table 9 - CIP test results Calculated Head grade NaCN Lime Gold Extraction, % Copper Extraction, % Au Cu kg/t kg/t 30 h 30 h 1.05 468 0.52 3.10 82.3 14.1 1.01 415 0.58 3.09 82.0 15.6 1.04 444 0.40 3.09 85.4 14.4 1.15 465 0.46 3.11 86.7 14.4 The results presented in Table 9 show that after 30 hours, the gold extraction ranged between 82.0% and 86.7%. The four CIP tests returned the highest gold extractions achieved in the 2010 feasibility testwork program and were used as the basis of the 2010 feasibility design. Based on these results, an extraction of 82% was recommended for process plant design. The higher extractions achieved in the CIP tests are difficult to explain, as the carbon was not added until after 24 hours of leaching. This could be a function of inaccuracy in the gold-on-carbon assays, used in calculating the gold extraction. Variability Tests Leach tests were performed on the four composites made up from drill hole VB08-041, which represented the oxide and transitional zones in the ore body. The tests conditions were: grind size of 150 µm 40% w/w solids initial ph of 11 with lime 4 hours of pre-conditioning initial cyanide concentration of 0.1% w/v, run in decay mode no air or oxygen sparging 30 hours leaching with solution assay taken at 30 hours. Date: January 2013 Page 18 of 92
The results of the four variability tests are summarised in Table 10. Table 10 - Drill hole VB08-041 leach test results Composite Calculated Head grade NaCN Lime Gold Extraction, % Copper Extraction, % Au g/t Cu ppm kg/t kg/t 30 h 30 h 3A 1.04 993 1.31 4.49 38.7 58.7 3A 2.23 949 1.26 4.50 16.6 57.2 3B 0.45 654 1.25 4.23 63.2 79.2 3B 0.34 701 1.25 4.22 65.6 79.2 3C 0.57 444 0.60 4.46 86.1 48.6 3C 0.53 440 0.66 4.49 83.9 50.0 3D 0.62 476 0.78 5.45 90.0 56.7 3D 0.75 502 0.83 5.49 90.8 57.8 The results of the variability tests summarised in Table 10 show a higher copper extraction for the four oxide and transitional samples. This is consistent with the experience of previous Mt Todd operations that mined ore from the oxide and transitional zones. Sample 3B marks the start of the transitional material, where the dominant copper species is secondary copper as distinct from the copper oxides in Sample 3A. Due to the high cyanide soluble copper levels in the oxide and transitional ores and the resulting high cyanide consumption, blending of material from these zones should be practised. 1.3.8 Thickening Testwork Dynamic thickening test work was undertaken on leach feed and leach tail samples of Composite 1. The tests were completed at ph 8 (to simulate the conditions for pre-leach thickening) and ph 11 (tailings thickening) at 150 µm P80. The test work showed that the settling rate is ph dependent, with a higher settling rate achieved at ph 11. On this basis a tailings thickener was incorporated into the flowsheet. A design settling rate of 0.11 m2/t/d (or 0.38 t/m2h) was selected to achieve an underflow solids density of 55% w/w. The selected settling rate was for a conventional thickener and RDi proposed a 65 m conventional thickener. A high-rate thickener was selected for the 2010 feasibility study flowsheet to reduce the plant foot print and in the absence of high-rate thickener test work a typical settling rate of 1.5 t/m2h was adopted. 1.3.9 Rheology Testwork Basic rheology tests were performed, but the slurry settled quickly at 35% to 45% w/w solids and the viscosity could not be measured. This concludes the review of previous testwork. The following section relates to sample selection for the current test work program. Date: January 2013 Page 19 of 92
1.4 Sample Selection, 2011 Test Work Campaign Samples used for the 2011 metallurgical test work program (ECS testwork program) were sourced from eight holes from the 2010/2011 drilling program. The drill holes were orientated to intersect the main Batman ore body beneath the existing pit to be representative of the ore within the feasibility study pit shell. All samples from drill holes labelled VB11 were drilled in 2011, logged, packaged and shipped directly to the laboratory for processing. Drill holes labelled MHT were drilled and logged during 2010 and were stored in cold storage before being transported to the laboratory in 2011. Figure 1 shows the plan view of the drill holes within the Batman deposit for the 2011 metallurgical test work campaign. Figure 1 - Plan View of 2011 Metallurgical Drill Holes 1.4.1 Geology Background The Batman ores are classified as sulfide ore (or fresh), mixed ore (or transitional) and non-sulfide ore (or oxide), based on degree of weathering. The test work samples were predominantly logged as sulfide ore, although it is understood that due to the degree of weathering in the existing pit, it was difficult to distinguish between lithologies and the transitional and sulfide zones. The life of mine (LOM) production forecast for each ore type is shown in Table 11. Date: January 2013 Page 20 of 92
Table 11 - Mine Plan Summary Ore Type LOM Production, k t % of Total Reserves Non-sulfide 1,136 0.66 Mixed 1,917 1.11 Sulfide 168,894 98.22 Non-sulfide 1,136 0.66 TOTAL 171,946 100.00 Due to the dominance of sulfide ore in the mine plan and in the drill core samples available, the test work program focused predominantly on the performance of this ore type. As was demonstrated during the 2010 feasibility testwork program, material from the oxide zone has significantly different cyanide soluble copper levels compared to the sulfide ore, due to the difference in copper mineralisation. Oxide ore samples were selected from drill holes VB11-009 and VB11-010 and tested as variability samples to confirm the findings of the previous PFS testwork program and to predict the impact of processing the remaining oxide reserves. Taking into consideration the difficulty in distinguishing between lithologies and the transition and sulfide zones, a decision was made to combine the sulfide ore and transitional ore, reducing the main ore types to two sulfide and oxide. Furthermore, based on the leach results obtained during the 2011 test work program, the sulfide and oxide zones are classified based on pit depth. Test work has shown samples sourced from the top 40 m of the pit show different mineralogical characteristics compared to samples sourced from deeper in the pit, highlighted by the difference in cyanide soluble copper extraction in the leach test work. Copper mineralisation in the oxide zone consists predominately of oxide copper minerals and secondary copper sulfide minerals (mainly chalcocite and bornite), which are known cyanocides (ie. highly cyanide soluble), affecting reagent consumptions and possibly also gold extractions. Primary copper species (ie. chalcopyrite) dominate the copper mineralisation in the sulfide ore. All samples sourced from depths below 40 m are classified as sulfide. As the majority of the oxide material has already been mined by previous operators and the remaining oxide ore represents only 0.66% of the total resource, it was important to consider this ore separately, so that the results do not artificially impact on the calculated averages for recovery and reagent consumption. Only two host rock lithologies, greywacke and siltstone, were identified and logged during the 2010/2011 drilling program. However, there was no clear classification or delineation between the greywacke dominant or siltstone dominant domains within the pit shell and consequently the variability test work did not include evaluation of the two lithologies. 1.4.2 Comminution Samples Twenty comminution samples were selected from the 2010/2011 drilling program. The sample origins (drill hole and interval) are provided in detail in Appendix 1. A summary of the comminution samples is shown in Table 12. The samples were selected to represent each of the drill holes, with three or four composite samples selected from each. All composite samples from drill holes MHT-001 and MHT-004 were selected to allow comparison to the comminution test work completed by Polysius on the same samples. Date: January 2013 Page 21 of 92
Table 12 - Comminution samples by origin Sample Composite No. Drill Hole Interval Comminution 1 #04 MHT-003 383 394 Comminution 2 #10 MHT-003 448 457 Comminution 3 #13 MHT-003 478 489 Comminution 4 #24 VB11-001 434 445 Comminution 5 #27 VB11-001 467 480 Comminution 6 #28 VB11-001 512 529 Comminution 7 #42 VB11-002 374 384 Comminution 8 #47 VB11-002 423 432 Comminution 9 #52 VB11-002 470 479 Comminution 10 #55 VB11-002 498 523 Comminution 11 #56 MHT-001 227 279 Comminution 12 #57 MHT-001 279 328 Comminution 13 #58 MHT-001 328 403 Comminution 14 #60 MHT-004 285 335 Comminution 15 #61 MHT-004 335 383 Comminution 16 #62 MHT-004 383 436 Comminution 17 #63 MHT-004 436 499 Comminution 18 #65 VB11-003 353 363 Comminution 19 #69 VB11-003 396 407 Comminution 20 #74 VB11-003 452 467 1.4.3 Metallurgical Composites Four main ore type composite samples were generated from drill core available from the 2010/2011 drilling program: master composite MHT-001 drill hole composite, used for HPGR testwork MHT-004 drill hole composite, used for HPGR testwork 2010 HPGR composite, used for additional HPGR testwork. Date: January 2013 Page 22 of 92
The origin (drill hole and interval) of each sample selected for the ore type composites is provided in Appendix 1.1. The samples were selected to give a spatial representation of the Batman deposit. The master composite is made up of 3 kg sub-samples of the variability composite samples 1 to 79 to create sufficient mass for the planned test work program. Each variability composite represents approximately a 10 m interval by drill hole depth, which was selected to ensure there was sufficient mass of each composite for test work (approximately 90 kg). The shortage of available drill core and the mass requirement of the master composite did not allow the creation of a master composite sample that closely reflected the design head grades. Variability samples 80 to 99 from drill holes VB11-009 and VB11-010 were late in arriving at ALS Ammtec and were not be included in the master composite recipe. The MHT-001 and MHT-004 composites were made up from 2010 drill core that was sourced from cold storage. These samples were selected by Vista Gold specifically for HPGR test work. Sample preparation was completed at ALS Ammtec before the samples were shipped to Polysius in Germany. Sub-samples of the MHT-001 and MHT-004 composites were kept at ALS Ammtec for the variability leach test work program. The 2010 HPGR composite was made up from 2010 drill core that was sourced from cold storage. The testing of 2010 samples was not part of the original testwork program. Due to the inconclusive nature of the HPGR versus conventionally crushed leach tests outlined in Section 1.9.1 and the shortage of 2011 samples to complete further HPGR testwork, this composite was created from 1 m intervals of core samples from nine 2010 drill holes available on site. 1.4.4 Leach Variability Samples Leach variability samples were selected to be representative of the various levels of oxidation and spatial representation and variations in grade that make up the Batman ore deposit. A total of 99 variability composites were constructed from core available from the 2010/2011 drill program. Full details of drill hole, intersection and gold assay used to make up each of the variability samples are provided in Appendix 1.1. 1.5 Head Assays 1.5.1 Main Composites The head assays of the four main composites are shown in Table 13. Salient points are: The gold head grade of the master composite is higher than the design head grade of 0.85 g/t. The master composite was used to determine the optimal leach conditions and not to calculate process plant recovery or reagent consumption. The 2010 HPGR composite sample assay is very similar to the master composite apart from the arsenic assay, therefore the two leach tests are comparable. The MHT-001 and MHT-004 composites have a lower gold and copper grade than the other composites. The head grades of the MHT-001 and MHT-004 composites are less important as they were dedicated to comminution testwork. Date: January 2013 Page 23 of 92
Table 13 - Main Ore Type Composite Head Assay Analyte Unit Master Composite MHT-001 Composite MHT-004 Composite 2010 HPGR Composite Au (1) g/t 1.06 0.80 0.49 1.01 Au (2) g/t 1.02 0.71 0.45 0.79 Au (3) g/t 1.15/1.23 N/A N/A 1.02 Au (average) g/t 1.01/0.96 N/A N/A 0.92 Ag g/t 1.1 0.5 0.4 0.9 As ppm 290 40 130 8.92 Cu ppm 482 218 206 590 CN Sol. Cu ppm 204 36 84 N/A C TOTAL % 0.15 0.12 0.15 0.12 C ORGANIC % < 0.03 <0.03 0.03 <0.03 Fe % 5.5 5.08 4.84 6.82 Hg ppm < 0.1 0.1 <0.1 0.1 Pd ppm < 0.02 <0.02 <0.02 N/A Pt ppm < 0.02 <0.02 <0.02 N/A S TOTAL % 1.32 0.68 0.8 1.34 S SULFIDE % 1.04 0.54 0.6 1.16 True SG (1) 3.35 2.81 2.78 2.80 1.5.2 Variability Composites The head assay of the variability samples are shown in Table 14. The following conclusions are drawn from the variability composites presented in Table 14 without significant exception, the gold grades show little variability (difference between Au1 and Au2), indicating that it is unlikely that there is coarse free gold present the total carbon (CTOTAL) levels in samples vary significantly. However, this is thought to be a function of carbonate levels, as the organic carbon content is very low for all samples, hence the risk of preg-robbing is considered negligible for all composites. Figure 2 shows a comparison of the expected gold grade and the analytical assays of each sample as determined by ALS Ammtec. This indicates no bias in the results when comparing the geological assays to the assays completed by ALS Ammtec. There were an equal number of higher results reported by ALS Ammtec as there were for the geology assays. Date: January 2013 Page 24 of 92
Figure 2 - Grade comparison expected verses versus actual assays - variability composites Date: January 2013 Page 25 of 92
Table 14 - Variability composite head assay Variability Composite Drill Hole Au 1 Au 2 Ag As C TOTAL C ORGANIC Cu CN Sol. Cu S TOTAL S SULFIDE (g/t) (g/t) (g/t) (ppm) (%) (%) (ppm) (ppm) (%) (%) TRUE SG Var-1 MHT-003 0.42 0.43 0.3 50 0.15 < 0.03 108 26 0.48 0.40 2.78 Var-2 MHT-003 0.43 0.42 2.5 50 0.33 < 0.03 168 34 0.92 0.80 2.78 Var-3 MHT-003 1.29 1.20 1.0 30 0.21 0.06 346 144 1.34 1.10 2.81 Var-4 MHT-003 2.16 2.28 1.1 40 0.18 0.06 414 116 1.76 1.48 2.78 Var-5 MHT-003 1.53 1.47 0.8 20 0.24 0.09 656 274 1.66 1.38 2.81 Var-6 MHT-003 1.62 1.58 1.2 40 0.24 < 0.03 454 220 1.4 1.20 2.81 Var-7 MHT-003 1.76 1.83 1.5 50 0.18 < 0.03 570 284 1.56 1.36 2.77 Var-8 MHT-003 1.89 2.04 1.2 30 0.15 < 0.03 698 360 1.54 1.32 2.82 Var-9 MHT-003 1.35 1.36 1.1 70 0.12 < 0.03 470 228 1.52 1.24 2.79 Var-10 MHT-003 0.91 0.87 0.9 40 0.12 < 0.03 384 170 1.24 1.06 2.80 Var-11 MHT-003 0.70 0.70 1.2 40 0.15 < 0.03 362 158 0.92 0.70 2.75 Var-12 MHT-003 1.08 1.08 1.0 40 0.24 < 0.03 330 156 1.02 0.88 2.80 Var-13 MHT-003 1.01 0.98 1.2 70 0.12 < 0.03 228 80 1.30 1.12 2.78 Var-14 MHT-003 0.42 0.42 3.1 190 0.21 < 0.03 732 210 1.38 1.06 2.81 Var-15 VB11-001 0.54 0.51 1.0 50 0.15 < 0.03 886 438 1.34 1.12 2.76 Var-16 VB11-001 1.79 1.93 1.4 820 0.06 < 0.03 592 258 1.34 1.08 2.76 Var-17 VB11-001 1.20 1.22 0.8 20 0.06 < 0.03 540 156 1.56 1.24 2.80 Var-18 VB11-001 0.93 1.16 2.3 120 0.39 < 0.03 1014 420 1.30 1.20 2.79 Var-19 VB11-001 0.54 0.53 2.2 970 0.36 < 0.03 1546 638 1.32 1.22 2.88 Date: January 2013 Page 26 of 92
Variability Composite Drill Hole Au 1 Au 2 Ag As C TOTAL C ORGANIC Cu CN Sol. Cu S TOTAL S SULFIDE (g/t) (g/t) (g/t) (ppm) (%) (%) (ppm) (ppm) (%) (%) TRUE SG Var-20 VB11-001 1.32 1.32 2.1 410 0.81 < 0.03 1118 508 1.74 1.48 2.82 Var-21 VB11-001 1.22 1.14 1.1 30 < 0.03 < 0.03 1100 480 2.06 1.70 2.79 Var-22 VB11-001 0.86 0.74 0.4 20 < 0.03 < 0.03 510 100 1.50 1.16 2.81 Var-23 VB11-001 0.9 0.72 0.4 10 0.03 < 0.03 522 78 1.48 1.20 2.77 Var-24 VB11-001 1.46 1.43 0.5 180 0.09 < 0.03 464 106 1.60 1.16 2.83 Var-25 VB11-001 1.93 2.14 1.0 30 0.06 < 0.03 742 210 1.66 1.42 2.75 Var-26 VB11-001 1.01 1.04 0.6 < 10 < 0.03 < 0.03 662 122 1.56 1.24 2.81 Var-27 VB11-001 0.71 0.70 0.5 60 < 0.03 < 0.03 4740 96 1.10 0.88 2.78 Var-28 VB11-001 0.51 0.54 0.5 50 < 0.03 < 0.03 522 96 1.38 1.02 2.80 Var-29 VB11-002 2.31 2.35 1.4 110 0.06 < 0.03 528 238 0.90 0.72 2.80 Var-30 VB11-002 3.83 3.76 1.9 11100 0.12 < 0.03 680 268 2.10 1.98 2.79 Var-31 VB11-002 0.57 0.52 1.1 110 0.09 < 0.03 250 104 0.52 0.48 2.77 Var-32 VB11-002 0.33 0.36 1.0 150 0.51 0.06 156 60 0.40 0.32 2.74 Var-33 VB11-002 0.23 0.22 1.0 40 0.18 < 0.03 122 66 0.30 0.22 2.79 Var-34 VB11-002 0.39 0.37 8.4 140 0.63 < 0.03 762 72 0.90 0.90 2.83 Var-35 VB11-002 0.36 0.39 1.4 50 0.18 < 0.03 312 90 1.04 1.04 2.80 Var-36 VB11-002 0.86 0.91 2.2 90 0.18 < 0.03 486 128 1.34 1.32 2.77 Var-37 VB11-002 1.25 1.18 2.0 180 0.30 < 0.03 642 120 1.50 1.50 2.82 Var-38 VB11-002 0.75 0.68 1.3 320 0.21 < 0.03 680 240 1.34 1.34 2.80 Var-39 VB11-002 3.04 3.79 0.7 60 0.15 < 0.03 820 50 1.70 1.52 2.84 Date: January 2013 Page 27 of 92
Variability Composite Drill Hole Au 1 Au 2 Ag As C TOTAL C ORGANIC Cu CN Sol. Cu S TOTAL S SULFIDE (g/t) (g/t) (g/t) (ppm) (%) (%) (ppm) (ppm) (%) (%) TRUE SG Var-40 VB11-002 1.04 1.10 0.7 40 0.09 < 0.03 842 42 1.92 1.86 2.81 Var-41 VB11-002 1.90 1.97 0.6 20 0.06 < 0.03 492 204 1.82 1.70 2.84 Var-42 VB11-002 1.18 1.27 0.7 20 0.09 < 0.03 498 74 1.48 1.40 2.78 Var-43 VB11-002 0.64 0.64 < 0.3 10 0.12 < 0.03 370 104 1.10 1.00 2.82 Var-44 VB11-002 2.25 2.10 1.4 60 0.12 < 0.03 652 86 2.04 1.84 2.78 Var-45 VB11-002 3.78 3.77 2.0 30 0.09 < 0.03 1088 166 2.5 2.38 2.88 Var-46 VB11-002 0.44 0.5 0.5 < 10 0.03 < 0.03 390 58 1.28 1.20 2.80 Var-47 VB11-002 0.62 0.6 0.5 20 0.06 < 0.03 536 156 1.20 1.08 2.83 Var-48 VB11-002 0.41 0.42 1.6 630 0.06 < 0.03 868 150 1.58 1.44 2.81 Var-49 VB11-002 0.44 0.41 0.6 < 10 0.06 < 0.03 622 80 1.30 1.20 2.79 Var-50 VB11-002 1.04 1.21 1.0 20 < 0.03 < 0.03 926 48 2.04 1.82 2.84 Var-51 VB11-002 0.62 0.64 0.8 < 10 0.03 < 0.03 784 46 1.48 1.32 2.76 Var-52 VB11-002 0.62 0.67 0.7 10 < 0.03 < 0.03 614 98 0.94 0.82 2.79 Var-53 VB11-002 0.98 0.84 1.0 10 < 0.03 < 0.03 808 50 1.30 1.24 2.78 Var-54 VB11-002 0.29 0.26 0.5 780 < 0.03 < 0.03 310 118 0.94 0.90 2.81 Var-55 VB11-002 0.13 0.11 0.4 80 0.09 < 0.03 358 354 0.98 0.88 2.78 Var-56 MHT=001 0.72 0.80 0.3 20 0.12 < 0.03 216 110 0.72 0.62 2.81 Var-57 MHT=001 1.56 1.37 0.4 80 0.09 < 0.03 328 284 0.72 0.64 2.77 Var-58 MHT=001 0.53 0.59 < 0.3 30 0.12 < 0.03 200 98 0.24 0.20 2.79 Var-59 MHT-004 0.31 0.37 0.5 340 0.12 < 0.03 150 132 0.34 0.32 2.74 Date: January 2013 Page 28 of 92
Variability Composite Drill Hole Au 1 Au 2 Ag As C TOTAL C ORGANIC Cu CN Sol. Cu S TOTAL S SULFIDE (g/t) (g/t) (g/t) (ppm) (%) (%) (ppm) (ppm) (%) (%) TRUE SG Var-60 MHT-004 0.25 0.26 0.4 140 0.18 0.06 136 54 0.36 0.30 2.74 Var-61 MHT-004 0.51 0.51 0.6 110 0.105 0.03 154 82 0.50 0.46 2.79 Var-62 MHT-004 0.58 0.54 0.5 150 0.12 < 0.03 238 194 0.84 0.70 2.77 Var-63 MHT-004 0.69 0.64 0.9 90 0.09 < 0.03 342 72 1.18 0.98 2.82 Var-64 VB11-003 0.5 0.56 0.3 30 0.21 0.03 130 192 0.98 0.78 2.80 Var-65 VB11-003 0.53 0.59 0.4 30 0.27 0.09 178 214 0.68 0.54 2.80 Var-66 VB11-003 1.06 1.00 0.6 30 0.21 0.03 236 244 0.98 0.78 2.77 Var-67 VB11-003 0.94 0.92 0.3 50 0.12 < 0.03 336 86 1.12 0.98 2.83 Var-68 VB11-003 1.36 1.25 0.6 30 0.12 < 0.03 398 110 1.10 0.90 2.81 Var-69 VB11-003 1.62 1.47 0.6 20 0.12 < 0.03 596 118 1.42 1.10 2.79 Var-70 VB11-003 2.68 2.72 0.5 10 0.09 < 0.03 544 234 1.82 1.32 2.83 Var-71 VB11-003 1.11 1.15 < 0.3 80 0.09 < 0.03 226 48 0.88 0.64 2.76 Var-72 VB11-003 0.45 0.43 < 0.3 20 0.06 < 0.03 208 40 0.62 0.54 2.78 Var-73 VB11-003 0.60 0.52 0.7 120 0.30 0.06 530 204 0.70 0.62 2.76 Var-74 VB11-003 0.48 0.46 < 0.3 20 0.12 < 0.03 246 94 0.70 0.54 2.79 Var-75 VB11-003 0.19 0.22 0.4 40 0.06 < 0.03 184 72 0.44 0.34 2.75 Var-76 VB11-003 0.48 0.43 0.5 70 0.09 < 0.03 256 84 0.94 0.80 2.82 Var-77 VB11-003 0.44 0.40 0.5 80 0.09 < 0.03 238 86 0.74 0.50 2.75 Var-78 VB11-003 0.38 0.42 1.2 40 0.12 < 0.03 468 172 1.00 0.84 2.82 Var-79 VB11-003 0.13 0.13 0.5 40 0.12 < 0.03 140 74 0.36 0.26 2.74 Date: January 2013 Page 29 of 92
Variability Composite Drill Hole Au 1 Au 2 Ag As C TOTAL C ORGANIC Cu CN Sol. Cu S TOTAL S SULFIDE (g/t) (g/t) (g/t) (ppm) (%) (%) (ppm) (ppm) (%) (%) TRUE SG Var-80 VB11-009 0.47 0.58 < 0.3 60 < 0.03 < 0.03 238 204 0.82 0.72 2.81 Var-81 VB11-009 0.61 0.61 < 0.3 40 0.03 < 0.03 230 184 0.82 0.70 2.76 Var-82 VB11-009 0.38 0.34 < 0.3 30 0.06 < 0.03 162 134 0.78 0.68 2.82 Var-83 VB11-009 0.18 0.19 < 0.3 30 0.03 < 0.03 148 80 0.70 0.62 2.75 Var-84 VB11-009 0.54 0.49 0.4 20 0.09 < 0.03 304 134 0.86 0.74 2.80 Var-85 VB11-009 0.54 0.48 0.4 20 0.06 < 0.03 258 100 1.04 0.90 2.79 Var-86 VB11-009 1.26 1.44 0.4 10 0.06 < 0.03 460 76 1.30 1.10 2.83 Var-87 VB11-009 0.85 0.67 0.6 20 0.03 < 0.03 824 140 2.10 1.60 2.79 Var-88 VB11-009 1.25 1.21 0.5 130 0.06 < 0.03 516 124 1.42 1.24 2.81 Var-89 VB11-009 0.53 0.4 0.7 20 0.06 < 0.03 934 304 1.60 1.40 2.37 Var-90 VB11-010 0.38 0.38 < 0.3 40 0.03 < 0.03 176 118 0.78 0.70 2.79 Var-91 VB11-010 0.64 0.62 0.3 20 < 0.03 < 0.03 220 152 0.82 0.74 2.80 Var-92 VB11-010 0.42 0.41 0.5 50 0.03 < 0.03 222 124 0.82 0.80 2.76 Var-93 VB11-010 0.19 0.18 < 0.3 30 0.03 < 0.03 144 54 0.70 0.60 2.79 Var-94 VB11-010 0.30 0.29 0.4 20 0.09 < 0.03 210 86 0.84 0.76 2.79 Var-95 VB11-010 0.60 0.59 0.7 20 0.06 < 0.03 532 186 1.82 1.44 2.85 Var-96 VB11-010 0.67 0.58 0.4 20 0.06 < 0.03 274 68 1.06 0.88 2.78 Var-97 VB11-010 0.73 0.56 0.5 40 0.06 < 0.03 702 174 1.34 1.12 2.75 Var-98 VB11-010 0.81 0.66 1 30 0.09 < 0.03 738 212 2.58 2.20 2.85 Var-99 VB11-010 0.33 0.38 0.3 10 < 0.03 < 0.03 536 112 1.48 1.24 2.77 Date: January 2013 Page 30 of 92
1.5.3 Cyanide Soluble Copper Assays A cyanide soluble copper assay was undertaken for all composites using bottle roll leach tests at an elevated NaCN concentration of 5,000 ppm for a duration of 7.5 hours. The NaCN concentration used in this test was much higher than will be experienced in the Mt Todd process plant and therefore these results are comparative only and do not reflect the expected level of copper leaching. At the less intense cyanide concentrations proposed for the Mt Todd operation, not all of the cyanide soluble copper will leach into solution, as has been demonstrated in the 2011 leach tests, summarised in Figure 3. Figure 3 shows that the actual cyanide soluble copper, at the leach conditions outlined in Section 1.8.6 and with a 90 µm grind size, is considerably lower than the assays listed in Table 14. Figure 3 - Cyanide soluble copper relationship 1.6 Mineralogy Six variability core samples were subjected to semi-quantitative mineralogical analysis to determine gold deportment and identify other minerals present in the ore. 1.6.1 Quantitative Mineralogy The quartz vein running through each of the core samples was isolated and thin section sized rectangular blocks containing this vein were cut from the received core samples. The sections were then suitably prepared for QEMSCAN3 analysis for quantitative mineralogy and Zeiss SEM Particle Scanner for the gold mineral search. The major mineral abundances for the six variability samples analysed are summarised in Table 15. 3 Quantitative Evaluation of Minerals by Scanning Electron Microscopy Date: January 2013 Page 31 of 92
Table 15 - Mt Todd Mineral Abundance Mineral Grouping Abundance (%w/w) Var-16 Abundance (%w/w) Var-19 Abundance (%w/w) Var-26 Abundance (%w/w) Var-33 Abundance (%w/w) Var-39 Abundance (%w/w) Var-91 Gold 0.09 0.00 0.00 0.00 0.00 0.02 Silver 0.00 0.00 0.00 0.00 0.00 0.00 Bismuth 0.09 0.00 0.00 0.00 0.00 0.02 Pyrite 16.79 37.33 0.80 4.04 13.30 23.50 Pyrrhotite 0.23 0.03 16.82 0.12 0.41 0.46 Chalcopyrite 0.94 0.06 0.84 0.07 0.06 0.09 Sphalerite 0.00 0.01 0.14 0.08 0.00 0.00 Galena 0.03 0.18 0.02 0.04 0.04 0.00 Arsenopyrite 0.06 0.00 0.00 0.00 0.00 0.00 Quartz 75.50 56.35 50.44 12.16 26.99 57.58 Mica/Clays 6.15 0.13 30.66 70.68 58.00 14.53 Carbonates 0.02 5.90 0.02 2.12 0.64 0.10 Fe Oxides 0.07 0.01 0.09 1.62 0.09 3.62 Minor Silicates 0.00 0.00 0.04 7.00 0.24 0.02 Minor Phases 0.02 0.01 0.13 2.08 0.24 0.05 Total 99.99 100.01 100.00 100.01 100.01 99.99 The QEMSCAN analysis highlights the following: the indicative mineralogy of the samples shows a dominance of quartz and mica/clays in all samples, which is expected since the quartz vein was targeted several sulfides were detected in the samples, including pyrite, pyrrhotite, chalcopyrite, sphalerite and galena pyrite is the dominant sulphide except in Var-26 where a large pyrrhotite grain was detected in the measurement area, which is contrary to the conclusions of the 2010 feasibility study chalcopyrite shows a significant occurrence in Var-16 and Var-26, which are both sourced from the same drill hole VB-11 sphalerite and galena were detected only in trace amounts the manual SEM EDS analysis of the detected gold grains during the gold search found that the majority of the gold occurred as both pure gold and argentian gold the mineralogical data above should be taken only as indicative, as measurement of a single 2D cross-sectional surface cannot produce statistically robust mineral populations. Date: January 2013 Page 32 of 92
1.6.2 Gold Mineralogy Gold mineralogy was determined by identifying gold grains using a Zeiss SEM particle scanner. Selected grains were then validated using x-ray spectrum spot analyses (Energy Dispersive Detector). Var-16 showed nuggety" gold disseminated throughout the quartz vein, with metallic bismuth also intercepted with the gold Var-19 revealed only a single argentian gold grain of ~20 µm within the quartz vein, with no further gold grains detected during the field scan, perhaps due to the small number and size of the gold grains present in the sample Var-26 detected only a single gold grain intergrown with pyrrhotite/pyrite in close association with metallic bismuth. The field scan detected a few pixels of gold along the fracture running through the quartz vein no gold was detected during the gold search analysis of either Var-33 or Var-39, with both samples revealing heavy pyrite mineralisation along the quartz vein single pixel sized gold grains were detected in Var-91, closely associated with metallic bismuth. A summary of gold deportment is shown in Table 16. Table 16 - Gold in Batman samples Sample Type Size (µm) Liberation Au mineral / Host Mineral Var-16 Core 27 Free Native gold Var-19 Core 20 Free Native gold Var-26 Core 25 Free Native gold Var-33 Core 0 - - Var-39 Core 0 - - Var-91 Core 26 Free Native gold The following conclusions are drawn for gold mineralization assessment: the grain sizes represent only the upper limit as the analysis resolution of 15 µm means the minimal reportable grain size is ~22 µm, which supports the findings of the gravity recovery testwork the results should be used only as a qualitative characterisation of gold due to the selectivity of the core slice used in the analysis the gold mineralisation for all samples occurred within the quartz vein, or close to the boundary between the vein and the host rock the majority of the gold occurred as both pure gold and argentian gold. Date: January 2013 Page 33 of 92
1.6.3 Particle and Grain Size The particle and grain size was calculated for each of the mineral groupings. The results are summarised in Table 17. Mineral Grouping Table 17 - Average particle and grain sizes Average Particle and Grain Sizes (µm) Var-16 Var-19 Var-26 Var-33 Var-39 Var-91 Gold 27 0 25 0 0 26 Silver 26 23 0 0 0 0 Bismuth 25 0 25 0 0 24 Pyrite 297 843 91 277 351 260 Pyrrhotite 44 37 336 35 54 29 Chalcopyrite 232 41 118 53 46 66 Sphalerite 48 36 91 31 40 35 Galena 43 47 33 27 76 0 Arsenopyrite 102 45 34 32 106 72 Quartz 729 2,143 159 32 81 265 Mica/Clays 78 87 92 119 157 68 Carbonates 45 439 45 49 89 35 Fe Oxides 59 52 47 25 29 103 Minor Silicates 34 31 32 28 26 37 Minor Phases 43 34 34 24 29 38 Comments include: the calculated average grain size values should be taken as only indicative as the size of the bulk material wall rock and vein is over estimated and does not occur as grains the pyrite and pyrrhotite are very coarse relative to other minerals the chalcopyrite in Var-16 and Var-26 is very coarse while the data suggests coarse mica/clays, the textural data suggests that the mica/clays are fine grained. Due to the resolution of the analysis it was not possible to differentiate individual grains. 1.7 Comminution Testwork Twenty comminution samples were selected (refer Section 1.4.2) for the following tests: SMC Test Bond ball mill work index (BWi). The testwork was complemented by: BWi determinations on the master composite sample at different closing screen sizes BWi comparison between conventionally crushed material and HPGR product BWi determinations on selected leach variability samples at various closing screen sizes Unconfined compressive strength (UCS) tests on selected samples crushing work index test (CWi) on eight of the comminution samples. Date: January 2013 Page 34 of 92
1.7.1 SMC Test The SMC Test results are summarized in Table 18, presenting the two properties used in comminution circuit design: the JK drop-weight test parameter (Axb) and the apparent ore specific gravity. The values for other parameters measured, drop-weight index (DWI), coarse ore work index (Mia), high pressure grinding rolls work index (Mih), crushing ore work index (Mic) and JK abrasion parameter (ta) are recorded in the JKTech Pty Ltd report 11001/P71 SMC Test Report on Twenty Samples from Mt Todd Project, which appears as an appendix in the ALS Ammtec testwork report, found in Appendix 1.2. Table 18 - SMC Test results Sample Composite No. SG Axb DWi Category Comminution 1 Var-04 2.49 21.0 11.7 very hard Comminution 2 Var-10 2.79 26.8 10.4 very hard Comminution 3 Var-13 2.8 24.0 11.8 very hard Comminution 4 Var-24 2.78 21.9 12.9 very hard Comminution 5 Var-27 2.78 21.0 12.9 very hard Comminution 6 Var-28 2.80 26.5 10.6 very hard Comminution 7 Var-42 2.81 22.1 12.8 very hard Comminution 8 Var-47 2.82 21.4 13.0 very hard Comminution 9 Var-52 2.74 25.3 10.8 very hard Comminution 10 Var-55 2.79 19.0 14.7 very hard Comminution 11 Var-56 2.78 30.8 9.0 hard Comminution 12 Var-57 2.7 24.6 10.9 very hard Comminution 13 Var-58 2.73 27.9 9.7 very hard Comminution 14 Var-60 2.75 25.7 10.8 very hard Comminution 15 Var-61 2.77 25.7 10.1 very hard Comminution 16 Var-62 2.77 27.7 12.1 very hard Comminution 17 Var-63 2.80 23.1 10.5 very hard Comminution 18 Var-65 2.77 26.7 13.29 very hard Comminution 19 Var-69 2.80 24.3 11.68 very hard Comminution 20 Var-74 2.75 25.4 10.96 very hard Maximum 2.82 30.8 14.7 Minimum 2.49 19.0 9.0 Average 2.76 24.6 11.53 75 th Percentile 2.80 22.1 12.8 Date: January 2013 Page 35 of 92
Table 18 shows the Batman ore is consistently very competent throughout the ore body, with the following key points: average SG of 2.76 75th percentile Axb value of 22.1 75th percentile DWi value of 12.8 kwh/m³. 1.7.2 Crushing Work Index Testwork on Selected Variability Composites The results of the CWI tests on selected Batman variability samples are presented in Table 19. Table 19 - CWI Results Composite No. CWI ( kwh/t) Var-04 7.5 Var-10 5.0 Var-13 4.7 Var-27 7.4 Var-42 5.3 Var-52 5.8 Var-57 8.4 Var-59 6.1 Table 19 includes: maximum value of 8.4 kwh/t minimum value of 4.7 kwh/t average value of 6.3 kwh/t 75th percentile value of 7.4 kwh/t. 1.7.3 Unconfined Compressive Strength Tests on Selected Samples The results of the UCS tests on samples from the MHT-001 and MHT-004 drill holes are presented in Table 20. Date: January 2013 Page 36 of 92
Table 20 - UCS Results Drill Hole Composite Number UCS (MPa) MHT-001 Var-56 127.1 MHT-001 Var-56 64.4 MHT-001 Var-57 69.8 MHT-001 Var-57 82.3 MHT-001 Var-58 39.4 MHT-001 Var-58 70.3 MHT-004 Var-58 120.0 MHT-004 Var-59 103.6 MHT-004 Var-59 131.8 MHT-004 Var-60 91.9 MHT-004 Var-60 174.2 MHT-004 Var-61 104.8 MHT-004 Var-61 182.3 MHT-004 Var-62 13.5 MHT-004 Var-62 54.9 MHT-004 Var-63 88.2 MHT-001 Var-63 127.1 Table 20 includes: maximum value of 182.3 kwh/t minimum value of 13.2 kwh/t average value of 94.9 kwh/t 75th percentile value of 121.7 kwh/t. 1.7.4 Bond Ball Mill Work Index Tests on Comminution Composites The results of the BWi tests on the Batman comminution samples at a 106 µm closing screen are presented in Table 21. Date: January 2013 Page 37 of 92
Table 21 - BWi Results (106 µm closing screen) Sample Composite No. BWI ( kwh/t) Comminution 1 Var-04 25.0 Comminution 2 Var-10 24.8 Comminution 3 Var-13 27.7 Comminution 4 Var-24 24.4 Comminution 5 Var-27 26.6 Comminution 6 Var-28 27.7 Comminution 7 Var-42 24.6 Comminution 8 Var-47 26.3 Comminution 9 Var-52 28.0 Comminution 10 Var-55 25.5 Comminution 11 Var-56 27.4 Comminution 12 Var-57 28.1 Comminution 13 Var-58 27.6 Comminution 14 Var-60 27.8 Comminution 15 Var-61 28.2 Comminution 16 Var-62 25.1 Comminution 17 Var-63 26.7 Comminution 18 Var-65 24.6 Comminution 19 Var-69 23.6 Comminution 20 Var-74 24.9 Table 21 shows that there is little variability between the comminution samples. All samples displayed above average hardness: maximum value of 28.2 kwh/t minimum value of 23.6 kwh/t average value of 26.2 kwh/t 75th percentile value of 27.7 kwh/t. Date: January 2013 Page 38 of 92
BWI, kwh/t TESTWORK REPORT FOR THE MT TODD GOLD PROJECT FEASIBILITY STUDY 1.7.5 Bond Ball Mill Work Index Tests on Master Composite at Various Closing Screen Sizes The results of the BWi tests on the master composite at different closing screen sizes are shown in Table 22 and represented graphically in Figure 4. Table 22 - BWI at different closing screen sizes for the master composite sample Sample Closing screen size (µm) BWi ( kwh/t) Master composite 75 24.1 Master composite 106 25.1 Master composite 150 26.2 Master composite 212 28.0 Master composite 300 30.5 Figure 4 - Closing screen size versus average BWi for master composite sample 32 30 28 26 24 22 20 0 50 100 150 200 250 300 350 Closing Screen Size, µm The plot of Bond Ball Mill work index ( BWi ) versus closing screen aperture shown in Figure 4 indicates that the Mt Todd Batman ore exhibits an unconventional trend whereby the BWi values decrease as the closing screen size (and the corresponding grind size) decreases. More typically, the BWi increases with fineness of grind (decreasing closing screen size). This atypical relationship of BWi to fineness of grind has significant implications for modelling of the comminution circuit and the determination of the optimum grind size. While unconventional, this relationship has been observed previously by ALS Ammtec at several operations in the Eastern Goldfields of Western Australia and therefore is not unique to the Batman ore. For this reason an additional phase of test work was undertaken on available variability samples to determine if this inverse trend was consistent across the Batman deposit. Date: January 2013 Page 39 of 92
1.7.6 Additional Bond Ball Mill Work Index Tests on Variability Composites The objective of the additional BWi tests on the variability composite samples at various closing screen apertures was to strengthen the database of BWi results and to determine whether the unconventional trend observed in Figure 4 is consistent across the Batman deposit. The results of this test work are summarised in Table 23. Table 23 - BWI at different closing screen sizes for various variability samples Closing Screen Size μm 75 106 125 150 180 212 300 P 80, μm 53 75 90 106 125 150 212 Var-4 - - 24.2 24.4 25.0 25.3 - Var-10 - - - - 24.8 - - Var -13 - - - - 26.8 29.1 - Var -24 - - - 23.4 24.4 25.0 - Var -27 - - 26.3 26.5 27.8 28.5 - Var -28 25.9 26.1 26.6 27.2 26.7 27.7 29.9 Var - 42 - - - - 25.4 - - Var - 47 - - - 27.0 27.6 - - Var - 52 - - - - 27.3 - - Var - 55 22.9 23.7 23.5 24.2 25.8 26.8 28.2 Var - 56 - - - 27.9 28.3 29.0 - Var - 57 - - - 27.6 27.9 28.5 - Var - 58 - - - 27.6 27.4 28.8 - Var - 60 - - - 25.8 25.9 27.4 - Var - 61 - - - - 27.4 - - Var - 62 - - - 26.5 25.1 26.3 - Var - 65 - - - 24.5 24.4 26.0 - Var - 69 - - - 23.6 24.6 25.1 - Var - 74 - - 23.5 23.6 23.8 24.6 26.5 Min 22.9 23.7 23.5 23.4 23.8 24.6 26.5 Max 25.9 26.1 26.6 27.6 28.3 29.1 29.9 Ave 24.4 24.9 24.8 25.7 26.1 27.0 28.2 75th Percentile 25.2 25.5 26.3 27.2 27.4 28.5 29.0 The average BWI for each closing screen size is plotted in Figure 5. Date: January 2013 Page 40 of 92
Average BWI, kwh/t TESTWORK REPORT FOR THE MT TODD GOLD PROJECT FEASIBILITY STUDY Figure 5 - Closing screen size versus average BWi for variability samples 29.0 28.5 28.0 27.5 27.0 26.5 26.0 25.5 25.0 24.5 24.0 0.0 50.0 100.0 150.0 200.0 250.0 Closing Screen Size, µm Figure 5 confirms a strong trend whereby BWi increases with closing screen size. 1.7.7 Comminution Testwork on HPGR Product BWi results on MHT-001 and MHT-004 HPGR Product Samples, Polysius Further comminution test work was completed by Polysius to determine the effect of high pressure grinding on the BWi. High pressure grinding often produces micro-cracks in the progeny particles, reducing the overall particle strength and generating a greater distribution of fine material in the ball mill feed which in turn reduces the downstream ball mill energy requirements. In theory, this should result in a reduced BWi for material that has been treated in a HPGR compared to conventionally crushed product. The results of this testwork are summarised in Table 24, comparing the BWi of HPGR product against a conventionally crushed product. Feed and product particle size distributions for the MHT-001 and MHT-004 samples are shown in Figure 6 and Figure 7. Note that this testwork was undertaken using a 200 µm closing sieve, before the determination of the optimum grind size. Table 24 - Bond Work Index (200 µm closing screen), HPGR Product vs. Conventionally Crushed feed, Polysius Sample Conventional Crushing BWi (kwh/t) HPGR BWi (kwh/t) MHT-001 27.1 25.1 MHT-004-25.7 24.5 Date: January 2013 Page 41 of 92
Figure 6 - MHT-001 Feed and Product Size Distributions Figure 7 - MHT-004 Feed and Product Size Distributions Date: January 2013 Page 42 of 92