Royal Nickel Corporation A PRELIMINARY ASSESSMENT OF THE DUMONT PROPERTY LAUNAY AND TRÉCESSON TOWNSHIPS, QUEBEC, CANADA. 30 September, 2010

Transcription

1 Royal Nickel Corporation A PRELIMINARY ASSESSMENT OF THE DUMONT PROPERTY LAUNAY AND TRÉCESSON TOWNSHIPS, QUEBEC, CANADA 30 September, 2010 William J. Lewis, B.Sc., P.Geo. Ing. Alan J. San Martin, MAusIMM Richard M. Gowans, P.Eng. David Penswick, P.Eng. Michel Lemieux, Eng., M.Sc. Pierre Primeau, P.Eng. Colin Hardie, P.Eng. SUITE BAY STREET, TORONTO ONTARIO, CANADA M5H 2Y2 Telephone (1) (416) Fax (1) (416)

2 Table of Contents Page 1.0 SUMMARY INTRODUCTION GEOLOGY AND MINERAL RESOURCES MINING AND GEOTECHNICAL ISSUES Geotechnical Issues Open Pit Design Production Schedule Mining Fleet Requirements METALLURGICAL TESTWORK PROCESS AND INFRASTRUCTURE DESIGN Proposed Flowsheet Process Design Criteria Infrastructure TAILINGS MANAGEMENT FACILITY ENVIRONMENTAL ASPECTS CAPITAL COST ESTIMATE OPERATING COST ESTIMATE EVALUATION Base Case and Upside Case Sensitivity Analysis CONCLUSIONS RECOMMENDATIONS Proposed Exploration Program Further Recommendations INTRODUCTION AND TERMS OF REFERENCE RELIANCE ON OTHER EXPERTS PROPERTY DESCRIPTION AND LOCATION DUMONT PROPERTY MINERAL CLAIM AGREEMENTS Griffis International Ltd. Mineral Claims Marbaw International Nickel Corporation Mineral Claims Sheridan-Ferderber Mineral Claims Royal Nickel Claims ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY HISTORY ROYAL NICKEL EXPLORATION HISTORICAL MINING AND PRODUCTION DUMONT PROPERTY RESOURCE AND RESERVE ESTIMATES i

3 7.0 GEOLOGICAL SETTING REGIONAL GEOLOGICAL SETTING CONTACT RELATIONS, AREAL EXTENT AND AGE OF THE DUMONT INTRUSION DEPOSIT TYPES MINERALIZATION DISSEMINATED NICKEL MINERALIZATION Nickel Mineralogy CONTACT-TYPE NICKEL-COPPER-PGE MINERALIZATION OTHER TYPES OF PGE MINERALIZATION EXPLORATION EXPLORATION PROGRAM Resource Definition Drilling Structural Drilling and Modelling Geotechnical Drilling and Studies Pilot Plant Test Drill Holes (NQ) Geotechnical (Overburden) Drill Holes Pilot Plant Sample Drill Holes (PQ) Geological Mapping Mineralogical Sampling Overburden Modelling EXPLORATION RESULTS Results of the Geotechnical (Overburden) Drilling Program Pilot Plant Test Drill Holes (NQ) Results Results of the Pilot Plant Sample Drill Holes (PQ) Results Results of the Mineralogical Sampling Program Results of the Overburden Modelling Program Results of the Crushing Testwork Hardness Domain Sampling DRILLING RESOURCE DEFINITION DRILLING PROGRAM RESULTS OF THE 2010 SECTIONAL DRILLING PROGRAM SAMPLING METHOD AND APPROACH ASSAY/GEOCHEMICAL SAMPLING MINERALOGICAL MAPPING SAMPLING PILOT PLANT SAMPLING SAMPLE PREPARATION, ANALYSES AND SECURITY SAMPLE COLLECTION AND TRANSPORTATION Core Logging and Sampling Sample Preparation and Analysis Control Samples ii

4 13.2 MINERALOGICAL MAPPING SAMPLING Sample Definition and Sampling Sample Preparation and Analysis Control Samples PQ DRILLING RESULTS OF THE QA/QC PROGRAM MICON COMMENTS ON THE QA/QC PROGRAM DATA VERIFICATION ADJACENT PROPERTIES MINERAL PROCESSING AND METALLURGICAL TESTING OVERVIEW TECHNOLOGY PARTNERS PHASE 2 TESTWORK Comminution Testwork Nickel Recovery Testwork RECOVERY EQUATIONS CONCLUSIONS FUTURE WORK MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES MICON UPDATED 2010 RESOURCE ESTIMATE FOR THE DUMONT PROPERTY Modelling Methodology Compositing Data Rock Density Assay Sample Nickel Grade Statistics Variography Data Interpolation Strategy Inside the Resource Block Model Model Validation and Post Processing Future Work MINERAL RESOURCE CLASSIFICATION MINERAL RESOURCE ESTIMATES Economic Discussion and Cut-off Grade August, 2010 Mineral Resource Estimates DUMONT PROPERTY EXPLORATION POTENTIAL OTHER RELEVANT DATA AND INFORMATION GEOTECHNICAL WORK Open-Pit Rock Mechanics Open-Pit Soil Mechanics Waste Impoundments MINE DESIGN Overview iii

5 Techno-Economic Model LG Ultimate Pit Shell Detailed Mine Design Base Case Design MINING Overburden Mining Rock Mining Equipment Maintenance Mining Fleet Selection Support Equipment PROCESSING Design Criteria Primary and Secondary Crushing Drying Tertiary and Quaternary Crushing Grinding Ferro-Nickel Circuit Nickel Sulphide Circuit Scavengers Concentrate Dewatering and Load-Out Tailings Thickening Product Quality INFRASTRUCTURE Site Layout Off-Site Infrastructure Logistics MANPOWER ENVIRONMENTAL STUDIES Environmental Management Plan ENVIRONMENTAL PERMITTING MARKETING Nickel Sulphide Concentrate Ferro-nickel Concentrate Commercial Terms SCHEDULE CAPITAL COST ESTIMATE Summary Mining Capital Processing Capital Tailings Management Facility Infrastructure Capital Indirect Capital Initial Capital Contingency Sustaining Capital Working Capital iv

6 18.12 OPERATING COST ESTIMATE Summary Labour Energy Mining Operating Costs Processing Operating Costs General and Administrative Operating Costs EVALUATION Key Assumptions Economic Results Sensitivity Analysis INTERPRETATION AND CONCLUSIONS RECOMMENDATIONS EXPLORATION PROGRAM FURTHER RECOMMENDATIONS REFERENCES SIGNATURE PAGE CERTIFICATES v

7 List of Tables Page Table 1.1 Summary of the Mineral Resource Estimate at a Cut-off Grade of 0.20% Nickel... 2 Table 1.2 In-Pit Diluted Portion of the Mineral Resources... 4 Table 1.3 Mining Fleet Base Case... 5 Table 1.4 Average Composite Sample Parameters and Estimated Nickel Recoveries... 6 Table 1.5 Capital Cost Summary Table 1.6 Estimated Site Operating Cost Summary Table 1.7 Summary Metrics Table 1.8 Sensitivity of NPV to a Selection of Discount Rates Table 2.1 Participants in Dumont Preliminary Assessment Table 2.2 List of Abbreviations Table 9.1 Nickel Bearing Mineral Abundance by Mineralization Type Table 10.1 Summary of the Royal Nickel Drilling Programs on the Dumont Property by Category Table 10.2 Dumont Property 2010 Exploration Expenditures Table 10.3 Summary for the Oriented Structural Drill Holes Table 10.4 Summary for the Pilot Plant Sample Drill Holes (PQ) Program Table 10.5 Summary of Principal Minerals in Mineralogical Samples Received to August 16, Table 10.6 Summary of the Drop-Weight Breakage Evaluation Table 10.7 Summary of the SMC Break Evaluation Table 10.8 Summary of the UCS Results Table 11.1 Sectional Drilling Program Drill Hole Collar Information Table 11.2 Table 12.1 Table Drilling Program Mineralized Intersections at a 0.25% Nickel Cut-off Grade EXPLOMIN TM Mineralogical Sample Preparation Procedure at ALS- Chemex Laboratories Summary of the Specifications for the Standard Reference Material samples Table 13.2 SGS Lakefield Daily Quality Checks for Qemscan Analysis Table 16.1 Phase 2 Metallurgical Testwork Composite Samples vi

8 Table 16.2 Average Composite Sample Parameters and Estimated Nickel Recoveries Table 16.3 Metallurgical Testwork Results From Domain Composite Samples Table 17.1 List of Interpreted Structural Domains for the Dumont Project Table 17.2 Block Model versus Solid Volume Check Table 17.3 Missing Assays by Structural Domain within the Resource Wireframe Table 17.4 Statistics for Results for Nickel Raw Samples from All Domains 1 (D1) to 7 (D7) Table 17.5 Statistics for Results for Density Raw Samples from All Domains Domain 1 (D1) to 7 (D7) Table 17.6 Experimental Variogram Parameters Table 17.7 Variogram Models Used for Percent Nickel Table 17.8 Search Parameters used for the Ordinary Kriging Interpretation Table 17.9 Example Corrections for Over-Smoothing in Awaruite Content Table Dumont Cut-off Grade Calculation Table Table Table Summary of the Dumont Mineral Resources at a 0.20% Nickel Cutoff Grade in all Domain Solids by Category and Structural Domain: Cumulative Summary of the Dumont Mineral Resources Between the 5500 and 9100 Section Lines at a 0.20% Nickel Cut-off Grade by Category and Structural Domain: Cumulative Summary of the Measured, Indicated and Inferred Mineral Resource in the Seven Structural Domain Solids at a Cut-off of 0.20% Nickel Table 18.1 Participants in Dumont Preliminary Assessment Table 18.2 Cost Inputs to the LG Algorithm Table 18.3 Nested Pit Shell Phases Table 18.4 Base Case Detailed Production Schedule (Years 1 15) Table 18.5 Base Case Detailed Production Schedule (Years 16+) Table 18.6 Mining Fleet Base Case Table 18.7 Capital Cost Summary Table 18.8 Mining Capital Cost Summary Base Case Table 18.9 Process Plant Capital Cost Summary Table Tailings Management Capital Cost Summary Table Infrastructure Capital Cost Summary vii

9 Table Indirect Capital Cost Summary Table Initial Capital Cost Contingency Table Sustaining Capital Cost Summary Table Estimated Site Operating Cost Summary Table Estimated Process Operating Cost Table Estimated General and Administrative Operating Cost Table Summary Metrics Table Annual Cash flow Base Case (80,000 t/d) Table Annual Cash flow Upside Case (100,000 t/d) Table Sensitivity of NPV to a Selection of Discount Rates Table 20.1 Proposed Budget for Work in 2011 and viii

10 List of Figures Page Figure 1.1 Simplified Metallurgical Flow Diagram... 6 Figure 1.2 Processing Flowsheet... 8 Figure 1.3 Base Case (80,000 t/d) LOM After-Tax Net Cash Flow Figure 1.4 Upside Case (100,000 t/d) LOM After-Tax Net Cash Flow Figure 1.5 Sensitivity of NPV - Base Case (80,000 t/d) Figure 1.6 Sensitivity of NPV - Upside Case (100,000 t/d) Figure 4.1 Dumont Property Location Map Figure 4.2 Dumont Property Mineral Claims Figure 4.3 Figure 5.1 Location Map Showing the Arctic-St. Lawrence Drainage Divide and Agricultural Zone Lands as they Relate to the Dumont Property Map Indicating Access and Geographical Features in the Area of the Dumont Property Figure 5.2 Access Road for the Dumont Property Figure 7.1 Magnetometer Survey of the Dumont Property Figure 7.2 Geological Map of the Dumont Sill Figure 7.3 Cross-Sectional View on Line 8100E Showing the Position of the Three Nickel-Enriched Layers Figure 9.1 Photo of the Dumont Mineralization in Core (Field of View = 5 cm) Figure 9.2 Sulphide Mineralization Type (EXP_18) Figure 9.3 Alloy Mineralization Type (EXP_43) Figure 9.4 Mixed Mineralization Type (EXP_15) Figure 9.5 Block Model for the Dumont Project Illustrating the Mineralization Type Distribution Figure 10.1 Geological Map Showing the Location of the Non-Resource Drilling Figure 10.2 Typical Domaining of a Drill Hole (09-RNC-218A-I) Figure 10.3 Drill Hole Locations with Mineralogical Mapping Sample Results and Planned Sampling Figure 10.4 Overburden Thickness Map Figure 11.1 Locations of the Resource Definition Drill Hole Collars on the Dumont Property Figure 11.2 View of a Rouillier Drill Rig (July, 2010 Site Visit) ix

11 Figure 11.3 Figure 11.4 Figure 11.5 Plan and Sectional View (7000E) Illustrating the Drill Hole Distribution and Assay Results Plan and Sectional View (7100E) Illustrating the Drill Hole Distribution and Assay Results Plan and Sectional View (7200E) Illustrating the Drill Hole Distribution and Assay Results Figure 12.1 Assay Sampling Protocol Figure 12.2 EXPLOMIN TM Sampling Protocol Figure 13.1 Core Logging Facilities in Amos Figure 14.1 Royal Nickel Office Core Logging Facilities in Amos Figure 14.2 Interior View of the Logging Facilities in Amos Figure 16.1 Simplified Metallurgical Flow Diagram Figure 16.2 Locations of Domain Composite Samples Figure 16.3 Figure 17.1 Cross-Section Showing a Typical Downhole Distribution of Domain Composite Samples Magnetic Intensity Map of the Dumont Project with Interpreted Lineaments Figure 17.2 Drilling Information used for the Mineral Resource Model Figure 17.3 Plot of SG versus Nickel Grade for Domain 1 at the Dumont Property Figure 17.4 Plot of SG versus Nickel Grade for Domain 2 at the Dumont Property Figure 17.5 Plot of SG versus Nickel Grade for Domain 3 at the Dumont Property Figure 17.6 Plot of SG versus Nickel Grade for Domain 4 at the Dumont Property Figure 17.7 Plot of SG versus Nickel Grade for Domain 5 at the Dumont Property Figure 17.8 Plot of SG versus Nickel Grade for Domain 6 at the Dumont Property Figure 17.9 Plot of SG versus Nickel Grade for Domain 7 at the Dumont Property Figure Figure Figure Figure Location of the Seven Structural Domain Solids Involved in the Mineral Resource Estimate and their Structural Boundaries (Grey) Block Model for the Dumont Project Illustrating the Grade Distribution along with the Drill Hole Distribution Block Model for the Dumont Project Illustrating the Distribution of the Mineralization Type with the Drill Hole Distribution Block Model for the Dumont Project Illustrating the Resource Categories with the Drill Hole Distribution Figure 18.1 Evaluation of Nested Shells Figure 18.2 LG Ultimate Pit Shell x

12 Figure 18.3 End of Mine Life Pit Faces Figure 18.4 Base Case Mining Schedule Figure 18.5 Base Case Schedule for Mill Feed Figure 18.6 Base Case Nickel Production Figure 18.7 Processing Flowsheet Figure 18.8 Site Layout at the End of Open Pit Operations Figure 18.9 Project Development Schedule Figure LOM Mining Unit Costs and One-Way Haulage Distance Figure Base Case LOM After-Tax Net Cash Flow (80,000 t/d) Figure Upside Case LOM After-Tax Net Cash Flow (100,000 t/d) Figure Sensitivity of NPV - Base Case (80,000 t/d) Figure Sensitivity of NPV - Upside Case (100,000 t/d) xi

13 1.0 SUMMARY 1.1 INTRODUCTION At the request of Royal Nickel Corporation (Royal Nickel), Micon International Limited (Micon) has prepared this Preliminary Assessment of Royal Nickel s Dumont Property located in western Quebec, Canada. The study is based on Micon s August, 2010 mineral resource estimate and comprises an open pit mine plan, production schedule, fleet requirements and designs for processing plant, tailings management facility and other required infrastructure. Mining, processing and infrastructural capital and operating costs are estimated and a discounted cash flow forecast is given. A base case evaluation of the property is presented, based on a processing throughput rate of 80,000 t/d. A second ( Upside ) case is also shown, based on a throughput of 100,000 t/d, utilizing the maximum electrical grid power available to the project on a planning basis. The Dumont Property is located in the western portion of Quebec approximately 25 km west of the city of Amos, 60 km northeast of the industrial and mining city of Rouyn-Noranda and 70 km northwest of the city of Val D Or. Royal Nickel advises that the Dumont Property consists of 138 contiguous mineral claims totalling 5, ha. Four sets (blocks) of claims are recognised: 50 claims staked by Royal Nickel itself, 24 claims purchased outright from Griffis International Ltd; 58 claims purchased from Marbaw International Nickel Corporation (subject to a future consideration and a retained NSR royalty) and 6 claims purchased from Sheridan-Ferderber (subject to a retained NSR royalty). The mineral claims confer subsurface rights only. Surface rights for approximately 25% of the property are held privately, the rest is public land. There are no known formal native land claims covering the Dumont Property. 1.2 GEOLOGY AND MINERAL RESOURCES The Dumont sill lies within the Abitibi subprovince of the Superior geologic province of the Archean age Canadian Shield. The sill is one of several mafic to ultramafic intrusive bodies that form an irregular, roughly east-west alignment, between Val d Or, Quebec and Timmins, Ontario. It comprises a lower ultramafic zone which averages 450 m in true thickness and an upper mafic zone about 250 m thick. The ultramafic zone is subdivided into the lower peridotite, dunite and upper peridotite subzones. Cumulus nickel sulphide and alloy minerals occur in parts of the dunite subzone and locally in the lower peridotite. Pentlandite ((Ni,Fe) 9 S 8 ), heazlewoodite (Ni 3 S 2 ) and awaruite (Ni 2.5 Fe) are the principal nickel minerals with lesser amounts of millerite (NiS). Historically, three mineralized layers have been identified within the dunite subzone using a 0.35% Ni cut-off grade. The middle layer has the highest average nickel grade of 0.5% and is the most laterally extensive, persisting over a strike length of 2,400 m with an average thickness of 24 m. A higher grade zone within the middle layer averages 0.71% nickel over a strike length of 730 m and has a true thickness of 14 m. 1

14 The property has been explored sporadically since the early 1970 s. Royal Nickel acquired the property in 2006 and since then has conducted its own exploration, including in-fill and step-out drilling, metallurgical testwork and engineering studies. The effective date of the current resource estimate is August 16, The estimate is based on the exploration database which contains a total of 70,577 m of assay results from 223 drill holes that Royal Nickel has obtained through its 2007 to 2010 drilling programs. The total metreage for the 223 holes available for the resource estimate is 90,212 m. Micon s estimate of the mineral resource is based on the geological information and assaying data for the Dumont Property available as of April 22, 2010, a structural model developed for Royal Nickel by Itasca Consulting, economic parameters developed in preparation for this preliminary assessment, and a resource block model prepared by Golder Associates that models the abundance of pentlandite, heazlewoodite, awaruite, olivine, magnetite, serpentine, brucite and coalingite, as well as the grade distributions of nickel, copper, cobalt, chromium, platinum, palladium and gold, and specific gravity. Micon reviewed the block model extensively and in some cases the model was refined in discussions with Royal Nickel. The final resource classification criteria remain the same as those used in previous reports. Table 1.1 Summary of the Mineral Resource Estimate at a Cut-off Grade of 0.20% Nickel (As of August 16, 2010) Mineral Resource Category Resource (000 t) Grade (% Ni) Nickel (000 t) Contained Ni (Million lbs) Measured 155, Indicated 1,003, ,707 5,967 Measured+Indicated 1,159, ,154 6,952 Inferred 581, ,451 3,198 Mineral resources that are not mineral reserves do not have demonstrated economic viability. There are no mineral reserves presently identified on the Dumont Property. The stated mineral resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues, unless stated in this report, to the best knowledge of the authors. The impact of mining, metallurgical, infrastructure and other factors that could materially affect this mineral resource estimate are described in Section 18 of this technical report. The mineral resource estimate as of the effective date of August 16, 2010 is compliant with the current CIM standards and definitions required by NI and is, therefore, reportable as a mineral resource by Royal Nickel. There is the potential to find additional mineralization on the Dumont Property. This potential includes extensions to the current zones of mineralization in the resource model both at depth and on the northwest and southeast ends of domains 1 and 7. Exploration either 2

15 at depth or on the flanks of the deposit may contribute further mineral resources to the Dumont Property. 1.3 MINING AND GEOTECHNICAL ISSUES The mining aspects of the preliminary assessment comprised studies of the safe slope angles for the open pit, calculation of the NSR value of the recoverable mineralisation within each element of the 3D block model of the resource, provisional estimates of unit operating costs for each material, open pit optimization to determine the economic limits of mining, scheduling of material movement to mill, stockpile or waste dumps so as to maximize NPV, and estimation of capital and operating costs for the required mining fleet Geotechnical Issues Genivar assessed the safe slope angles for the open pit in unconsolidated overburden and in rock. Genivar concluded that, based on its analysis of soil types at similar projects in the Dumont area, safe slope angles for the open pit walls in overburden would be: 3H:1V (18 ) in fine-grained material slope angles 2H:1V (27 ) in coarse-grained material Applying a factor of safety of 1.5, it was concluded that in rock, safe slope angles would be: 47 on the hanging wall of the open pit 52 on the footwall of the open pit Impoundments for waste rock will have slopes of 24 (2.25H:1V), composed of 15 m lifts with face angles equivalent to the angle of repose of approximately 34 (1.5H:1V) with 11 m berms. The impoundments will reach a maximum height of 10 lifts, or 150 m Open Pit Design In order to determine the economic limits of open pit mining, NPV Scheduler software was used to apply the Lerchs-Grossmann (LG) algorithm to the recoverable NSR values and provisional unit cost estimates for each element in the 3D resource block model, taking account of the slope angles given above. NSR values were based on recovery equations derived from regression analysis of results from test work on representative composite samples. The equations also considered expected differences in the metallurgical response of the three styles of mineralization: Sulphide mineralization, mainly pentlandite and heazlewoodite Alloy mineralization, dominated by the alloy mineral awaruite Mixed mineralization, containing significant levels of sulphides and alloy styles 3

16 The recovery equations allowed the recovery of Ni to nickel sulphide and ferro-nickel concentrates to be estimated discretely for every block in the resource model. The recovered nickel was then combined with assumptions regarding percentage payables for the different concentrates, treatment and refining charges (TC/RCs) and the long-term Ni price to yield an NSR value for each block. The LG algorithm produces, in addition to the ultimate limits for the open pit, a series of nested shells used to guide the sequence of mining. The detailed pit design incorporates a high-grade, low-strip-ratio starter pit and 5 pushbacks into progressively lower-grade and/or higher-strip-ratio material. The optimal parameters for the mining schedule including milling rate, low-grade stockpile strategy and cut-off grade were selected through iteration. The optimized NSR cut-off value is C$14.00/t. Table 1.2 summarizes the in-pit diluted portion of the mineral resources included in the mine plan. Note that the scoping-level engineering design is not sufficient to justify classification of this material as a mineral reserve. No value was attributed to the inferred resource within the open pit: for the purpose of the preliminary assessment it was treated as waste rock. Table 1.2 In-Pit Diluted Portion of the Mineral Resources Mineral Resource Category Resource (000 tonnes) Grade (% Ni) Nickel (000 t) Contained Ni (million lbs) Measured 135, Indicated 760, ,071 4,565 Measured + Indicated 896, ,458 5,418 Inferred 29, Production Schedule Through analysis of multiple schedules reflecting different rates of mining and processing, an optimal schedule has been arrived at which accelerates mining to a rate which exceeds that required to keep the mill supplied. In doing so, the opportunity arises to selectively feed the mill with higher value material while stockpiling the rest. Once the open pit has been completed, it can be used as a repository for tailings arising from milling of the stockpiled material. This strategy results in a higher NPV than one in which the mining rate is matched to the milling rate Mining Fleet Requirements The mining fleet requirement is summarized in Table 1.3. In addition to purchase of the new units identified here, provision has been made in the sustaining capital estimate for rebuilding some of this equipment. 4

17 Table 1.3 Mining Fleet Base Case Overburden Rock Size Example No. Reqd Size Example No. Reqd Unit Init. Sust. Init. Sust. Rotary Drill n/a mm Ø P&H XP Production FEL 13 m 3 Cat m 3 LeTorneau L Rope Shovel n/a - 55 m 3 P&H 4100 XPC Haul Truck 100 t Cat t Cat Track Dozer 13 m 3 Cat D m 3 Cat D Rubber Tyre Dozer 8 m 3 Cat m 3 Cat Grader 16-ft blade Cat 16M ft blade Cat 24M Water Tanker 40 t Cat t Cat One additional unit is required for the Upside Case. - One less unit is required for the Upside Case 1.4 METALLURGICAL TESTWORK Initial metallurgical testwork was carried out on samples from the Dumont Property in , and focused on grinding, flotation and magnetic separation but achieved an overall nickel recovery of around 55%. Royal Nickel s testwork commenced with Phase 1 in at SGS Minerals in Lakefield, Ontario under the management of Royal Nickel s independent metallurgical consultants, Mineral Solutions. Tests were performed on composite samples representing the three different styles of mineralization that have been identified, namely sulphide, alloy and mixed. Different flowsheets were used for each style of mineralization, but all incorporated wet grinding as the initial stage. Encouraging results were obtained for sulphide mineralization, but the recovery and concentrate grade for both the alloy and mixed mineralization samples was poor due to excessive slimes and high viscosity. Phase 2 of the current program, which began in late 2008, was also managed by Mineral Solutions. During this phase, the process concept changed significantly, with the focus on pre-treatment to remove chrysotile fibres and brucite slimes. With the resulting reduction in slimes and lowered pulp viscosity, nickel recovery and concentrate grades improved markedly. This flowsheet was then developed into a standard process test to perform variability analysis on domain composite samples. The metallurgical performance for the three styles of mineralization (sulphide, alloy and mixed) using a common flowsheet was estimated by Mineral Solutions using the results from the Phase 2 metallurgical testwork program. These estimates, based on the standard simplified flowsheet presented in Figure 1.1, are summarized in Table

18 Figure 1.1 Simplified Metallurgical Flow Diagram Defibering Dynamic Classification Grinding and De-sliming Stage 1 Mag. Sep. Rougher Flash Float Flotation Conc. Fluff Float Tailings Slime Float Tailings Mag. Sep. Cleaner and Re grind Tailings Flotation Conc. Flotation Conc. Final Mag. Conc. Grinding and De-sliming Stage 2 Tailings Rougher and Scavenger Flotation Cleaner Flotation Flotation Conc. Tailings Tailings Table 1.4 Average Composite Sample Parameters and Estimated Nickel Recoveries Description Units Style of Mineralization Sulphide Alloy Mixed Average Minimum % Ni Sample Maximum % Ni Grades Average % Ni Average Ni Deportment Ni Recovery to Rougher Concentrate Cut-off % Ni Ni in Pentlandite % of Total Ni Ni in Heazlewoodite % of Total Ni Ni in Awaruite % of Total Ni Ni in Silicates % of Total Ni Minimum 1 % of Contained Ni Maximum 1 % of Contained Ni Average 1 % of Contained Ni Adjusted Recovery 2 % of Contained Ni Recovery to rougher concentrate, excludes contribution from fibre and slimes scavenger circuits. 2 Adjustment includes contribution from fibre and slimes scavenger circuits which increases overall recovery to rougher concentrate by a minimum of 6% (alloy and mixed mineralization) to a maximum of 8% (sulphide mineralization). These estimates are based on recovering approximately 50% of the contained Ni reporting to these scavenging circuits. 6

19 Regression analysis was used to develop nickel recovery equations. Each equation was applied to the entire modeled resource for domains 1 through 7 and an average recovery was calculated for each of the three types of mineralization (sulphide, alloy, and mixed). These calculated recoveries were compared to the average recoveries calculated by Mineral Solutions using results from repeated flowsheet testing of the three types of mineralization. The equations selected for use in the preliminary assessment were those that resulted in the lowest variance between calculated and idealized recoveries. These equations then had adjustment factors applied to account for the variance between calculated results and the average recoveries estimated by Mineral Solutions. The equations also include a provision for additional recovery of nickel from the fibre and slimes removal scavenger circuits. In February, 2010, Royal Nickel commissioned Minerals Associates Inc. (Minerals Associates) to design and construct a continuous mini pilot plant (MPP) on the Phase 1 preliminary assessment at a throughput of kg/h. This plant was completed in July, 2010 and testing of samples commenced in August, The MPP testwork is being performed to confirm the laboratory metallurgical performance (recovery, concentrate grades and reagent dosage) for the various mineralization types. Flowsheet optimization work will further investigate both sulphide and magnetic cleaning circuits (concentrate grade and recovery), and optimize reagent and energy costs. A trade-off study to evaluate alternative primary grinding options will also be completed. 1.5 PROCESS AND INFRASTRUCTURE DESIGN Proposed Flowsheet The key elements of the flowsheet are: Four-stage crushing, with Vertical Shaft Impact (VSI) crushers used for tertiary and quaternary crushing. Ore drying, to ensure moisture content in VSI feed does not exceed 2%. De-fibering of crushed mill feed using air classifiers. Grinding by ball mills, followed by de-sliming. Scavenger circuits to recover approximately 50% of nickel associated with the fibres and slimes (or approximately 5% of total contained nickel). Magnetic separation, with the magnetic concentrate further separated into ferro-nickel and nickel sulphide concentrates by flotation. Flotation of the magnetic separation tails to recover the bulk of nickel sulphide concentrate, which will be combined with concentrates from the scavenger and magnetic circuits. 7

20 Cleaning of the separate ferro-nickel and nickel sulphide concentrates. The fibre, slimes and rock tails will be combined into a single tailings product for disposal. The proposed flowsheet for the concentrator is shown in Figure 1.2. Figure 1.2 Processing Flowsheet Process Design Criteria The concentrator design was developed by BBA, based on the conceptual metallurgical flowsheet provided by Mineral Solutions and the following design criteria: The plant will treat 80,000 t/d, with feed grading an average of 0.27% Ni over the life of operation. The plant will operate continuously, 24 h/d, 365 d/y. Conventional gyratory and cone crushers used for primary and secondary crushing will have availability averaging 75%. VSI crushers used for tertiary and quaternary crushing will have average availability of 85%. 8

21 Grinding and flotation area equipment availability will be 95%. Dryers have been sized to treating 32% of the total feed, and reducing the moisture content from 4% to 2%. Equipment sizes are based on an assumed weight recovery to cleaner concentrate of approximately 2.0%. (The actual weight recovery is forecast at 0.5%). The average Bond ball work index is 21.3 kwh/t Infrastructure The following site infrastructure will be required to support mining and processing operations: Administration building. Concentrate warehouse. Fuel farm, with a facility to separately store diesel and fuel oil, mainly for the open pit and for the concentrator driers, respectively. Total capacity of the fuel farm will be approximately 10 days average consumption, or 2 ML of diesel and 3 ML of fuel oil. Mine workshop based on the maximum fleet size of t haul trucks, and associated support equipment. Electrical sub-station at 120 MW and associated reticulation system. Sewage treatment and a landfill site. Off-site infrastructure includes a 28 km electrical power line to connect with the grid near Amos. The water balance developed for the preliminary assessment indicates that, on average, at least 95% of process water requirements could be met through the re-use of process water (from the TMF) or from inflows to the open pit that would be captured in a sump. The remainder, averaging 337 m 3 /h (or 93 L/s), will be drawn from a pumping station to be constructed on the Villemontel River immediately south of the Dumont Property and approximately 3 km from the concentrator. 1.6 TAILINGS MANAGEMENT FACILITY Tailings from the fibre, slimes and flotation streams will be combined into a single product, which will have a relatively low density of 45% solids after thickening. During operation of the open pit, underflow from the tailings thickener will be pumped into the tailings management facility (TMF) using positive displacement pumps. After open-pit operations are complete, tailings generated from the treatment of lower-grade stockpiles will be pumped into the mined-out pit using lower-cost centrifugal pumps. 9

22 Studies to date indicate the waste products at Dumont will be benign, with no acid generated. Thus, the tailings will not require sub-aqueous disposal; nor will it be necessary to line the TMF with an impervious membrane. The TMF will be a conventional terrestrial facility that uses approximately 20% of the waste rock from the open pit for construction of an impoundment dyke. Clay in the underlying overburden, where present, will essentially act as a low permeability membrane. On closure, the TMF will contain 618 Mt of tailings, while the impoundment dykes will contain 197 Mt of waste rock. The maximum height will be 41 m. Golder has prepared the preliminary designs for the TMF and construction costs have been estimated using the cost model for the earthmoving fleet. Golder PasteTec has identified the pumping system that will be required, and its estimated cost. 1.7 ENVIRONMENTAL ASPECTS Environmental studies are well advanced, with the following work having been completed: Three phases of environmental baseline studies were completed during the period in order to establish the pre-development environmental condition of the property and identify potential areas of impact. A preliminary geochemistry study on a representative sample of mineralization, waste and potential tailings from the Dumont deposit to determine acid rock drainage and leaching characteristics. Future studies are planned and include: Further geochemical analysis to fully understand and predict the behaviour of tailings and waste rock, with a particular focus on the potential for metal leaching. Construction of an experimental in-situ tailings cell to quantify the potential for carbon sequestration under operating conditions by the serpentine component of the tailings. Hydrological studies to quantify the impact of proposed operations on the local water table and a nearby aquifer-bearing esker. Quantifying the impact of mining operations on existing wetlands and fish habitats, and identifying opportunities for mitigation. Characterization of the soils in the area that would be impacted by operations. Once the project scope is finalized during the pre-feasibility study, a Project Notice will be submitted to the Quebec Ministère du Dèveloppement durable, de l Environnement et des Parcs (MDDEP), or Ministry of Sustainable Development, Environment and Parks. MDDEP will accordingly advise on the scope and requirements of an environmental impact study (EIS). 10

23 The project scope is such that this study would be assessed jointly at the provincial and federal levels under the Canada-Quebec Cooperation Agreement. It is expected this assessment could take up to two years from the time of submission of the Project Notice before the granting of a Certificate of Authorization to commence construction. The assessment period would run in parallel with the feasibility study and detailed engineering and the overall impact on the project s critical path would thus be minimal. The scope of the project will result in permits being required from a number of federal departments, including: Department of Fisheries and Oceans Canada (DFO) as baseline studies have identified several species of fish inhabiting wetlands within the footprint of disturbance. Natural Resources Canada (NRCan) as storage and manufacture of explosives requires a licence from this agency. 1.8 CAPITAL COST ESTIMATE The capital cost estimate (±40%) summarized in Table 1.5 is expressed in real January, 2010 terms and assumes a long-term exchange rate of US$0.90/C$1.00. US$ millions Table 1.5 Capital Cost Summary Base Case (80,000 t/d) Upside Case (100,000 t/d) Initial Capital Mine $448 $457 Process Plant $709 $859 Tailings Dam $124 $138 Infrastructure $131 $152 Indirects $242 $274 Contingency $369 $424 Sub-Total $2,023 $2,304 Sustaining Capital Mine Fleet $331 $354 Mill $367 $361 Tailings Dam $159 $153 Closure $100 $100 Contingency $181 $182 Sub-Total $1,139 $1,150 Total Capital $3,162 $3,454 11

24 1.9 OPERATING COST ESTIMATE A summary of operating costs (±40%) is provided in Table 1.6. Table 1.6 Estimated Site Operating Cost Summary Area Units Base Case (80,000 t/d) Upside Case (100,000 t/d) Mining US$/t mined US$/t treated Processing US$/t treated G&A US$/t treated Sub-Total Site Costs US$/t treated US$/lb Ni TC/RCs US$/lb Ni By-Product Credit US$/lb Ni (0.16) (0.16) Net C1 Cash Costs + US$/lb Ni C1 costs include mining, processing, site administration and refining, net of by product credits 1.10 EVALUATION Base Case and Upside Case The evaluation included the following key macroeconomic assumptions: A long-term price for nickel of $7.50/lb. Sensitivity analysis considered a range of ±10%, or $6.75/lb $8.25/lb. A long-term price for cobalt of $12.00/lb. As the contribution of cobalt is only 1.7% of total NSR, it was not included in the sensitivity analysis. A long-term exchange rate of US$/C$ = Sensitivity analysis considered a range of ±10%, or US$/C$ = Long-term oil price of $80/bbl. The impact of the variation in oil price was included within the sensitivity analysis of changes in total operating costs. Table 1.7 summarizes key metrics for the base case and the upside case. Figure 1.3 illustrates the LOM free cash flow (FCF) for the base case, while Figure 1.4 illustrates the upside case. This preliminary assessment is preliminary in nature; and there is no certainty that the preliminary assessment will be realized. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. Inferred resources have not been included in this preliminary assessment. 12

25 Table 1.7 Summary Metrics Item Units Base Case (80,000 t/d) Upside Case (100,000 t/d) Mill Feed 1 million tonnes Grade % Ni Waste million tonnes 1,115 1,115 Stripping Ratio ore:waste Concentrator Recovery % of contained Ni 65.5% 65.5% Payables % of recovered Ni 92.5% 92.5% Recovered Ni million lb 3,551 3,551 Payable Ni million lb 3,286 3,286 Mill Throughput thousand t/d Project Life years Peak Ni million lb/y Annual Payable Ni million lb/y Site Operating Costs US$/tonne ore Net C1 Costs 2 US$/lb Ni Initial Capital US$ million 2,023 2,304 Sustaining Capital US$ million 1,139 1,150 Total Capital US$ million 3,162 3,454 Pre-Tax NPV 10% US$ million 1,073 1,433 Pre-Tax IRR % Post-Tax NPV 10% US$ million Post-Tax IRR % Note 1: Diluted in-pit measured and indicated resources Note 2: C1 costs include mining, processing, site administration and refining, net of by product credits. Figure 1.3 Base Case (80,000 t/d) LOM After-Tax Net Cash Flow 13

26 Figure 1.4 Upside Case (100,000 t/d) LOM After-Tax Net Cash Flow Sensitivity Analysis Figure 1.5 and Figure 1.6 illustrate the sensitivity of the base case and upside case, respectively, to the potential variation of ±10% in the following key input assumptions: Long-term nickel price (±10%: $6.75/lb $8.25/lb) Long-term exchange rate (±10%: US$/C$ = ) Average concentrator recovery (±10%: 59.0% 72.1%) Average head grade (±10%: 0.247% Ni 0.301% Ni) Total capital costs (±10%: $2,846 million $3,478 million) Site operating costs (±10%: $9.70/t $11.86/t) TC/RCs (±10%: $1.06/lb $1.30/lb) Both cases display similar behaviour, with results most sensitive to the nickel price (a 1% change has an impact of $40 million on the base case and $47 million on the upside case), head grade and recovery (for either, a 1% change has an impact of $38 million on the base case and $44 million on the upside case) exchange rate of C$/US$ (a 1% change has an impact of $24 million on the base case and $28 million on the upside case). Results are equally sensitive to variation in capital and operating costs (for either, a 1% change has an impact of $17 million on the base case and $19 million on the upside case). Variation in TC/RCs would have a lesser impact (a 1% change will have an impact of $7 million on the base case and $8 million on the upside case). The project NPV for the Base Case and Upside Case over a range of discount rates is given in Table

27 Figure 1.5 Sensitivity of NPV - Base Case (80,000 t/d) Figure 1.6 Sensitivity of NPV - Upside Case (100,000 t/d) Table 1.8 Sensitivity of NPV to a Selection of Discount Rates IRR (%) US$ million NPV at Discount Rate 8% 9% 10% Case Undiscounted Cash Flow Base Case Pre-Tax ,178 1,682 1,353 1,073 Post Tax , Upside Case Pre-Tax ,182 2,098 1,742 1,433 Post Tax ,226 1,

28 1.11 CONCLUSIONS The preliminary assessment has identified the following as generating optimal economic returns from development of the Dumont Property: An open pit designed to extract 2,011 Mt of total material, including 171 Mt overburden and 1,840 Mt of ore and waste rock. Mining should be performed using the largest class of equipment available, including 360-t haul trucks. For the hanging wall rocks, an acceptable factor of safety (1.5) will be achieved with a 47 slope angle, while the more competent footwall rocks will allow the same factor of safety to be achieved using a 52 slope. In overburden, slope angles will be reduced to 18 in fine-grained material and 27 in coarse-grained material. A mining rate that will be accelerated relative to the requirements of the mill, with open-pit mining completed in 20 years compared to the 31-year life-of-project. The accelerated mining rate will allow an elevated NSR cut-off value of C$14.00/t to be applied in the initial years of project life as well as providing a void for impounding approximately 30% of the tailings that will ultimately be produced. A steady-state milling rate of 80,000 t/d will maximize economic returns within the amount of grid power currently available to the project. Assuming utilization of the maximum electrical grid power available to the project on a planning basis, returns could be enhanced by increasing throughput to 100,000 t/d. A process flowsheet that includes 4-stage dry crushing followed by the removal of chrysotile fibres and brucite slimes before extraction of separate nickel sulphide and ferro-nickel concentrates. The crushing and screening operations will be equipped with dust extraction systems to ensure personnel are not exposed to any fibres. Chrysotile will be mixed with one or more of the other tailings products to ensure there are no fugitive airborne emissions from the TMF. This scope of design is estimated to require an initial capital investment of $2,023 million, and have average net C1 costs of $3.96/lb Ni. At a long-term Ni price of $7.50/lb, the posttax NPV 10% is estimated to be $488 million, while the IRR would be 14.1%. Undiscounted payback occurs near the end of the fifth year of production. Sensitivity studies suggest that the base case is sufficiently robust to withstand an adverse change of more than 10% in nickel price, head grade or process recovery, and at least 20% adverse change in capital, operating costs and C$/US$ exchange rate. The project is less sensitive to TC/RC terms. Consequently, Micon concludes that continued development of the Dumont Property, including work to complete a Preliminary Feasibility Study, is justified. 16

29 1.12 RECOMMENDATIONS Proposed Exploration Program Following upon the results of the Preliminary Assessment, Royal Nickel plans to continue to explore the Dumont Property in 2011 with the goal of refining the deposit model and collecting geotechnical and environmental data, and completing metallurgical testwork to support the completion of a Preliminary Feasibility Study. Based on the results of the Preliminary Feasibility Study, further exploration, geotechnical and metallurgical work will be initiated later in 2011 and continue into 2012 to support a Feasibility Study. The total expenditure for the 2011 exploration and drilling program, further metallurgical and other studies to complete the Preliminary Feasibility Study and initiate data collection for the Feasibility Study is estimated to be approximately C$29,000,000. The objectives of the work to be completed in 2011 are as follows: Outline additional resources that may occur inside the currently proposed pit shell. Increase drilling density in high value portions of the deposit that are currently in the indicated category. Continue to refine the geometallurgical model of the Dumont deposit based on drilling, geochemistry, mineralogy and metallurgical testing completed to date. Collect geotechnical data on rock mechanics and overburden properties in order to refine models of pit wall slopes and to select locations for surface infrastructure. Continue to characterize the environmental behaviour of tailings and waste rock. Engage stakeholders in consultation. Characterize local hydrological and hydrogeological regimes. Continue to operate a pilot plant based on the standard test procedure developed by Royal Nickel for the Dumont mineralization to demonstrate the commercial viability of the process. Complete a Preliminary Feasibility Study based on the geometallurgical model and metallurgical process development results to optimize parameters for a Feasibility Study. Initiate additional exploration and metallurgical work to support a Feasibility Study as indicated by the Preliminary Feasibility Study. 17

30 Royal Nickel has also proposed to spend a further C$30,700,000 in 2012 to complete the Feasibility Study on the Dumont Project. However, this expenditure will be based on the results and conclusions contained in the Preliminary Feasibility Study. Micon has reviewed Royal Nickel s proposal for further exploration and studies on its Dumont Property and considers that the budget for the proposed program is reasonable. Micon recommends that Royal Nickel implements the program as proposed, subject to either funding and other matters which may cause the proposed program to be altered in the normal course its business activities or alterations which may affect the program as a result of exploration activities themselves Further Recommendations Further development of the Dumont Property should address key risks and opportunities that have been identified through the following work program: Resource modelling. While the opportunity to extend the resource model along strike and at depth has been noted, the current measured and indicated resource (>1,100 Mt above a 0.20% Ni cut-off grade) is sufficient for the purposes of the Preliminary Feasibility Study. Resource modelling should focus on updating the database of mineralogical measurements and revising the forecast proportions of various Nibearing minerals accordingly. Geometallurgy. The geometallurgical model should continue to be refined based on drilling, geochemistry, mineralogy and metallurgical testwork. Geotechnical studies. Additional rock and soil mechanics holes should be drilled and samples taken that would allow the mechanical properties of overburden and rock to be determined. These studies could lead to revised designs for pit slopes and waste impoundments. A preliminary geotechnical study of the area proposed for the primary crusher and process plant should also be performed to gain a better understanding of civil work requirements and construction costs. Hydrological studies. Holes drilled for geotechnical samples should be used to gain a better understanding of the expected inflows to the pit. Measurements should also be taken to determine the flow rate in the Villemontel River, and the sustainable flow rate that would be available to the project. Metallurgical testing should focus on three areas: o Establishing the repeatability of results achieved to date, through testing of largerscale samples that capture the range of variability in mineralogy and associated metallurgical response that will be experienced during commercial operations, thereby improving confidence in the recovery forecasts. 18

31 o Optimizing project value by determining set points for Ni recovery, concentrate grades and costs that yield maximum NPV and IRR. o Investigating alternatives to 4-stage dry crushing that may be lower cost and/or require less power and thus be amenable to increased scale. The potential benefit of using natural gas in place of fuel oil to power the dryers should be investigated. Tailings produced from the metallurgical pilot plant should be tested to establish the optimal concept for impoundment. It is possible that the current concept (combining three tailings products into a single stream with a relatively low solids content) may be replaced with an alternative that entails creating separate higher- and lower-density streams. If soil conditions allow it, the ultimate height of the TMF may be increased in order to reduce the overall footprint of the facility. Environmental and socio-economic studies and public consultation should proceed with the objective of permitting the project in line with the feasibility timing and assume that there will be a joint provincial and federal review process. Anticipating the expected requirements of pending revisions to Quebec s Mining Act, closure costs in the cash flow model are split into (i) bond payments, incurred over the first five years of operation (amounting to 100% of the costs for which bonds are required), and (ii) decommissioning and monitoring costs, reflected as a lump sum at the end of the mine life. Estimated reclamation costs and bonding requirements should be reassessed in the next phase of development. Discussions with potential off-takers should be held, towards obtaining heads of agreement regarding off-take terms and conditions. Results from the various tests and studies listed above should be incorporated into a revised mine design and LOM schedule. Once the revised open pit design, production schedule, and modified flowsheet have been completed, they should be used to generate new capital and operating cost estimates to the level of accuracy normally associated with a Preliminary Feasibility Study, and an updated economic evaluation could then be prepared. 19

32 2.0 INTRODUCTION AND TERMS OF REFERENCE At the request of Mr. Tyler Mitchelson, President and CEO of Royal Nickel Corporation (Royal Nickel), Micon International Limited (Micon) has been retained to review and compile the results of independent work carried out for Royal Nickel comprising a Preliminary Assessment of its Dumont Property located in western Quebec, Canada. Participants in the study are listed in Table 2.1. The role of each participant is more fully described below. Table 2.1 Participants in Dumont Preliminary Assessment Activity Lead Organization Qualified Person Project Management A.St. Jean Royal Nickel Geology Exploration and Database A.St. Jean Royal Nickel A. San Martin Mineralogy S. Downing SGS W. Lewis Resource Model O. Tavchandjian Golder A. San Martin Resource Estimate W. Lewis Micon W. Lewis Geotechnical Slopes, Designs M. Garon Genivar Hydrology O. Fala Genivar D. Penswick Mining Concept Selection D. Penswick Independent LG Pit Design A.von Wielligh Prysm Resources Detailed Pit Design and Schedule R. Kear Independent D. Penswick Design Review M. Garon Genivar Processing Flowsheet Concepts and Design R. Salter Mineral Solutions Testwork J. Marois CTMP R. Gowans Engineering Design C. Hardie BBA C.Hardie Tailings Design, Concept R. Ouellet Golder M. Lemieux Pumping P. Primeau Golder Paste Tec P. Primeau Environmental Baseline Studies and Testwork D. Blanchet Genivar Water Balance C. Hardie BBA n/a Closure Plan D. Penswick Independent Schedule and Execution Plan D. Penswick Independent D. Penswick Capital Cost Estimate Mining D. Penswick Independent D. Penswick Processing, Infrastructure M. Fitzgibbon BBA C.Hardie Tailings M. Lemieux Golder M. Lemieux Operating Cost Estimate Mining D. Penswick Independent D. Penswick Processing C. Hardie BBA C.Hardie Tailings M. Lemieux Golder M. Lemieux G&A D. Penswick Independent Evaluation and Report D. Penswick Independent D. Penswick 20

33 The geological setting of the property, mineralization style and occurrences, and exploration history were described in earlier reports by Golder (2010), Lewis and San Martin (2010, 2009 and 2008), Lewis (2007), Caron (2004), Oswald (1988), Duke (1986), Honsberger (1971) and in various government and private publications. The relevant sections of those reports are reproduced herein. The Preliminary Assessment is based on the mineral resource estimate prepared by Micon and disclosed in its most recent technical report entitled NI Technical Report, Mineral Resource Estimate for the Dumont Property Launay And Trécesson Townships, Quebec, Canada dated August 30, Prior to that, Micon had prepared four other Technical Reports. In chronological order, these are: NI Technical Report on the Dumont Property, Launay and Trécesson Townships, Quebec, Canada dated August 15, 2007; NI Technical Report, Preliminary Mineral Resource Estimate for the Dumont Property, Launay and Trécesson Townships, Quebec, Canada dated April 30, 2008; NI Technical Report, Updated Mineral Resource Estimate for the Dumont Property, Launay and Trécesson Townships, Quebec, Canada dated January 23, 2009; and NI Technical Report, Mineral Resource Estimate for the Dumont Property, Launay and Trécesson Townships, Quebec, Canada dated April 5, Geotechnical studies were carried out for Royal Nickel by Genivar engineers, working under the supervision of independent mining engineer David Penswick, P.Eng. Open pit optimization was conducted by Prysm Resources (Pty) Ltd and open pit mine design and scheduling by Robin Kear. This work, carried out under the supervision of David Penswick, was also reviewed by Genivar. Mining capital and operating cost estimates were prepared by David Penswick. Quebec s Centre de Technologie Minérale et de Plasturgie (CTMP) carried out a metallurgical testwork program under the supervision of independent consulting metallurgist Robert Salter, Ph.D, P.Eng., who prepared the process flowsheet. The results of this work, and the development of the metallurgical parameters used in the evaluation of the property, have been reviewed by Micon s principal metallurgist, Richard Gowans, P.Eng. Process and infrastructural designs, water balance and associated capital and operating cost estimates (±40%) were prepared by BBA Engineering of Quebec. Tailings management facility designs have been prepared by Golder Associates Ltd (Golder) while Golder Paste Technology Ltd (Golder PasteTec) was responsible for estimates related to tailings pumping equipment. David Penswick estimated costs for construction of earthworks using the fleet operating cost model. 21

34 Mr Lewis has undertaken five visits to the Dumont Property, the most recent of which took place July 20-23, Previous visits were on May 8-9, 2007, January 29-31, 2008, October and 20-22, 2008, and October 16-18, Mr Penswick visited the property on August 10, The prime reports prepared by Royal Nickel s consultants as components of the preliminary assessment are listed as follows: Aker Solutions (various) Dumont Nickel Conceptual Study Update August, 2008 Genivar (O. Fala) Restricted Hydrogeological Study (for the) Dumont Nickel Project November, 2009 Genivar (various) Conceptual Review Dumont Property November 24, 2009 Genivar (L. Li) Preliminary Stability Analysis of Slopes January, 2010 Genivar (D. Blanchet) Preliminary Slope Evaluation of the Overburden Dumont Project April 26, 2010 Prysm Resources (A. Swart) Dumont Nickel Project Open-Pit Optimization Study June, 2010 D. Penswick Recovery Equations Used in Scoping Study June 2, 2010 Golder (G. Warren) GeoMetallurgical Modeling of the Dumont project June 28, 2010 BBA (various) Scoping Study for the Dumont Nickel Ore Project (draft) July 16, 2010 R.M. Kear Dumont Project Preliminary Mine Plans August, 2010 Micon (Lewis and San Martin) NI Technical Report Mineral Resource Estimate for the Dumont Property August 30, 2010 All currency amounts are stated in US dollars or Canadian dollars, as specified, with commodity prices typically expressed in US dollars. Quantities are generally stated using the Système International d Unités (SI) or metric units, the standard Canadian and international practice, including metric tonnes (t), kilograms (kg) or grams (g) for weight, kilometres (km) or metres (m) for distance and hectares (ha) for area. Wherever applicable, imperial units have been converted to SI units for reporting consistency. Table 2.2 provides a list of the abbreviations used throughout this report. 22

35 Table 2.2 List of Abbreviations Term Abbreviation Barrel(s) bbl Canadian dollar C$ Canadian National Railway CNR Cobalt Co Cubic metre(s) m 3 Cubic metres per hour m 3 /h Days per year d/y Degree(s) o Degrees Celsius o C Foot(feet) ft Front-end loader FEL Gram(s) g Gold Au Hectare(s) ha High pressure grinding roll HPGR Hour(s) h Hour(s) per day h/d Inch(es) in Index of rock strength IRS Internal rate of return IRR Iron Fe Kilogram(s) kg Kilograms per hour kg/h Kilometre(s) km Kilowatt(s) kw Kilowatthour(s) kwh Kilowatthours per tonne kwh/t Lerchs-Grossmann LG Life-of-mine LOM Litre(s) L Litres per second L/s Low intensity magnetic separator LIMS Magnesium/magnesium oxide Mg/MgO Megawatt(s) MW Metre(s) m Metres per second m/s Micron(s) μm Milligrams mg Millimetre(s) mm Millimetres per year mm/y Million M Million tonnes Mt Million tonnes per year Mt/y Million years old Ma Minute(s) min Net present value NPV Net present value at discount rate of 10% per year NPV 10% Net smelter return NSR Neutralization potential NP 23

36 Term Abbreviation Nickel Ni Palladium Pd Parts per billion ppb Parts per million ppm Platinum Pt Platinum group elements PGE Pound(s) lb Pounds per year lb/y Preliminary Feasibility Study PFS Provincial water quality objectives PWQO Quality Assurance/Quality Control QA/QC Rock quality designation RQD Second s Semi-autogenous grinding SAG Square metre(s) m 2 Square kilometre(s) km 2 Tailings management facility TMF Three-dimensional 3D Tonne(s) t Tonnes per cubic metre t/m 3 Tonnes per day t/d Tonnes per hour t/h Tonnes per year t/y Treatment charge/refining charge TC/RC Unconfined compressive strength UCS United States dollars US$ or $ Vertical-axis time domain electromagnetic VTEM Vertical shaft impact VSI Volt(s) V 24

37 3.0 RELIANCE ON OTHER EXPERTS Micon has reviewed and analyzed data provided by Royal Nickel, its consultants and previous operators of the property and, augmented by its direct field examination, has drawn its own conclusions therefrom. Micon has not carried out any independent exploration work, drilled any holes or carried out any extensive program of sampling and assaying on the property. However, during its field visit in 2007, Micon did collect five quarter-core samples from the zone of mineralization located on the Dumont Property. The results of this sampling program are contained in the August 15, 2007 and April 30, 2008 Technical Reports. While exercising all reasonable diligence in checking, confirming and testing it, Micon has relied upon Royal Nickel s presentation of its project data and that of previous operators of the Dumont Property, in formulating its opinion. The various agreements under which Royal Nickel holds title to the mineral claims for this project have not been reviewed by Micon, and Micon offers no legal opinion as to the validity of the mineral title claimed. A description of the property, and ownership thereof, is provided for general information purposes only. Comments on the state of environmental conditions, liability, and estimated costs of closure and remediation have been made where required by NI In this regard Micon has relied on the work of Genivar and other experts it understands to be appropriately qualified, and Micon offers no opinion on the state of the environment on the property. The statements are provided for information purposes only. The descriptions of geology, mineralization and exploration used in this report are taken from reports prepared by various companies or their contracted consultants. The conclusions of this report rely on data available in published and unpublished reports supplied by the various companies which have conducted exploration on the property, and information supplied by Royal Nickel. The information provided to Royal Nickel was supplied by reputable companies or government agencies and Micon has no reason to doubt its validity. The structural information used in the current mineral resource estimate was originally developed by an external consultant, Itasca Consulting Canada Inc. (Itasca Consulting), and it has significantly changed the manner in which the deposit has been modelled. Micon and Royal Nickel simplified the original structural model into seven separate domains for use in the resource model. Some of the figures and tables for this report were reproduced or derived from historical reports written on the property by various individuals and/or supplied to Micon by Royal Nickel. Most of the photographs were taken by the authors of this report during their respective site visits. In the cases where photographs, figures or tables were supplied by other individuals or Royal Nickel they are referenced below the inserted item. 25

38 4.0 PROPERTY DESCRIPTION AND LOCATION The Dumont Property is located in the western portion of the Canadian province of Quebec. Specifically, the property is located primarily in Launay and partly in Trécesson Townships in the Abitibi Region. The location of the project is shown in Figure 4.1. The longitude and latitude for the Dumont Property are approximately N, W. The UTM coordinates are approximately 5,391,500N, 688,400E with UTM zone 17 using the NAD83 Datum. The property is located approximately 25 km west of the city of Amos, approximately 60 km northeast of the industrial and mining city of Rouyn- Noranda and 70 km northwest of the city of Val D Or. Royal Nickel has advised that the Dumont Property consists of 138 contiguous mineral claims totalling 5, ha. The mineral claims confer the subsurface rights only. Approximately 25% of the surface rights for the property are held privately and the balance is public land. See Figure 4.2 for a claim map of the Dumont Property. In cases where private land has been crossed during the exploration program, agreements have been reached with the surface owners for access to the drill sites. There are no known formal native land claims covering the Dumont Property. Figure 4.3 shows the extent of the lands that are classed as an agricultural zone within the meaning of the Loi sur la protection du territoire et des activités agricoles respecting the preservation of agricultural land and agricultural activities. Mining activity on these lands would require rezoning and exclusion of these lands from the agricultural zone by the Quebec Agricultural Land Commission (CPTAQ). This exclusion must be requested by the local municipality. The application for exclusion must demonstrate that there are no suitable non-agricultural lands available for the stated purpose in the municipality. Royal Nickel does not expect that exclusion for the purpose of developing the Dumont Project would be unreasonably withheld. Drilling on the public land (Crown land) is being conducted under a forestry operational permit with the Quebec Ministry of Natural Resources whereby stumpage fees are paid for timber cut in order to access drill sites. 26

39 Figure 4.1 Dumont Property Location Map 27 Figure supplied by Royal Nickel Corporation.

40 Figure 4.2 Dumont Property Mineral Claims 28 Figure supplied by Royal Nickel Corporation.

41 Figure 4.3 Location Map Showing the Arctic-St. Lawrence Drainage Divide and Agricultural Zone Lands as they Relate to the Dumont Property 29 NOTE: Black circles represent Royal Nickel drill holes. Figure supplied by Royal Nickel Corporation.

42 The mineral claim boundaries coincide with the established township Lot and Range boundaries. The mineralized zones which are the subject of this report are located mainly in Ranges V, VI and VII on Lots 46 to 62 of Launay Township and in Range V on Lots 1 to 3 of Trécesson Township. The property is covered by a layer of glacial overburden and swamp land and mineralization subcrops approximately 30 m below the present surface. The mineral claims are held 100% by Royal Nickel. The mineral claims can be subdivided into four blocks, three of which were obtained from separate parties and as a consequence each block is the subject of a separate underlying agreement. The details of the underlying mineral claim agreements are described below. 4.1 DUMONT PROPERTY MINERAL CLAIM AGREEMENTS Griffis International Ltd. Mineral Claims The Griffis International Ltd. (Griffis) mineral claim block is comprised of 24 mineral claims totalling 1, ha. This block of claims was originally held by Griffis, but a 100% interest in the claims was sold and transferred to Royal Nickel under an agreement dated January 15, The agreement with Griffis is not subject to any further future consideration, work commitment requirement or NSR Marbaw International Nickel Corporation Mineral Claims The Marbaw International Nickel Corporation (Marbaw) mineral claim block is comprised of 58 contiguous mineral claims totalling 2, ha. This block of claims was originally held by Marbaw, but a 100% interest in the claims was sold and transferred to Royal Nickel, for future consideration, under an agreement dated March 8, Future consideration consisted of the following: 1) Issuance of 7 million shares in Royal Nickel to Marbaw upon the property being placed into commercial production or upon transfer of the property to a third party. 2) Payment of C$1,250,000 to Marbaw on March 8, This amount has been paid by Royal Nickel. Royal Nickel has also committed to incurring a minimum expenditure of C$8,000,000 on the property prior to ceasing operations. This commitment was met in The Marbaw mineral claims are also subject to a 3% NSR royalty payable to Marbaw and Royal Nickel has the right to buy back half of the 3% NSR for C$10,000,000 at any time. 30

43 4.1.3 Sheridan-Ferderber Mineral Claims The Sheridan-Ferderber mineral claim block comprises 6 contiguous claims totalling ha. The claims were originally held 50% by Terrence Coyle and 50% by Michel Roby, but they were optioned to Patrick Sheridan and Peter Ferderber under an agreement dated October 26, The option agreement was subsequently assigned to Royal Nickel through an agreement dated May 4, Royal Nickel s option to acquire 100% interest in this block of mineral claims was exercised by the completion of C$75,000 in work on the claims before October 26, 2008 and by paying C$10,000 to Coyle-Roby by October 26, 2007 and C$30,000 to Coyle-Roby by October 26, The claims were transferred 100% to Royal Nickel on August 25, These claims are subject to a 2% NSR royalty payable to Terrence Coyle (1%) and Michel Roby (1%). Royal Nickel has the right to buy back half of this 2% NSR for C$1,000,000 at any time. An advance royalty of C$5,000 per year is also payable to Coyle-Roby beginning in Royal Nickel Claims In March, 2009, Royal Nickel staked a contiguous buffer block of 50 claims totalling 2, ha to the southwest of the Dumont Property. There is no known mineral resource on these claims. Royal Nickel holds a 100% interest in the claims, which are not subject to any royalty or other underlying agreement. Micon is unaware of any outstanding environmental liabilities attached to the Dumont Property and is unable to comment on any remediation which may have been undertaken by previous companies. 31

44 5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY The Dumont Property is readily accessible from the city of Amos, via Quebec provincial Highway 111 which runs along the southern boundary of the property in an east-west direction. The property is accessed from Highway 111 via a network of forestry and township boundary roads which branch off the highway. Royal Nickel has constructed over 3.5 km of gravelled all-weather roads to provide year-round access from Highway 111 along the axis of the Dumont sill between lines 5200E and 10400E. These roads and a gravel pit on the property have been permitted by the Quebec Ministry of Natural Resources. The Canadian National Railway (CNR) crosses the property just north of Highway 111. The major centre for the area is the city of Amos with a population of 12,584 according to the last Canadian census (2006). Amos has a municipal airport, but scheduled flights from major centres land at either the Rouyn-Noranda or Val d Or airports. Rouyn-Noranda and Val d Or are the main regional centres with populations of approximately 39,925 and 31,130, respectively. See Figure 5.1 for a map indicating the access and various other geographical features in the area of the Dumont Property. The Dumont Property is located approximately 25 km west of the city of Amos, approximately 60 km northeast of the industrial and mining city of Rouyn-Noranda and 70 km northwest of the city of Val d Or. Access to Amos from Rouyn-Noranda is via provincial Highways 117 and 109 and access from Val d Or is via provincial Highways 117 and 111. Both Rouyn-Noranda and Val d Or are well established mining communities and skilled labour for mining is readily available in these centres. The closest accommodations are located in Amos, which has several motels, hotels and restaurants. The Dumont Property is situated in the Abitibi region of northwestern Quebec. The property exhibits low to moderate relief up to a maximum of 30 m and lies between 310 m and 350 m above sea level. The climate at the Dumont Property is continental with an average annual temperature of 1.2 C, a monthly high of 23 C in July and a monthly low of -23 C in January. Total average annual precipitation is 918 mm. While field exploration work can be conducted year-round, drill access in low-lying boggy areas is best during the frozen winter months. Also, periodic heavy rainfall or snowfall can hamper exploration at times during the summer or winter months. Water for diamond drilling programs is obtained from several creeks which run through the property and is generally pumped to the drill sites. However, fresh water can also be supplied by the nearby Villemontel River. Power to the property may be supplied by the Hydro Quebec Amos substation. 32

45 Figure 5.1 Map Indicating Access and Geographical Features in the Area of the Dumont Property 33 Figure supplied by Royal Nickel Corporation.

46 Wildlife on the property consists of moose, black bear, beaver, rabbit and deer. Some logging has been conducted on the property with the wood being used primarily for pulp. Figure 5.2 is a view of the current access road on the Dumont Property built by Royal Nickel. Figure 5.2 Access Road for the Dumont Property 34

47 6.0 HISTORY The general exploration history for the Dumont Property is described in the technical report dated August 30, 2010, NI Technical Report, Mineral Resource Estimate for the Dumont Property, Launay and Trécesson Townships, Quebec, Canada (Lewis and San Martin, August, 2010). The reader is referred to this report for details of exploration between 1935 and While the presence of ultramafic and mafic rocks has been known on the Dumont Property since 1935, the presence of nickel within the rock sequence was only discovered in However, it was not until the 1970s that the existence and potential of the large low-grade nickel mineralization was first recognized. 6.1 ROYAL NICKEL EXPLORATION Royal Nickel raised money to conduct an initial exploration drilling program in 2007 which consisted of five twin holes to confirm the historic drilling results, followed by a further 38 infill and step-out drill holes totalling 15,606 m. The results of the initial drilling program were discussed in a Technical Report entitled NI Technical Report on the Dumont Property, Launay and Trécesson Townships, Quebec, Canada dated August 15, Following this initial phase, Royal Nickel continued to conduct a comprehensive exploration program on the Dumont sill which consisted of two parts implemented progressively throughout The first part of the exploration program was comprised of: A 200 m spaced sectional drilling program designed to increase confidence in the historical resource defined around the central zone mineralization between sections 7400E and 8700E, down to a vertical depth of approximately 350 m from surface. A helicopter-borne Geotech vertical-axis time domain electromagnetic (VTEM) and magnetic survey over the entire property. The second part of the program began in late 2007 and continued through This portion was designed to: Provide further confidence on the central zone resource on 100 m spaced sections. Complete drilling on 200 m spaced sections in other zones of promising mineralization. Complete an initial evaluation of the relatively unexplored portions of the Dumont sill. Test targets generated by the VTEM survey. 35

48 Forty-three diamond drill holes totalling 17,288 m were drilled in Following the successful drilling program, Micon conducted a preliminary resource estimate on the Dumont Property, the details of which are contained in the Technical Report entitled NI Technical Report, Preliminary Resource Estimate for the Dumont Property Launay and Trécesson Townships, Quebec, Canada dated April 30, The exploration program in 2008 focused on completing diamond drilling on regularly spaced sections and 40,803 m were drilled in 96 holes up to October 31, Results were received from 89 holes totalling 37,638 m and these were included in the updated mineral resource estimation discussed in the January, 2009 Technical Report. In addition, two short drill holes totalling 231 m were completed to gather metallurgical samples. In addition to the drilling, a surface mapping program was carried out over the Dumont Property, primarily to provide a structural geology framework for the modelling of the Dumont deposit. Field mapping was completed in July and August, Given the poor exposure over the Dumont sill, the mapping program focused on outcrops in the country rocks outside the sill, in order to gain an understanding on the local structural geology. A secondary purpose for the program was to identify areas of outcrop for potential infrastructure (mill) development. Information collected during this program was interpreted in association with airborne magnetics and LIDAR Bare Earth model (LIDAR) topography data and was used to update the historic maps. The exploration work completed in 2009 focused on diamond drilling on regularly-spaced sections in order to increase confidence in the resource reported in the January 23, 2009 Technical Report. In addition to this resource definition drilling, several programs intended to characterize the deposit and its environment were undertaken in order to support development studies. This included several holes completed to define structures to assist in geological interpretation and deposit modelling, to assess geotechnical properties of the rock, and to provide samples for metallurgical testing. Additionally, ongoing environmental baseline studies and mineralized material and waste rock characterization studies continued in In 2009, drilling was completed to yield a nominal 100 m by 100 m drill spacing over the Dumont deposit from section 5700E to 9000E. In addition, 50 m by 50 m spaced drilling was completed over two blocks centred on section 8250E and on section 6850E. The purpose of this drilling was to assess the variability of the deposit on the 50 m scale as compared to the 100 m scale. In total, 18,061 m of core drilling in 52 holes were completed in 2009 for the purpose of resource definition. For the purpose of defining major geological structures (faults) in the central portion of the deposit, 1,359 m were drilled in four oriented core holes. These holes were drilled parallel to the strike of the deposit and at high angles to the major structures that cross-cut the deposit. Data from these structural holes were combined with the global drill hole database and 36

49 surface mapping by John Fedorowich, Ph.D., P.Geo., of Itasca Consulting, to produce a first order structural model for the deposit that was used to delimit structural domains and help constrain the resource block model. In order to define rock mass characteristics and evaluate open pit slope angles on an indicative basis, data collection for a preliminary geotechnical study was carried out in Work associated with this study included the measurement and analysis of 1,503 m of core from drilling three oriented core holes near section 6800E, and a limited hydrogeological study between sections 6500E and 7500E. Drilling was also carried out to collect samples for bench-scale metallurgical variability testing and crushing testwork. A total of 2,964 m of drilling in eight holes was completed for metallurgical testwork, and three holes totalling 406 m were completed for crushing testwork. Additionally, a series of seven pilot drill holes totalling 1,757 m were completed to characterize the near-surface mineralization in order to select representative mineralization domains for sampling by large diameter drilling for mini pilot plant testing in The principal conclusions from the structural analysis completed by Itasca Consulting as follows: Macro-scale lineaments between 0.2 and 2.0 km in strike length cross-cut and offset the deposit at high angles, as represented within the first vertical derivative of the total magnetic field information. Ground truthing of lineaments in five surface mapping areas spanning the deposit and its environs found that meso-scale fabrics were similar to the overall pattern of the lineaments. Detailed structural mapping at 49 outcrops in these areas helped to define the penetrative regional foliation and faulting pattern. The exposed faults are <1 m true width, dominantly sub parallel to regional foliation, or at northeast - southwest trends, which cut the regional foliation. Faults observed in drill core are brittle-ductile structures that typically have slickenside lineated surfaces, cm-scale seams of gouge, an envelope of broken ground, shear and extensional serpentine veins, and subsidiary shear fracture sets. Serpentine veins appear to have developed together with the faulting. Three sets of faults are consistently observed in oriented core scanline traverses using various combinations of oriented core drill holes. The dominant set is belt parallel, which is also the orientation of regional foliation. The two other sets are at high angles to this, at northeast-southwest and east northeast south southeast orientations. Fault spacing is found to be somewhat variable. An average spacing of faults has been calculated at about 40 m, 70 m and 30 m for sets 1, 2 and 3, respectively. A fault model was developed for 17 pronounced structures within the project area. Plan and section analysis was carried out to obtain a best fit of lineaments and fault intercepts in drill core (oriented and un-orientated). Where possible, shear sense and 37

50 net slip orientations were determined by combining the measured offset of lithological contacts with fault plane lineation information obtained in oriented drill core. This allows for fault plane solutions, which help indicate drill targets for infill drilling where mineralization is offset. Small-scale features observed in oriented drill core intercepts of faults and associated veins can be understood within the larger scale fault pattern. The exploration work conducted or assays received up to the effective date of August 16, 2010 are described in Sections 10 and 11 of this report. 6.2 HISTORICAL MINING AND PRODUCTION No historical mining or production has been conducted on the Dumont Property. However, the Val D Or - Rouyn-Noranda region surrounding the Dumont Property has been a prolific mining area for the past 100 years. 6.3 DUMONT PROPERTY RESOURCE AND RESERVE ESTIMATES Historical mineral resource estimates, and subsequent Royal Nickel resource estimates are discussed in Lewis and San Martin, August, 2010, to which the reader is referred for detail. 38

51 7.0 GEOLOGICAL SETTING The following description of the geology of the Dumont Property, presented in Lewis and San Martin, August, 2010, to which the reader is referred for more detail. It is based on the description and the genetic model for the formation of the Dumont sill provided by Duke (1986). 7.1 REGIONAL GEOLOGICAL SETTING A thick supracrustal succession of Archean volcanic and sedimentary rocks underlies about 65% of the Abitibi belt and there is evidence to suggest that these supracrustal rocks lie unconformably upon a basement complex of sialic composition. The volcanic rocks are mainly of mafic composition although ultramafic, intermediate and felsic types are also present. The abundance of pillowed and non-vesicular lavas, together with the flyschoid character of much of the sedimentary component, demonstrates the prevalence of deep submarine conditions. However, the occurrence of some fluvial sedimentary rocks and airfall tuffs attest to occasional local non-marine conditions. Numerous small to medium sized synvolcanic intrusions reflect the range of compositions of the lavas themselves. The supracrustal rocks were deformed and intruded by granitic stocks and batholiths during the Kenoran event about 2,680 to 2,700 million years (Ma) ago. Folding along generally east-trending axes has commonly produced isoclinal structures. Regional metamorphism is predominantly greenschist and prehnite-pumpellyite facies except in the contact aureoles of the Kenoran granites where amphibolite grade is usually attained. The amphibolite facies metamorphism also occurs in the sedimentary rocks of the Pontiac Group. The Dumont sill is hosted by lavas and volcaniclastic rocks assigned to the Amos Group. The lavas may be traced eastwards through Amos and are part of the Barraute volcanic complex. Three cycles of mafic to felsic volcanism are recognized and the Dumont sill is one of at least five ultramafic-mafic complexes in the Amos area which occur at approximately the same stratigraphic level within the mafic lavas of the middle cycle. With the Amos (Landrienne) sill to the east of the Dumont sill, the host rocks of the sill are for the most part iron-rich tholeiitic basaltic lavas although some intermediate rocks were intersected in drilling of the footwall of the body at its eastern end. Although the volcanic rocks have been folded and now dip steeply, a penetrative deformational fabric is only locally developed. In the vicinity of the Dumont sill, pillows in the lavas are not strongly deformed and primary textures such as swallow-tail plagioclase microlites are preserved. However, the chemical compositions of many of the rocks are highly altered. Three main directions of faulting are recognized in the Amos area with the earliest being the east-trending set of bedding plane faults which are believed to have developed during the major period of folding. The second set of faults occurred during the intrusion of the granitic rocks which was accompanied by the development of steeply dipping faults that strike north to northwest. However, the most prominent faults strike northeast and probably postdate the granitic plutonism with the Dumont sill cut by a number of these northeast-trending faults. 39

52 7.2 CONTACT RELATIONS, AREAL EXTENT AND AGE OF THE DUMONT INTRUSION The property is covered by a layer of glacial overburden and swamp land and the mineralization subcrops approximately 30 m below the surface. Therefore, the contacts between the Dumont sill and its host rocks have not been observed in outcrop but, in overall attitude, the body appears to be conformable to the layering of the volcanic rocks and there is little doubt that it is a sill. Pillowed basalts exposed at the eastern end of the sill clearly indicate a northeast facing direction. The sill comprises a lower ultramafic zone and an upper mafic zone. Although less than about 2% of the bedrock surface of the intrusion is exposed in outcrop, the boundaries of the ultramafic zone can be drawn with some confidence on the basis of a magnetometer survey (Figure 7.1) and diamond drilling (Figure 7.2). Figure 7.1 Magnetometer Survey of the Dumont Property Figure supplied by Royal Nickel Corporation. The western end of the body has not been precisely outlined; however, the ultramafic zone is a lenticular mass at least 6,600 m in length with an average true thickness of 450 m. The true dip of the zone, 60 to 70 northeast, was determined in three cross-sections where either contact was intersected in two or more drill holes. The extent of the mafic zone is much less well defined and its top has not been observed in either outcrop or drill core. A drill hole near the middle of the sill intersected a true thickness of 175 m for the mafic zone, and outcrop constraints would allow a maximum of 400 m. It is believed that 250 m is a reasonable estimate of the average true thickness of the mafic zone and, therefore, the average true thickness of the Dumont sill as a whole could be 700 m. A number of faults at a 40

53 high angle to the long axis of the sill are suggested by offsets in the magnetic contours and also by the internal stratigraphy of the ultramafic zone. However, the inferential nature of these faults is emphasized and there are undoubtedly other faults which have not been shown. A dyke-like apophysis of peridotite, 30 m wide, occurs between the mafic zone and the overlying lavas at the eastern end of the intrusion. The contact between the gabbro and peridotite in this location is marked by a mylonite zone about a centimetre in width, and it is believed that the apophysis is a fault slice from the top of the ultramafic zone. This interpretation is supported by the composition and petrography of the peridotite. Two poorly exposed sills of gabbro occur about 1 km to the northeast of the Dumont body and may be consanguineous with the main sill. No feeder to the Dumont sill has been observed. See Figure 7.2 for a geological map of the Dumont sill based on Royal Nickel s work on the Dumont Property. Figure 7.2 Geological Map of the Dumont Sill Figure supplied by Royal Nickel Corporation. The age of the Dumont sill is not explicitly known. No radiometric age determination has yet been undertaken and there is little in the way of direct geological evidence. The conformable nature of the body together with the character of its differentiation, documented below, leave little doubt that it was emplaced as a virtually horizontal sill that was folded and faulted during the Kenoran event. It is reasonable to conclude that the Dumont sill is of late Archean age but is only slightly younger than the enclosing lavas; that is, about 2,700 Ma. 41

54 The sill is poorly exposed over a strike length of 6.5 km and comprises a lower ultramafic zone which averages 450 m in true thickness and an upper mafic zone about 250 m thick. The ultramafic zone is subdivided into the lower peridotite, dunite and upper peridotite subzones. The lower and upper peridotite subzones are olivine-chromite cumulates with variable amounts of intercumulus clinopyroxene. The dunite subzone is an extreme olivine adcumulate containing very small amounts of intercumulus chromite and clinopyroxene. Cumulus sulphide occurs in certain parts of the dunite subzone and also locally in the lower peridotite. The mafic zone comprises three subzones which are, from the base upwards, the clinopyroxenite, the gabbro and the quartz gabbro. The clinopyroxenite subzone is an extreme clinopyroxene adcumulate at its base but grades into clinopyroxene + plagioclase cumulate rocks in the overlying gabbro subzone. The quartz gabbro subzone includes both plagioclase + clinopyroxene cumulates and noncumulate gabbros that contain modal and normative quartz. Olivine and chromite are restricted to the ultramafic zone whereas plagioclase occurs only in the mafic zone. The magnesium/magnesium-iron ratios of the ferromagnesian cumulus phases increase gradually from the base of the sill upwards across the lower peridotite, undergo an abrupt increase at or just above the base of the dunite, remain essentially constant through the upper part of the dunite and into the base of the upper peridotite and then follow a normal iron enrichment trend upward through the overlying part of the intrusion. The parent magma of the intrusion is inferred to have been a peridotitic komatiite liquid containing about 27.5% MgO and carrying about 12.5% olivine (93% forsterite) in suspension. The lower peridotite subzone represents the accumulation of these phenocrysts and variable proportions of trapped intercumulus liquid. The cumulus chromite in the lower peridotite may also have been carried into the magma chamber in suspension or may have crystallized during settling of the olivine. The increase of forsterite content of olivine upwards across the lower peridotite reflects the amount of reaction of the cumulus olivine with intercumulus liquid before and during expulsion of the liquid from the crystal mush by filter pressing. The dunite, upper peridotite, clinopyroxenite and gabbro subzones accumulated by crystallization at or near the upward-migrating base of the convecting magma chamber. The base of the mafic zone represents a reaction relationship or distribution point whereby olivine and chromite are replaced as liquidus phases by clinopyroxene. Coarse grained, non-cumulate gabbros and pegmatoidal gabbro schlieren within the quartz gabbro subzone demonstrate that fractional crystallization of peridotitic komatiite magma can produce liquids containing as little as 5.5% MgO. Moreover, the presence of thin sills of similar evolved composition overlying the Dumont sill indicates that these liquids were mobile and mass balance considerations suggest that an amount of differentiated liquid equivalent to about 10% of the initial mass of magma was tapped off the main chamber. Some of the noncumulate rocks are very iron-rich (about 15% FeO), and differ in this respect from typical extrusive basaltic komatiites. Magmatic sulphides are restricted to the lower peridotite and dunite subzones, although in the former they represent mainly a post cumulus phase. In the dunite subzone olivine and sulphide are present in approximately their cotectic proportions and molten sulphide was apparently a cumulus phase. Three olivine-sulphide cumulate layers occur within the dunite 42

55 subzone but do not extend over the entire strike length of the sill. The middle layer has the highest average nickel grade (0.50%) and is the most laterally extensive, persisting over a strike of 2,400 m with an average true thickness of 24 m. A higher grade zone within the middle layer averages 0.71% nickel over a strike length of 730 m and has a true thickness of 14 m (Duke, 1986). The ultramafic rocks are pervasively serpentinized and serpentinization is overprinted by talc-carbonate alteration in places along the basal contact of the sill. The predominant secondary assemblage is lizardite + magnetite + brucite + chlorite + diopside ± chrysotile ± pentlandite ± awaruite ± heazlewoodite. Antigorite is developed locally, particularly in the uppermost ultramafic zone. Native copper occurs in some samples but not in those containing awaruite. Millerite occurs in steatitized rocks. The mafic zone rocks are ubiquitously altered to the assemblage actinolite + epidote + chlorite ± quartz. Primary textures are pseudomorphously preserved throughout most of the intrusion. The textures and assemblages of the secondary minerals are indicative of non-equilibrium, retrograde, low temperature alteration which may well have occurred as a result of an influx of water during the initial cooling of the intrusion. The sill was faulted and tilted into a steeply inclined attitude during the Kenoran event but no penetrative deformational fabric is evident, and the effects of regional metamorphism are minimal. Figure 7.3 illustrates the positions of the three nickel-enriched layers in the dunite subzone. The position of the basal contact of the sill is projected from off-section. Figure 7.3 Cross-Sectional View on Line 8100E Showing the Position of the Three Nickel-Enriched Layers Figure supplied by Royal Nickel Corporation. 43

56 8.0 DEPOSIT TYPES The following description of deposit type has been extracted from Lewis and San Martin, August, Magmatic nickel-copper-platinum group element (PGE) deposits occur as sulphide concentrations associated with a variety of mafic and ultramafic magmatic rocks. The magmas originate in the upper mantle, and an immiscible sulphide phase occasionally separates from the magma as a result of the processes occurring during emplacement into the crust. The sulphide phase generally partitions and concentrates nickel, copper and PGE elements from the surrounding magma. The heavy sulphide droplets once concentrated and separated from the magma tend to sink towards the base of the magma, and form concentrated pockets or layers of sulphides which crystallize upon cooling to form mineral deposits. The Dumont mineral deposit comprises olivine + sulphide cumulates which make up differentiated layers of the Dumont sill, an Archean komatiitic intrusion contained within the Archean Abitibi Greenstone Belt of northwestern Quebec. As such, it is usually classified with its most analogous counterpart, the Mt. Keith mineral deposit located in the Agnew- Wiluna Greenstone Belt within the Archean Yilgarn craton of West Australia. Greenstone belts are typical terranes found in many Archean cratons, and may represent intracratonic rift zones. The greenstone belts are generally composed of strongly folded, basaltic/andesitic volcanics and related sills, siliciclastic sediments, and granitoid intrusions which have been metamorphosed to greenschist and amphibolite facies, and typically adjoin tonalitic gneiss terranes. Komatiitic rocks form an integral part of some of these greenstone belts. Both the Dumont and Mt. Keith deposits have undergone pervasive serpentinization and local talc-carbonate alteration as a result of metamorphism to mid-upper greenschist facies. This alteration history has resulted in liberation of much of the nickel from nickel silicates (olivine) and consequent upgrading of the primary magmatic nickel-sulphide and nickel-alloy minerals through partitioning of nickel. 44

57 9.0 MINERALIZATION The following description of the mineralization for the Dumont deposit has been extracted from Lewis and San Martin, August, Two types of mineralization have been identified historically within the Dumont sill, the primary large low-grade to medium-grade disseminated nickel deposit (Duke, 1986) and the contact type nickel-copper-platinum group elements (PGE) occurrence discovered in 1987 (Oswald, 1987). Drilling by Royal Nickel has also identified discontinuous PGE mineralization associated with disseminated sulphides at lithological contacts in the layered intrusion and within the dunite. 9.1 DISSEMINATED NICKEL MINERALIZATION Nickel bearing sulphides and a nickel-iron alloy are enriched within three distinct layers of the dunite subzone, the upper layer, the middle layer, and the lower layer, and are broadly disseminated throughout the dunite and lower peridotite subzones. In thinner parts of the dunite subzone, fewer than three enriched layers may be present. Nickel mineralization continues at lower grades between the enriched layers Nickel Mineralogy Disseminated nickel mineralization is characterized by disseminated blebs of pentlandite ((Ni,Fe) 9 S 8 ), heazlewoodite (Ni 3 S 2 ), and the ferro-nickel alloy, awaruite (Ni 2.5 Fe), occurring in various proportions throughout the sill. Millerite (NiS) is also present in lesser amounts near host rock contact zones. These minerals can occur together as coarse agglomerates, often associated with magnetite, up to 10,000 µm (10 mm), or as individual disseminated grains ranging from 2 to 1,000 µm (0.002 to 1 mm). Figure 9.1 shows nickel mineralization in core from the Dumont Property. Figure 9.1 Photo of the Dumont Mineralization in Core (Field of View = 5 cm) Figure supplied by Royal Nickel Corporation. 45

58 The observed mineralogy of the Dumont deposit is likely a result of the serpentinization of a dunite which hosted a primary disseminated (intercumulus) magmatic sulphide assemblage. The serpentinization process whereby olivine reacts with water to produce serpentine, magnetite and brucite creates a strongly reducing environment where the nickel released from the decomposition of olivine is partitioned into low-sulphur sulphides and newly formed awaruite. Royal Nickel s mineralogical sampling program provides an analytical measure of the whole-rock mineralogy which is the basis for the classification of the disseminated nickel mineralization into three distinct mineralization types: sulphide, alloy, and a transition between these two members referred to as mixed. The mineralization types are defined by the ratio of nickel-bearing sulphides to awaruite ((pentlandite+heazlewoodite)/awaruite). Samples with a ratio of nickel-bearing sulphides to awaruite greater than 5.5 are classified as sulphide mineralization type. Samples with a ratio of nickel-bearing sulphides to awaruite less than 0.85 are classified as alloy mineralization type. The transition between these two mineralization types, samples with a ratio of nickel-bearing sulphides to awaruite of 0.85 to 5.5 are classified as mixed mineralization type. The sulphide mineralization type occurs in higher grade bands (over 0.35% nickel) parallel to the dip of, and principally in the centre of, the sill and is dominated by pentlandite and/or heazlewoodite with lesser awaruite (Table 9.1). Table 9.1 Nickel Bearing Mineral Abundance by Mineralization Type Mineralization Type % Nickel* % Pentlandite** % Heazlewoodite** % Awaruite** Mean Min Max Mean Min Max Mean Min Max Mean Min Max Sulphide Alloy Mixed * % Ni in the mineralogical samples (189 mineralogical mapping samples to April 22, 2010). ** Mineral abundance by SGS Lakefield Research (Explomin TM Particle Scan, 189 samples to April 22, 2010). Table provided by Royal Nickel Corporation. Sulphides occur as medium to coarse grained blebs associated with magnetite±brucite±chromite, in intercumulus spaces as a primary magmatic texture. Figure 9.2 shows an example of the mineralogical textures in the sulphide mineralization type. The alloy mineralization type is characterized by the presence of awaruite as fine, discrete grains and/or associated with intercumulus magnetite blebs. Alloy mineralization generally occurs parallel to the Dumont sill along the hangingwall and footwall boundaries. Figure 9.3 shows an example of the mineralogical textures in the alloy mineralization type. The mixed mineralization type typically occurs as halos around the sulphide and alloy mineralization types and often represents a transition from sulphide to alloy mineralization types. The mixed mineralization contains varying amounts of sulphide (pentlandite and 46

59 heazlewoodite) along with awaruite in similar quantities (Table 9.1). Mineralization can occur as coarse sulphide-magnetite blebs associated with awaruite or as finely disseminated discrete grains. Figure 9.4 shows an example of the mineralogical textures in the mixed mineralization type. Figure 9.2 Sulphide Mineralization Type (EXP_18) Figure supplied by Royal Nickel Corporation. Figure 9.3 Alloy Mineralization Type (EXP_43) Figure supplied by Royal Nickel Corporation. 47

60 Figure 9.5 shows the distribution of the pentlandite, heazlewoodite and awaruite mineralization types within the Dumont deposit. Figure 9.4 Mixed Mineralization Type (EXP_15) Figure supplied by Royal Nickel Corporation. Figure 9.5 Block Model for the Dumont Project Illustrating the Mineralization Type Distribution Figure supplied by Royal Nickel Corporation. 48

61 9.2 CONTACT-TYPE NICKEL-COPPER-PGE MINERALIZATION All contact-type nickel-copper-pge mineralization described is outside the current mineral resource estimate and the following description is provided for background information, summarized from Lewis and San Martin, August, Drilling by Royal Nickel has confirmed the occurrence and grade of the historically identified mineralization at the basal contact at the eastern end of the Dumont sill. Drill hole 08-RN-71 intersected 0.8 m of semi-massive pyrrhotite grading 0.99% nickel, 0.19% copper, 0.3 g/t platinum, 1.0 g/t palladium and 0.07 g/t gold at the contact between the Dumont intrusive and footwall volcanics. This contact-related mineralization appears to be restricted in extent. Royal Nickel has drilled several holes through the footwall of the Dumont intrusion, to test weak, vertical-axis Time Domain Electromagnetic (VTEM) anomalies. The holes intersected barren pyrrhotite-pyrite mineralization in the footwall volcanics in proximity to the contact, but, no nickel-bearing sulphides were found. 9.3 OTHER TYPES OF PGE MINERALIZATION Royal Nickel s drilling has further delineated three anomalous PGE horizons other than the basal contact type described above. In 2008, a PGE horizon associated with the pyroxenite layer, overlying the upper peridotite, was identified. This zone varies in thickness from 1.5 to 22.0 m with grades ranging from 0.01 to 0.16% nickel, 0.08 to 0.39 g/t platinum, and 0.04 to 0.34 g/t palladium. The second PGE horizon, which lies under the main sulphide body, was previously identified during research on the historical drilling (Brügmann, 1990). This zone ranges from 0.4 to 34.5 m thick with grades ranging from 0.18 to 1.37% nickel, 0.01 to 0.76 g/t platinum, and 0.01 to 0.14 g/t palladium. The remaining PGE horizon was discovered by Royal Nickel in 2008 and is located approximately 100 m below the lowest sulphide body near the dunite contact with the lower peridotite. This horizon is ranges from 1.0 to m thick with grades ranging from 0.09 to 0.49% nickel, to 0.84 g/t platinum, and 0.03 to 1.86 g/t palladium. These horizons are generally observed to be continuous along strike and dip where drilling is present. Samples from each PGE horizon were sent to Memorial University for analysis using scanning electron microscope. This work identified that the PGE phases are similar in all horizons and consist of three alloys: palladium/tin (Pd/Sn), platinum/copper (Pt/Cu), and platinum/nickel (Pt/Ni) which are intimately associated with nickel sulphides. This mineralization is not included in the 2010 mineral resource estimate. 49

62 10.0 EXPLORATION A description of the work conducted on the Dumont Property prior to Royal Nickel s 2010 exploration program is provided in Section 6.0. The details of Royal Nickel s exploration activities since its acquisition of the Dumont Property in 2007 are contained in prior Micon Technical Reports EXPLORATION PROGRAM Since acquiring the right to explore the Dumont Property, Royal Nickel has undertaken an aggressive exploration program to evaluate and develop the mineral resources on the property. The 2010 exploration programs completed to date include resource definition drilling, structural drilling and modelling, geotechnical (rock mechanics) drilling and studies, pilot plant test holes (NQ), geotechnical (overburden) holes, pilot plant sample holes (PQ), geological mapping, mineralogical mapping sampling and overburden modelling. Details of the Royal Nickel drilling programs are provided in Table Table 10.2 outlines the expenditures for the 2010 exploration program from January 1, 2010 to July 31, Resource Definition Drilling The 2010 resource drilling program consisted of four holes totalling 2,353 m. These holes were drilled between sections 6900E and 7100E for the purpose of delineating the unconstrained down-dip extent of mineralization in this portion of the deposit. Previous drilling indicated that local faulting in this area has resulted in the relative offset of mineralization in structural domain 5 to the northeast, along the southeasterly dipping fault that defines the boundary between domains 4 and 5. Previous 100 m by 100 m drilling coverage in this area was inadequate to define this mineralization down to 450 m in depth because of the fault offset and the difficulty associated with drilling through the fault. The 2010 resource definition drilling program indicates that the mineralization in domain 5 extends down-dip to the northeast below the fault. Details of the drilling program are provided in Section Structural Drilling and Modelling The structural drilling program defined major geological structures (faults) within the central portion of the deposit. The results have been incorporated into the drilling database and applied in the August, 2010 mineral resource estimate contained in Section 17 of this report. 50

63 Table 10.1 Summary of the Royal Nickel Drilling Programs on the Dumont Property by Category to Total Purpose of Drilling Number of Total Number of Total Number Number of Total Total Metres Holes Metres Holes Metres of Holes Holes Metres Twin Hole 5 1, ,681 Sectional Resource Definition , , , ,912 Structural 4 1, ,359 Geotechnical (Rock Mechanics) 3 1, ,503 Pilot Plant Test Holes (NQ) 7 1, ,757 Total Drilling included in the Current Resource Estimate ,212 Metallurgical Sample Drilling , ,194 Crushing Testwork Sample Drilling Geotechnical Drilling (overburden) PQ Test Hole Drilling 13 2, ,785 Total , , , ,701 Table supplied by Royal Nickel Corporation.

64 Table 10.2 Dumont Property 2010 Exploration Expenditures (From January 1, 2010 to July 31, 2010) Category Cost (C$) Drilling Resource drilling 258,243 Metallurgical drilling 1,075,947 Assaying 174,998 Subtotal Drilling 1,509,188 Metallurgy Process development and testwork 1,287,540 Mineralogy 57,528 Mini pilot plant 955,978 Process plant scoping 69,104 Subtotal Metallurgy 2,370,150 Geology Geometallurgy 193,780 Environmental Environmental studies 26,392 Environmental permitting 0 Independent audits 1,906 Subtotal Environmental 28,298 Development Studies Mining engineering /geotechnical studies 142,625 Scoping studies 257,887 Subtotal Development Studies 400,512 Property Mining rights renewal 10,035 Surface rights access and options to purchase 540,318 Subtotal Property 550,353 Subtotal 2010 program 5,052,282 Salaries 848,100 Administration 282,317 Total 2010 program 6,182,699 Table of expenditures provided by Royal Nickel Corporation Geotechnical Drilling and Studies The geotechnical (rock mechanics) drilling program defined rock mass characteristics and evaluated open pit slope angles. The results were incorporated into the drilling database and the mineral resource estimate of August,

65 Pilot Plant Test Drill Holes (NQ) A series of seven pilot plant test drill holes totalling 1,757 m was completed in The purpose of the pilot plant test hole (NQ) program was to characterize the near surface mineralization such that representative mineralization domains could be selected for sampling by larger diameter (PQ) drill holes. The assay results have now been incorporated into the drilling database and the resource estimate, and have been used to define metallurgical domain composites for pilot plant testing. These holes have also undergone mineralogical sampling for the purpose of metallurgical domain composite characterization Geotechnical (Overburden) Drill Holes The geotechnical (overburden) drill hole program consisted of five holes totalling 104 m. This initial program was designed to characterize the overburden material located above the indicated resources in order to aid in engineering work for a future preliminary assessment. The program also allowed for the installation of three piezometers for future groundwater measurements and will serve as the basis for future geotechnical investigations. The data regarding the depth to bedrock for these drill holes is currently available; however, the final overburden characterization report is pending Pilot Plant Sample Drill Holes (PQ) The pilot plant sample hole (PQ) drilling program consisted of thirteen holes totalling 2,785 metres. This program was based on the results of the pilot plant test hole (NQ) program described previously. The objective of the drilling was to provide representative mineralogical variability in a larger sample size for testwork at Royal Nickel s pilot plant located in Thetford Mines, Quebec. Multiple holes were planned on each site in order to acquire a sufficient sample of each metallurgical domain Geological Mapping Seventeen days of surface mapping were completed between June 9 and July 15, The mapping program focused on a package of 10 contiguous claims to the west of the Dumont sill. The mapping team completed 100 m spaced traverses to locate, describe and sample surface outcrops. A description of the vegetation was made at every 50 m interval along each traverse. The data collected will be used to assess potential mining infrastructure development Mineralogical Sampling Mineralogical sampling of Dumont core began in The mineralogical sampling program uses the SGS Lakefield Research (SGS Lakefield) EXPLOMIN TM Particle Scan analysis to provide detailed mineralogical information on mineral assemblages, nickel deportment, liberation, alteration and the variability of these factors. Mineralogical samples were taken for the purpose of metallurgical domain composite characterization and for the purpose of mineralogical mapping of the Dumont deposit. 53

66 Mineralogical mapping sample locations were planned so as to provide representative spatial and compositional down drill hole traces for holes on even numbered sections from 5800E to 9000E, with the goal of providing comprehensive representation of the mineralogical variability of the deposit. A total of 732 mineralogical mapping samples were collected as of August 16, Metallurgical domain composite characterization samples were selected on an ongoing basis to represent the mineralogy of each metallurgical domain composite as defined for testwork. This includes all domain composites described in Section 16.4 as well as all metallurgical composites defined in the pilot plant test drill holes. A total of 70 metallurgical domain characterization samples were collected as of August 16, The sampling and analytical procedures for both types of samples are identical and described in Sections 12.2 and Overburden Modelling The overburden modelling program was completed in June, 2010, to incorporate the new data from the geotechnical (overburden) drill hole program with all the existing drill hole data, outcrop mapping data, and LIDAR survey data to produce one comprehensive overburden thickness map for the Dumont Property. The overburden thickness map is bounded to the northeast by the arctic watershed and to the south by Highway 111. This model will be used to assess potential mining infrastructure development and will serve as the basis to plan future geotechnical studies EXPLORATION RESULTS Results of the non-resource definition drilling program, including geotechnical (overburden) holes, pilot plant sample holes (PQ), geological mapping, mineralogical sampling, and overburden modelling, are described in this section. Results of the exploration program related to the sectional resource definition drilling program are described in Section Figure 10.1 is a geological map showing the collar locations of the non-resource drilling. 54

67 Figure 10.1 Geological Map Showing the Location of the Non-Resource Drilling Figure supplied by Royal Nickel Corporation Results of the Geotechnical (Overburden) Drilling Program The geotechnical (overburden) drill hole program consisted of five holes totalling 104 m. The geotechnical drill hole details are summarized in Table The analysis and conclusions of this study are pending. The 2010 geotechnical (overburden) drill hole program will serve as the basis for future geotechnical investigations. Table 10.3 Summary for the Oriented Structural Drill Holes Drill Hole UTM Coordinates Elevation Final Azimuth Dip Section Piezometer Number Easting (m) Northing (m) (ASL m) Depth (m) ( ) ( ) Line Installed GD E Yes GD E Yes GD E Yes GD E No GD E No Total Table supplied by Royal Nickel Corporation. 55

68 Pilot Plant Test Drill Holes (NQ) Results Assay results and logging observations from the pilot plant test holes were used to subdivide each entire drill hole into metallurgical domain composites. These metallurgical domain composites are defined as zones of homogeneous metallurgical response in accordance with the principles described in Section Parameters used to define the extent of each composite include the nickel-to-sulphur ratio of the rock, nickel grade, alteration intensity, nickel mineralogy and grain size. Figure 10.2 displays typical domaining of a drill hole.. Figure 10.2 Typical Domaining of a Drill Hole (09-RNC-218A-I) Figure supplied by Royal Nickel Corporation. 56

69 Results of the Pilot Plant Sample Drill Holes (PQ) Results The pilot plant sample hole (PQ) drilling program consisted of twelve holes totalling 2,785 m. The details for these drill holes are summarized in Table Table 10.4 Summary for the Pilot Plant Sample Drill Holes (PQ) Program Drill Hole UTM Coordinates Elevation Final Azimuth Dip Section Number Easting (m) Northing (m) (ASL m) Depth (m) ( ) ( ) Line 10-RN-216-P E 10-RN-216-P E 10-RN-217-P E 10-RN-217-P E 10-RN-217-P E 10-RN-218-P E 10-RN-218-P E 10-RN-218-P E 10-RN-218-P E 10-RN-222-P E 10-RN-222-P E 10-RN-222-P E 10-RN-222-P E Total 2,785.0 Table supplied by Royal Nickel Corporation. The pilot plant sample drill holes (PQ) are sampled according to the variability domain composites defined in the pilot holes. Samples are stored on site in Amos until required for testwork in the Royal Nickel pilot plant Results of the Mineralogical Sampling Program Analytical results from the mineralogical mapping samples and the mineralogical domain composite characterization samples were used to develop the mineralogical block model for the Dumont deposit. Statistics for the major mineral components of the 336 mineralogical samples received to August 16, 2010 are given in Table 10.5Error! Reference source not found.. The location of all drill holes where sample results have been incorporated into the mineralogical block model, and all planned mineralogical mapping samples are shown in Figure

70 Table 10.5 Summary of Principal Minerals in Mineralogical Samples Received to August 16, 2010 Mineral Mass (%) Median Mean Max Min Serpentine Magnetite Brucite Olivine Pentlandite Heazlewoodite Awaruite Table supplied by Royal Nickel Corporation. Figure 10.3 Drill Hole Locations with Mineralogical Mapping Sample Results and Planned Sampling Figure supplied by Royal Nickel Corporation Results of the Overburden Modelling Program The overburden modelling program was completed in June, 2010, and incorporates the data from the geotechnical (overburden) drill hole program, as well as all available data from surface mapping, drilling programs completed since 2007, and LIDAR topography survey data, into one comprehensive overburden thickness map (Figure 10.4) for the Dumont 58

71 Property. The data analysis and interpolation model were completed for an area bounded by the arctic watershed and Highway 111 to estimate overburden thickness. This model will be used to assess potential mining infrastructure development and will serve as the basis to plan future geotechnical studies. Figure 10.4 Overburden Thickness Map Figure supplied by Royal Nickel Corporation Results of the Crushing Testwork Hardness Domain Sampling A sampling program for the purpose of providing samples representative of the variability of rock hardness and grindability for the Dumont deposit was designed in consultation with Xstrata Process Support (XPS) to provide data for the design of the potential crushing plant and grinding circuits. Four hardness domains were defined based on data obtained by point load testing and in accordance with statistical analysis completed by XPS. Three of these hardness domains (Domains 1 to 3) correspond to variation in serpentinization alteration and a fourth domain represents fault zones. Holes 09-RN-215 and 09-RN-210 were drilled to provide samples for Domains 1 and 2. Hole 09-RN-211 was drilled to provide samples for Domain 3. The fault zone sample was a composite sample collected from fault zones of holes drilled for the three other domains. The samples chosen are believed to be representative of the properties displayed by the majority of the rock in the respective domains. These samples have been submitted for a full JKTECH drop-weight evaluation, semiautogeneous (SAG) grinding, and unconfined compressive strength (UCS) testing at Hazen Research Inc. (Hazen) under the supervision of XPS. 59

72 The results of the crushing testwork domain sampling are summarized in Table 10.6, Table 10.7 and Table

73 Table 10.6 Summary of the Drop-Weight Breakage Evaluation Value Parameter Domain 1 Domain 2 Domain 3 SG (by weighing in water and air) JKSimMet parameters A (maximum breakage) b (relation between energy and impact breakage) A x b (overall AG SAG hardness) t a (abrasion parameter) Resistance to impact breakage Medium Soft Moderately soft Resistance to abrasion breakage Medium Soft Moderately soft Table supplied by Royal Nickel Corporation. Table 10.7 Summary of the SMC Break Evaluation Value Parameter Domain 1 Domain 2 Domain 3 SG (by weighing in water and air) SMCT parameters A (maximum breakage) b (relation between energy and impact breakage) A x b (overall AG SAG hardness) SMC Test DW i kwh/m DW i % M ia kwh/t M ih kwh/t M ic kwh/t t a Table supplied by Royal Nickel Corporation. Table 10.8 Summary of the UCS Results HRI Domain Compressive Strength (psi) , , , , , ,640 Table supplied by Royal Nickel Corporation. 61

74 11.0 DRILLING The following has been extracted from Lewis and San Martin, August, Earlier drilling programs are described in Section 6.0 of this report RESOURCE DEFINITION DRILLING PROGRAM The 2010 resource drilling program consisted of four holes totalling 2,353 m drilled between sections 6900E and 7100E, for the purpose of delineating the unconstrained down-dip extent of mineralization in this part of the deposit. The sectional resource drilling program, initiated in 2007, was designed to maintain a nominal 100 m spacing between holes within the plane of the section and along strike, except in the 50 m by 50 m variability testing blocks where 50 m spacing was maintained. The program was designed to define mineralization down to a nominal depth of 400 m from surface (-90 m elevation). In places, drilling has investigated mineralization down to a depth of 650 m (-340 m elevation). Figure 11.1 illustrates the location of all holes completed during the sectional resource definition and exploration drilling program. Figure 11.1 Locations of the Resource Definition Drill Hole Collars on the Dumont Property Figure supplied by Royal Nickel Corporation. 62

75 A summary of all of the drilling conducted on the property since 2007 has been provided previously in Table Table 11.1 summarizes the details of the 2010 resource definition holes. Drill Hole Number UTM Easting (m) Table 11.1 Sectional Drilling Program Drill Hole Collar Information UTM Northing (m) Elevation (ASL m) Final Depth (m) Azimuth ( ) Dip ( ) Section Line In Previous NI Technical Report 10-RN E No 10-RN E No 10-RN E No 10-RN E No Table provided by Royal Nickel Corporation. Royal Nickel contracted Forages M. Rouillier (Rouillier) of Amos, Quebec to conduct the drilling program. Rouillier used custom-built diamond drill rigs mounted on skids or selfpropelled tracked vehicles with NQ diameter diamond drill coring tools. Rouillier is an independent diamond drilling contractor that holds no interest in Royal Nickel. Figure 11.2 shows one of Rouillier s drills set up on a hole and drilling during Micon s July, 2010 site visit. Figure 11.2 View of a Rouillier Drill Rig (July, 2010 Site Visit) Drill hole directional surveys were conducted using a Maxibor down-hole survey tool which calculates the spatial coordinates along the drill hole path based on optical measurements of direction changes and gravimetric measurements of dip changes. 63

76 All geological, engineering and supervision portions of the drilling program were overseen by geological staff of Royal Nickel, principally Dr. John Guo, P.Geol., OGQ, and Mr. Lorne Burden, P.Geo., supervised by Mr. Alger St-Jean, P.Geo., Vice-President Exploration for Royal Nickel. The 2010 drilling program was designed by the geological staff of Royal Nickel with input from Micon, based on the analysis of the results of the previous exploration and drilling programs on the Dumont Property RESULTS OF THE 2010 SECTIONAL DRILLING PROGRAM The current drilling results are consistent with both the historical results obtained by prior operators and the previous drilling by Royal Nickel from 2007 to Both the historical and Royal Nickel drilling results indicate that the disseminated nickel mineralization is primarily constrained to the dunite unit. The drilling also indicates that the mineralization occurs as bands of variable nickel grade which are generally conformable to the orientation of the contacts of the Dumont sill and trend along an azimuth of 315 and dip 70 to the northeast. The 2010 resource drilling program consisted of four holes totalling 2,353 m drilled between sections 6900E and 7100E for the purpose of delineating the unconstrained down-dip extent of mineralization in this portion of the deposit. Previous drilling indicated that local faulting in this area has resulted in the relative offset of mineralization in structural domain 5 to the northeast, along the southeasterly dipping fault that defines the boundary between domains 4 and 5. Previous 100 m by 100 m drilling coverage in this area was inadequate to fully drill off this mineralization down to 450 m depth because of the fault offset and the difficulty in drilling through the fault. The 2010 resource definition drilling program demonstrated that the mineralization in domain 5 extends down-dip to the northeast below the fault, as shown in Figure 11.3, Figure 11.4 and Figure The mineral resource has been correspondingly expanded in this area. Table 11.2 is a compilation of mineralized intersections from the 2010 Royal Nickel drilling program at 0.25% nickel cut-off grade. Table 11.2 also contains mineralized intersections from the structural, geotechnical and pilot plant test holes described in an earlier Technical Report, dated April 5, At that time, assay results from the holes drilled as part of this program were pending. These results have now been received and incorporated into the drilling database and August, 2010 resource estimate. 64

77 Figure 11.3 Plan and Sectional View (7000E) Illustrating the Drill Hole Distribution and Assay Results Viewing Direction is Northwest (Azimuth = 315 ) Figure supplied by Royal Nickel Corporation. 65

78 Figure 11.4 Plan and Sectional View (7100E) Illustrating the Drill Hole Distribution and Assay Results Viewing Direction is Northwest (Azimuth = 315 ) Figure supplied by Royal Nickel Corporation. 66

79 Figure 11.5 Plan and Sectional View (7200E) Illustrating the Drill Hole Distribution and Assay Results Viewing Direction is Northwest (Azimuth = 315 ) Figure supplied by Royal Nickel Corporation. 67

80 Table Drilling Program Mineralized Intersections at a 0.25% Nickel Cut-off Grade Drill Hole ID 09-RN RN RN RN-208 Drill Hole Intersection (m) Assay Results From To Length Nickel (%) Purpose Drill Hole ID Drill Hole Intersection (m) Assay Results From To Length Nickel (%) RN Geotech RN Structural RN Structural Structural RN Pilot hole 10-RN RN Pilot hole RN Pilot hole RN RN Pilot hole RN RN Pilot hole Table provided by Royal Nickel Corporation. Purpose Pilot hole Pilot hole Resource Resource Resource Resource 68

81 12.0 SAMPLING METHOD AND APPROACH Descriptions of the historical sampling methods and approaches for the Dumont Property have been previously provided in Section 6 of this report, if available. The following has been extracted from Lewis and San Martin, August, Royal Nickel has implemented a QA/QC program. Micon discussed this program with both the consulting staff and Royal Nickel staff during the site visits in 2007, 2008, 2009 and Micon is satisfied that Royal Nickel has implemented an appropriate QA/QC program for the Dumont Property ASSAY/GEOCHEMICAL SAMPLING Diamond drilling sampling controls start after a run has been completed and the rods are pulled out of the drill hole. The core is removed from the core barrel and placed in core boxes. The capacity of each box depends on the diameter of core stored in it (1.5 m for PQ diameter, 2.40 m for HQ diameter or 4.50 m for NQ diameter). This follows standard industry procedures. Small wooden tags mark the distance drilled in metres at the end of each run. On each filled core box, the drill hole number and sequential box numbers are marked by the drill helper and checked by the geologist. Once the core box is filled at the drill site, the box is covered with a lid to protect the core and the box is sent to the core logging facility in Amos for further processing. In general, the core recovery for the diamond drill holes on the Dumont Property has been better than 98% and little core loss due to poor drilling methods or procedures has been experienced. Upon receipt at the Amos facilities, the core boxes are opened and placed sequentially on logging tables where the core is washed, logged and sampled. Core is sampled for geochemical assay by defining consecutive intervals of 1.5 m maximum length. The 1.5 m maximum sampling interval was chosen because it is generally close to the five-foot sample interval used during the historical drilling programs. The marked 1.5 m interval is cut down the middle to produce representative halves. One half is sent to ALS-Chemex for crushing and assaying while the remainder is placed back in the box for future reference (Figure 12.1). Figure 12.1 Assay Sampling Protocol Figure supplied by Royal Nickel Corporation. 69

82 During the assay/geochemical sampling process for resource definition, a unique sample tag was inserted into the sample bag, with each sample and each sample bag marked with its individual sample number. The bags containing the blank and standard samples were added into the sequential numbering system prior to being shipped to the assay preparation facilities of Expert Laboratories in Rouyn-Noranda or ALS-Chemex in Timmins. No field duplicate samples were added to the control samples during the sampling of the initial drilling program which twinned historical holes. Royal Nickel began regularly inserting field duplicate samples into the sampling stream beginning with drill hole 07-RN-04. The drill core sampling procedures practiced by Royal Nickel are among the commonly accepted procedures used throughout the mineral industry. Along with in-house standards, blanks and duplicates included in the sample stream, routine check assays are conducted on the samples by a second laboratory. Micon has reviewed the drill core sampling practices of Royal Nickel and finds that the sampling quality is both representative of the mineralization identified at the Dumont Property and of a sufficient nature upon which to base a mineral resource estimate. The significant 2010 drilling intersections for the Dumont Property have been previously summarized in Table The significant drilling intersections prior to 2010 are contained in the previous Micon Technical Reports MINERALOGICAL MAPPING SAMPLING The mineralogical mapping sampling program uses the previously drilled and sampled NQ half-core from the resource drilling programs. The core has been stored indoors in a dry storage facility since the completion of the drilling. The core from each hole is laid out on the core logging tables in sequential order. Using the previously acquired assay results from each hole, the geological staff classifies the core into one of three nickel mineral deportment domains: sulphide, alloy or mixed. These domains are then further subdivided based on nickel grade, grain size and degree of alteration. Once all the mineralogical domains are categorized within a hole, a representative sample is chosen from each domain and a location for thin section sampling is marked off within the sample interval. These samples are taken to coincide with the sample intervals defined in the original logging as described previously. The selected sample is given a new unique mineralogical sample ID, photographed, and sent to the core cutting area. Technicians quarter split the half-core. One quarter of the half-core is placed back into the box for reference and returned to storage, and the remaining quartered core is placed in another similarly labelled sample bag (Figure 12.2). Before the bag of quartered core is sealed, a representative piece, 3.0 to 8.0 cm in length is selected and placed in a separate storage box which acts as a quick mineralogical reference and further builds Royal Nickel s core sample library. The preselected thin section slice is cut from the quarter core that is selected for mineralogical analysis. The thin section slice is placed in a labelled sample bag. 70

83 Figure 12.2 EXPLOMIN TM Sampling Protocol The remaining half core is cut down the centre for mineralogical sampling Figure supplied by Royal Nickel Corporation. The sample bag containing the thin section slice is sent directly to SGS Lakefield for thin section preparation and EXPLOMIN TM field stitch analysis. The sample bag containing the quarter core is sent to the ALS-Chemex preparation laboratory in Timmins for stage crushing and pulverization as detailed in Table Table 12.1 EXPLOMIN TM Mineralogical Sample Preparation Procedure at ALS-Chemex Laboratories Mineralogical Sample Preparation Procedures WEI-21 Weigh and log received sample LOG-22 Log sample CRU-31 Crush entire sample to > 70% passing 2 mm SPL-21 Riffle split 100g for pulverizing PUL-35 Stage pulverize, two 100g splits to 90% passing 106 microns WSH-22 Wash pulverizer CRU-QC Crush to 70% passing 2 mm PUL-QC Pulverize to 90% passing 150 mesh Table supplied by ALS-Chemex. The first 100 g split of pulverized material is sent to SGS Lakefield where the sample is prepared for EXPLOMIN TM particle scan mineralogy and XRF Borate Fusion assay. The results are forwarded to Royal Nickel and imported directly into the database. 71

84 The other 100 g split of the pulverized material is retained by ALS-Chemex for chemical analyses. The reject material is sent back to the Royal Nickel s Amos office for storage. The results are forwarded to Royal Nickel and imported directly into the database PILOT PLANT SAMPLING PQ core metallurgical domain composite samples are selected based on nickel deportment, grade and alteration of the rocks as determined through assays and mineralogical sampling of the NQ pilot hole (as discussed in Section and in the previous Technical Report dated April, 2010). A 1.5 m PQ drilling grid was established around each NQ pilot hole to plan multiple PQ holes on the same site, in order to accommodate the larger sample volume required (approximately 1,800 kg/domain sample) while maintaining domain sample uniformity. As a result of hole proximity and the inherent difficulty and cost of PQ drilling in overburden, a percussion well drilling rig was employed to push casing into the bedrock for the multiple holes required on each of the sites. When the well drilling rig reached bedrock, the diamond drill returned to drill the PQ core domain samples. The sampling method for PQ core is identical to that described previously up to and including the geotechnical logging, after which there is a change in procedure. After geotechnical logging, the core is thoroughly washed to remove any drilling additives that may interfere with the metallurgical testwork. The PQ core is then checked for comparability to the pilot hole, by comparing lithological contacts, mineralization, alteration, and structural features. The core is then logged for lithology, and metallurgical domain composite samples are delineated which reflect those established in the pilot NQ hole. The core is then photographed and placed in short-term indoor storage to await sampling. The PQ sampling program is supervised by an independent qualified engineer provided by Stavibel Inc. (Stavibel) to ensure quality control of the sampling method and to certify chain of custody. The rock is weighed and transferred by domain sample from the core boxes directly into 200-L plastic barrels fitted with Schrader valves. The domain samples are kept separate and barrels are filled in sequential order. A barrel typically holds from 250 kg to 270 kg of rock. The engineer seals the full barrel and places a numbered tag on the closure to prevent or identify any possible tampering. The barrels are purged with nitrogen to prevent oxidation and degradation of the rock while the sample awaits metallurgical testwork. When the sample is required by Royal Nickel s metallurgical group, the barrels are shipped directly via road freight to the pilot plant in Thetford Mines, Quebec. 72

85 13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY The following has been extracted from Lewis and San Martin, August, Sample preparation, analysis and security details are only partly available for the historical work conducted on the Dumont Property during the 1970 s and 1980 s. Where the details are available, they are provided in Section 6.0 of this report. Prior to the initial drilling program conducted in 2007, Royal Nickel did not conduct any sample preparation or analysis, as no samples were collected from the property during the period leading up to the drilling program SAMPLE COLLECTION AND TRANSPORTATION Diamond drill core is stored in closed wood core boxes and is transported by the drill contractor to the Royal Nickel core logging facility in Amos twice a day. The core is stored sequentially hole by hole in racks for logging. The diamond drill core sampling is conducted by a team of several staff geological assistants under the close supervision of the Royal Nickel geologist in charge of the program on site. The Royal Nickel staff geologists are responsible for the integrity of the samples from the time they are taken until they are shipped to the preparation facilities in Rouyn-Noranda or Timmins Core Logging and Sampling Once the core boxes arrived at the logging facility in Amos, the boxes were laid out in order, the lids were removed and the head of the first box was marked in red to denote the starting point of the drill hole. The core was then laid out on the logging table, and washed to remove any grease and dirt which may have entered the boxes. Core logging consists of two major parts: geotechnical logging and geological logging. The geotechnical logging is completed first to check the core pieces for best fit and to determine core recovery, Rock Quality Designation (RQD), Index of Rock Strength (IRS) and magnetic susceptibility. The number of open (natural) fractures in the core is counted and the fracture surfaces are evaluated for their joint surface condition. Geological logging follows and is comprised of recording the lithology, alteration, texture, colour, mineralization, structure and sample intervals. All geotechnical and geological logging and sample data are recorded directly into a computerized database using Century System s DHLogger data logging software. During the core logging process the geologists define the sample contacts and designate the axis along which to split the core with special attention paid to the mineralized zones to ensure representative splits. All core which is classified as dunite by the geological logging is marked in 1.5 m intervals for sampling. Any mineralized sections outside the dunite are 73

86 also marked for sampling. Outside the dunite unit a minimum of one, 1.5 m control sample in every 10 m of core is taken. See Figure 13.1 for a photograph of the core logging facilities in Amos. Figure 13.1 Core Logging Facilities in Amos Samples are identified by inserting three identical pre-fabricated, sequentially-numbered, weather-resistant sample tags at the end of each sample interval. Once the core is logged, photographed and the samples marked, the core boxes are transferred to the cutting room for sampling. Sections marked for sampling are split using a diamond saw except in the case of extreme rock hardness where a hydraulic splitter is used. Once the core is split in half, one half is placed into a plastic sample bag and the other half is returned to the core box. The core cutting technicians verify that the interval on the sample tag matches the markings on the core and that the sample tag matches the sample number on the bag. The half of the cut core returned to the core box is then re-marked by the core technician with a grease pencil to indicate the end of the sample interval. The boxes containing the remaining half core are stacked and stored on site in the secure core storage facility. Duplicate, blank and standard samples are inserted into the sample stream at regular intervals using a sequential numbering scheme set up by Royal Nickel. Once the sample is placed in its plastic sample bag, the bag is secured with electrical tie wraps and the sample bags are placed into large fabrene sacks. Generally, seven sample bags are placed into each fabrene bag and then the bag is secured with an electrical tie wrap. The fabrene sample bags remain secured in the core shack in Amos until they are shipped to the laboratory by courier. The general shipping rate for the samples is once for every 100 to 150 samples. 74

87 After-hours access to the core logging, core cutting and core storage facilities, as well as the project office, is controlled by a zoned alarm system with access restrictions based on employee function Sample Preparation and Analysis As of June 1, 2008, Royal Nickel s samples have been prepared at ALS-Chemex s preparation facility in Timmins, Ontario and analyzed at ALS-Chemex s laboratory in Vancouver, British Columbia. Both the preparatory facility and assay laboratory have ISO 9001:2000 certification. Expert Laboratories, located in Rouyn-Noranda, Quebec is not ISO certified; however it does participated in the CANMET round robin proficiency testing twice yearly. John Guo P.Geol, OGQ, the Chief Project Geologist for the Dumont project, performs an annual inspection of Expert Laboratories in Rouyn-Noranda and the ALS- Chemex sample preparation facility in Timmins. Prior to June 1, 2008, all samples were assayed at Expert Laboratories and then all the pulps were re-assayed at ALS-Chemex. Currently, five percent of each assay batch returned from ALS-Chemex is randomly selected for check assay at Expert Laboratories. Once the samples reach ALS-Chemex s Timmins preparation laboratory each sample is dried as needed, crushed, and split into reject and a 250-g aliquot for pulverization. After pulverization the 250-g pulverized sample aliquot is again split into a 150-g master sample and a 100 g analytical sample. The 150-g master sample is stored in the Timmins facility for reference and the 100-g analytical sample is forwarded to the ALS-Chemex analytical laboratory for assaying in Vancouver. On receipt in Vancouver, the specific gravity of the analytical sample material is measured, and this is followed by a 35-element analysis using an aqua regia digestion and ICP-AES finish. Where reported nickel values exceed 4,000 ppm, a second analysis is completed from the 100-g analytical sample using a four acid total digestion with an ICP-AES finish. This 4,000 ppm threshold reanalysis was raised to 10,000 ppm on June 1, In addition, all samples are assayed for precious metals (gold, platinum, palladium) using a standard fire assay with an ICP-AES finish. After a holding period at the laboratories, all pulps and rejects are returned to Royal Nickel in Amos for long term storage. All analytical data are reconciled with the drill log sample records and recorded in the project database. For the purpose of geological and resource modelling, the ALS-Chemex aqua regia determinations are used for samples under 10,000 ppm nickel and the ALS-Chemex total digestion determinations are used for samples over 10,000 ppm nickel Control Samples As part of Royal Nickel s QA/QC procedures, a set of control samples comprised of a blank, a field duplicate and a standard reference material sample, are inserted sequentially into the sample stream. The cut core samples, along with the inserted control samples, are then shipped to the ALS-Chemex assay preparation facility in Timmins. 75

88 Blank Samples The blank samples used for the Dumont project consist of local esker sand. The esker sand is collected in 205-L drums by a local Amos construction contractor. Randomly selected samples were collected from the drum and assayed at ALS-Chemex to evaluate the composition of the sand and determine its suitability for use as a blank control sample. The assayed nickel grades from these samples range from 30 to 80 ppm. The qualified blank sample drum is sealed and placed at a clean place for further use. Royal Nickel sets 200 ppm nickel as the recommended upper limit of the blank sample value. The blank samples are submitted into the sample stream at the rate of approximately one for every 25 samples Duplicate Samples A duplicate sample is submitted into the sample stream at a rate of approximately one for every 25 samples. The sample and its duplicate consist of quartered core from the given sample interval. The remaining half-core is placed back into the core box for future reference Standard Reference Material Samples The Standard Reference Material Samples (SRMS) are inserted into the sample stream at the rate of approximately one for every 25 samples. Initially one high grade SRMS (OREAS 14P) was inserted into the sample stream for every three low-grade SRMS (OREAS 13P) submitted. On the phasing out of OREAS 13P and 14P, OREAS 70P is inserted into the sample stream at the same sample rate of one for every 25 samples. An exception to this occurs where logging personnel visually recognize zones of higher grade mineralization; through these high grade zones OREAS 72a is inserted. If the situation arises where the twenty-fifth sample is consistently located in between higher grade mineralization zones, a higher grade sample will be inserted outside the one-in-25 sequence to ensure that the higher grade zones are represented by standard reference materials. Four SRMS have been used in the project. The SRMS were prepared by Ore Research & Exploration Pty. Ltd. of Australia. Table 13.1 summarizes the specifications for the SRMS. Table 13.1 Summary of the Specifications for the Standard Reference Material samples Description Constituent Recommended Value 95% Confidence Interval Low High OREAS 13P Copper (ppm) 2,504 2,439 2,569 Gold (ppb) Nickel (ppm) 2,261 2,233 2,289 Palladium (ppb) Platinum (ppb) OREAS 14P Copper (%)

89 Description Constituent Recommended Value 95% Confidence Interval Low High Gold (ppb) Nickel (%) Palladium (ppb) Platinum (ppb) OREAS 70P Copper (ppm) Gold (ppb) Nickel (ppm) Aqua Regia 2,438 2,222 2,655 Nickel (ppm) 4 Acid 2,730 2,620 2,841 Palladium (ppb) <1 IND IND OREAS 72a Copper (ppm) Gold (ppb) Nickel (%) 4 Acid Palladium (ppb) Platinum (ppb) Table supplied by Royal Nickel Corporation MINERALOGICAL MAPPING SAMPLING The mineralogical mapping sampling program uses the SGS EXPLOMIN TM analysis to provide detailed mineralogical information on mineral assemblages, nickel deportment, liberation, alteration and the variability of these factors. Mineralogical samples were taken for the purpose of mineralogical domain composite characterization and for the purpose of mineralogical mapping of the Dumont deposit Sample Definition and Sampling The mineralogical mapping sampling program samples a quarter of the NQ core previously drilled for the resource definition program. In areas of interest, sample length and location are defined to coincide with previous assay sample intervals so that a direct comparison can be made between results obtained from assay/geochemical analyses and mineralogical sampling results. The selected mineralogical mapping samples are given a unique sample identification number (ID), photographed, and sent to the core cutting area. Mineralogical mapping sampling is usually completed in batches, where multiple samples are selected from each hole, then cut sequentially. The half-core remaining from the previous assay sampling is quarter-split to produce the mineralogical sample. A portion of the quartered core is cut further to produce a pre-selected portion of rock for thin section field stitch analysis. The selected portion for field stitch analysis and the quartered core are each placed in separate bags, and identified by the same unique mineralogical mapping sample ID. For QA/QC purposes, a piece of the quartered core selected for mineralogical particle scan analysis is selected from the sample bag and placed in the Royal Nickel mineralogical mapping sampling library. 77

90 Once a sample is placed in its plastic bag, the bag is secured with staples. Typically, seven sample bags are placed into a cardboard box and secured with tape. The sealed boxes remain secured in the Amos core logging facilities until they are shipped to the laboratory using a courier service. Samples are shipped at the rate of 50 to 100 samples at one time. Blanks and standard samples are inserted into the sample stream at regular intervals using a sequential numbering scheme set up by Royal Nickel. The sample bag with the thin section slice is sent directly to SGS Lakefield for thin section preparation and mineralogical analysis. The sample bag containing the quarter core is sent first to ALS-Chemex s Timmins preparation laboratory for stage crushing and assaying, with a split shipped to SGS Lakefield for mineralogical particle scan analysis Sample Preparation and Analysis Upon receipt at ALS-Chemex s Timmins preparation laboratory the mineralogical samples are prepared according to the procedure summarized previously in Table Geochemical Preparation and Analysis Samples are chemically analyzed at the ALS-Chemex Laboratory in Vancouver, for specific gravity, followed by a 35-element analysis using an aqua regia digestion and ICP-AES finish. Where reported nickel values exceeded 10,000 ppm a second analysis is completed using a four acid total digestion with an ICP-AES finish. In addition, all samples are assayed for precious metals (gold, platinum, palladium) using a standard fire assay with an ICP-AES finish. Analysis results are forwarded to Royal Nickel and imported directly into the database Mineralogical Preparation and Analysis Mineralogical analysis and sample preparation are completed by SGS Lakefield in the following manner: "Upon sample receipt, the Sample Log-on technician verifies the received samples according to the sample list provided by Royal Nickel geologists. Any extra sample(s), discrepancies in identification, damage, contamination, unsuitable samples, concerns, or hazards are recorded, and Royal Nickel is notified. Once sample receipt is verified, samples are forwarded to the mineralogist for sample login and LIMS [laboratory information management system] reporting. The samples are kept in the same order that they appear on the documentation provided by Royal Nickel. For sample tracking purposes within SGS Lakefield, LIMS numbers are assigned to incoming samples. The LIMS number reflects the type of work being performed on the samples, the source of the samples, and secondary information such a Reference, Project, Batch, Quote, Link, Note, Category, Supervisor, Priority, Warning, Charge ID, Date Received, Date Requested. 78

91 When the LIMS log-in has been completed, a Project file is created to hold all the paperwork pertaining to the project. The project file is labelled with the Project number, LIMS number, and the Client or Company name. A log-in checklist is attached to the Project file and completed. A chain of custody is created. Record LIMS information is recorded in Diamond Services/Mineralogy project list. The project file is placed in a red folder and given to the Mineralogy Project Supervisor. Once the folder is checked by the Mineralogy Project Supervisor it is returned to Sample Login. Any additional information is updated in LIMS and the project list. The signed Chain of Custody is photocopied and the original is mailed to the client. Active Mineralogy Samples are stored with labels containing the project number, LIMS number, and test required. All of the samples are placed in one of the LIMS numbered, large plastic bags, placed in the To Do box. A copy of the work order accompanies the samples. When all requested analyses have been completed, samples are brought to Sample Tracking for storage. Boxes are stored in the Sample Tracking Room in Mineralogical Services for six months. After six months, the box is inventoried and the mineralogist is contacted for further instructions. Sample Preparation Using a binocular microscope, the Mineralogist or Project Mineralogist identifies the areas of interests previously marked by Royal Nickel staff for thin section analysis. One polished section for each sample is prepared for field stitch analysis. Sections are ground and polished then coated with carbon for analysis. Crushed samples that are received later on from ALS-Chemex are first riffle-split into 2 parts (of ~125g), one for mineralogy and one for assay. Each sample is potted in moulds and the necessary amount of resin and hardener is added. The moulds are placed into the pressure vessel and left under pressure for 5 hours. The moulds are then labelled and backfilled with resin. Then they are placed in the oven. The sections are ground and polished followed by carbon coating. Qemscan Operation The block holder is loaded with the samples. Measurement parameters (for core samples, field scan mode with 10 µm resolution and for crushed samples, PMA mode with 3 µm resolution) are set up. Stage Set-Up, Focus Calibration, Beam optimization and BSE Calibration are performed at the start of each run. After the runs are completed, the daily quality checks are preformed as summarized in [Table 13.2]. Weekly calibration and checks are also preformed to verify the following: Stage Initialization, Tilt Check, Rotation Check, X-Ray Detector Check, Gun Set-up, Brightness and Contrast, Filaments and Vacuum. The detectors are checked every three months. The Qemscan Data Validation report includes a measurement validation table and an assay reconciliation chart. Qemscan data are compared to externally measured chemical assay data to ensure measurement accuracy. Minerals are double-checked optically. A technical check is preformed on all data by a senior mineralogist." 79

92 Table 13.2 SGS Lakefield Daily Quality Checks for Qemscan Analysis Task/Duty Operational Purpose Management Purpose Checking correctness of PS placement. Statistics will readily show if samples and parameters are mismatched. Proper scheduling and quality control protocols. Check that analyses have been performed successfully. Go-, no-go decision to perform sample exchange for next analysis batch. Keep track of scheduling, processing and project management. Keep track of the measurement statistics as a matter of record To assist in optimizing analysis parameters and analysis times. Table supplied by SGS Lakefield. Optimization of analyses is influenced by the interdependence of PS-packing density and point-spacing For reviewing parameter selection criteria. Resolution vs. speed. If additional statistics are required for particle or modal accuracy, additional PS s may be required. Establishing accuracy and precision of measurement. Analytical results are forwarded to Royal Nickel and imported directly into the database Control Samples As a part of SGS Lakefield standard QA/QC procedures for Qemscan analysis, a standard sample is run every week. There are currently three standard samples from different projects that are cycled each time. One of the standards used is a Royal Nickel data validation sample. As part of Royal Nickel s QA/QC procedures for geochemical assays, a set of control samples comprised of a blank and standard reference material sample, are inserted sequentially into the sample stream. The cut mineralogical samples along with the inserted control samples are then shipped to the ALS-Chemex for stage crushing and chemical analysis. The standard reference materials and blanks used are analogous to those described previously with the exception that the frequency of insertion is increased to approximately one in every 15 samples PQ DRILLING The procedure for the sampling of PQ core to obtain metallurgical domain composite samples has been described previously in Section RESULTS OF THE QA/QC PROGRAM Micon has reviewed Royal Nickel s QA/QC program and procedures which follow the best practice guidelines currently in place for the industry. Micon has also reviewed the results of the QA/QC program and finds that they are acceptable to be used in a mineral resource estimate. 80

93 13.5 MICON COMMENTS ON THE QA/QC PROGRAM Micon considers that, based on a review of the QA/QC program and data and on discussions with Royal Nickel personnel, Royal Nickel applies a reasonable degree of care and diligence in monitoring the sample results on the property. Micon considers that, in general, the QA/QC procedures and protocols employed at the Dumont Property are rigorous enough to ensure that the sample data are appropriate for use in mineral resource estimations. It is Micon s opinion that the database and the procedures in-place at the Dumont Property are appropriate for use in a mineral resource estimate. 81

94 14.0 DATA VERIFICATION The following has been extracted from Lewis and San Martin, August, Due to the extensive exploration history on the Dumont Property, there is a large amount of sometimes conflicting information regarding the exploration programs. The original data were not in electronic form and Royal Nickel has converted the historical data into a digital format as part of its exploration program. Royal Nickel s logging facilities in Amos are located in a warehouse/business frontage style building on the town s outskirts (Figure 14.1 and Figure 14.2). Figure 14.1 Royal Nickel Office Core Logging Facilities in Amos Figure 14.2 Interior View of the Logging Facilities in Amos 82

95 Micon s initial site visit to the Dumont Property was conducted from May 8 to 9, 2007 with the assistance of Mr. Roland Horst, President and CEO of Royal Nickel, Mr. Alger St-Jean, P.Geo., Vice-President of Exploration for Royal Nickel and Mr. Lorne Burden, P.Geo., a consulting geologist who is overseeing Royal Nickel s exploration program, along with a number of other individuals from Royal Nickel and its consultants. During the site visit, five quarter-core samples were taken to independently verify the mineralization on the property. Micon s verification sampling was discussed within the August, 2007 and April, 2008 Technical Reports. Also during the initial site visit, the drill sites for the twinned drill holes conducted by Royal Nickel and the core logging facilities in Amos were visited. During the site visit Micon also noted that the geology as indicated by the drill holes matched the descriptions contained in the historical reports. Micon conducted a second site visit to the Dumont Property and the Royal Nickel exploration offices in Amos, from January 29 to 31, This site visit was conducted to review the database and details for the resource estimate. This site visit was conducted after initial discussions were held at the Royal Nickel offices in Toronto regarding the database and parameters upon which the resource estimate would be based. The details regarding Micon s data verification for the second site visit are contained in the April, 2008, Technical Report. Micon conducted a third series of visits to the Dumont Property from October 16 to 18, 2008 and from October 20 to 22, These visits coincided with the review of the database and drilling program prior to undertaking the updated resource estimate which was based on the 2008 drilling program. Micon s third trip to the property constituted the second visits to the site for both Alan San Martin, MAusIMM, and William Lewis, P.Geo. Mr. San Martin visited the site from October 16 to 18, 2008 to review the database and discuss Royal Nickel s strategy to update the resource estimate using the new information collected during the 2008 drilling program. Mr. Lewis visited the site between October 20 and 22, 2008 to review the geological interpretation for the mineral resource estimate and review Royal Nickel s drilling program and QA/QC program and data. Royal Nickel had outlined its preferences regarding the resource estimate concepts it wished to pursue during a meeting held in Toronto prior to the site visits and during the visit alternative methods and concepts regarding the resource estimate were also discussed. Micon conducted a fourth visit to the Dumont Property from October 16 to 18, The purpose of this visit was the full review of the drilling database and assay certificate crosschecks to assure the data integrity for the resource update. Micon s fourth trip to the property constituted the third visit to the site for Alan J. San Martin, MAusIMM. While on site Mr. San Martin performed an exhaustive validation of the entire database using a specific Excel based tool called DigDB which allows working with multiple spreadsheet files at once. After the validation process was complete, there was only one error found in the specific gravity table. No other errors were found, providing a high level of confidence in the drilling database. 83

96 For updating the mineral resource estimate on the Dumont Property, the amount of data collected by Royal Nickel during its drilling programs justified conducting the estimate without using any of the historical drilling data. Royal Nickel has completed enough drilling to fully dispense with the historical drilling. For the August, 2010 mineral resource estimate, Micon first reviewed the Golder model on June 11, 2010 with representatives from both Golder and Royal Nickel in Micon offices in Toronto. The model and its parameters were reviewed in detail during the 3-hour discussion. Micon then conducted a fifth site visit to the Dumont Property between July 20 and 23, During the visit the updated database was reviewed in detail, a number of drill sites were visited. The statistical data comparing the 50 m, 100 m and 200 m drill spacing were also reviewed and the Golder geometallurgical model was examined and discussed in detail as the basis for a resource estimate. As a result of its five site visits and discussions, Micon is satisfied that Royal Nickel s exploration program, QA/QC program and the work on the corresponding database have been appropriately undertaken and that the database can be used as the basis of the August, 2010 mineral resource estimate. Micon is also satisfied that the Golder geometallurgical model is suitable for use as the basis of not only the current mineral resource estimate and that the methodology employed is appropriate for use in any further resource estimates which may occur in the future. 84

97 15.0 ADJACENT PROPERTIES There are no immediately adjacent mineral properties which affect the interpretation of the geology or exploration potential of the Dumont Property. 85

98 16.0 MINERAL PROCESSING AND METALLURGICAL TESTING 16.1 OVERVIEW Metallurgical test work of Dumont samples can be categorized as follows: Initial testwork was conducted during as part of the feasibility study authored by Caron, Dufour, Séguin & Associates (CDS). This work showed that a process that entailed grinding, flotation and magnetic separation could produce two concentrates with a combined recovery of approximately 55.4%. The sulphide concentrate (containing 87% of recovered Ni) would grade 20% Ni, while the magnetic concentrate containing the remaining 13% of recovered Ni would grade only 1.9% and would probably not be saleable. Phase 1 of the current test program was conducted during The testwork was performed by SGS Minerals in Lakefield, Ontario, under the management of Royal Nickel s independent metallurgical consultants, Mineral Solutions. Tests were performed on composite samples representing the three different styles of mineralization that have been identified, namely sulphide, alloy and mixed. Different flowsheets were used for each style of mineralization, but all incorporated wet grinding as the initial stage. Encouraging results were obtained for sulphide mineralization, but the recovery and concentrate grade for both the alloy and mixed mineralization samples was poor due to excessive slimes and high viscosity. Phase 2 of the current program, which began in late 2008, was also managed by Mineral Solutions. During this phase, the process concept changed significantly, with the focus on pre-treatment to remove chrysotile fibres and brucite slimes. With the resulting reduction in slimes and lowered pulp viscosity, nickel recovery and concentrate grades improved markedly. This flowsheet was then developed into a standard process test to perform variability analysis on domain composite samples TECHNOLOGY PARTNERS The following organizations contributed to the Phase 2 test program: Mineral Solutions, based in Lakefield, Ontario. Mineral Solutions was retained as metallurgical consultant responsible for management of the test program and the overall flowsheet design. Centre de Technologie du Minérale et de Plasturgie (CTMP), a Crown corporation of the Government of Quebec, with laboratories located in Thetford Mines. CTMP has expertise in primary chrysotile ores and assisted in the selection of processes for removing chrysotile and slimes, as well as the nickel recovery flowsheet. Xstrata Process Support (XPS), independent consultants based in Sudbury, Ontario. XPS was responsible for managing the testwork required to select a crushing circuit. 86

99 Hazen Research Inc. (Hazen), based in Golden, Colorado. Hazen was responsible for performing various tests to select a crusher circuit. SGS Lakefield Research (SGS), based in Lakefield, Ontario. SGS conducted flotation test work PHASE 2 TESTWORK Comminution Testwork The concept for de-fibering the mill feed has been selected as part of the comminution circuit and is based on processes that have been successfully applied to primary chrysotile ores for many years. These include dry processes for reducing material to a size of less than 1 mm. Samples for comminution test work were selected jointly by Royal Nickel, XPS and Mineral Solutions. Statistical analysis of results from point load tests led to the definition of the following four different rock type domains: Hardness Domain 1 Relict olivine zone Hardness Domain 2 Coalingite (magnesium, iron carbonate) zone Hardness Domain 3 Black competent serpentinite Hardness Domain 4 Fault zones with strong alteration Tests performed at Hazen included the JKTech drop-weight evaluation, semi autogenous (SAG) mill comminution and unconfined compressive strength (UCS). Results of this testwork were used to evaluate the various dry-crushing flowsheet alternatives. Based on the flowsheets used by primary chrysotile operations, the design criteria for the crusher flowsheet included a maximum product size for 80% of material (P80) of 0.84 mm. Initial size reduction will use a conventional primary gyratory crusher. Following comparative testing of high-pressure grinding rolls (HPGR) and vertical shaft impact (VSI) crushers, the latter have been incorporated into the base case process flow sheet. These results were used to select the VSI option in the comminution circuit. Three Bond ball mill grindability tests were completed during Phase 1 of the testwork program. The metric work indices results were relatively high, ranging from 22.7 to 26.0 kwh/t Nickel Recovery Testwork The Phase 2 program tested 10 composite samples that were selected to represent the three styles of mineralization that have been identified (sulphide, alloy and mixed). The nickel grade and deportment of these samples are presented in Table

100 Table 16.1 Phase 2 Metallurgical Testwork Composite Samples Composite Type Ni (%) Hz (5) Pn (%) Aw (%) Si (%) Comp 02 Sulphide Comp 12 Mixed Comp 18 Sulphide Comp 22A Alloy Comp 23B Sulphide Comp 08 Sulphide Comp 93 Sulphide Comp 131 Mixed Comp 130 up Sulphide Comp 130 low Mixed Notes: Hz - heazlewoodite, Pn - pentlandite, Aw - awaruite, Si - silicates. The metallurgical testing performed during the Phase 2 program essentially comprised the flowing: De-fibering Wet grinding and de-sliming Recovery of a ferro-nickel concentrate using magnetic separation Recovery of a nickel sulphide concentrate using flotation De-fibering CTMP performed dry crushing, screening and air classification tests on composites 02 and 12. The test regime was based on processes used by primary chrysotile producers in Quebec. These tests showed that, with a size reduction to 840 μm, liberation of chrysotile fibres from granular serpentine was mostly complete. With a simple air classification, it was possible to remove a chrysotile product that was depleted of nickel. The intensity of the air classification determined the amount of chrysotile extracted and associated nickel losses. From the samples tested, it was determined that a mass pull of 10% would result in rejection of 80% of chrysotile, with a loss of 7.0% 8.5% of contained nickel Wet Grinding and De-sliming Wet testing of the air classification concentrate (i.e., following removal of chrysotile) revealed that the viscosity of pulp was very high. The high viscosity, which interfered with grinding and flotation, was found to result from the presence of brucite. CTMP successfully devised a process for stage grinding and concurrent de-sliming using hydrocyclones. The initial stage of grinding achieved a P 80 of 150 μm before the initial stage of de-sliming. The thick, paste-like hydrocyclone overflow was discarded, while the coarse, free-flowing underflow was treated with a low-intensity magnetic separator (LIMS) to produce a magnetic product (containing alloy mineralization). 88

101 Tailings from the initial magnetic separation were reground to a P 80 of 74 μm before a second stage of de-sliming. The hydrocyclone overflow was again discarded, while the underflow was treated on the magnetic separator to scavenge any remaining alloy mineralization liberated in the second-stage grinding. The process was optimized with a mass pull of 5% at each of the two stages of de-sliming, with an associated loss of 5% of the contained nickel at each stage. This level of brucite rejection produced a consistently low viscosity pulp suitable for flotation. With a higher mass pull to the hydrocyclone overflow, no improvements in pulp viscosity were observed, but nickel losses increased significantly. Based on the metallurgical work to date, a single stage of grinding was selected for the scoping study. Testwork results suggested that satisfactory recovery to the magnetic concentrate could be achieved after a single stage of grinding Recovery of Ferro-nickel Concentrate The magnetic separation process described above results in a rougher concentrate that typically grades approximately 1% to 3% Ni and 45% to 60% Fe. As both the primary sulphide minerals (pentlandite and heazlewoodite) are weakly magnetic, the rougher concentrate includes a significant amount of nickel in sulphide minerals. The combined sulphide-alloy composition would be acceptable if concentrate were treated conventionally (by smelting). However, it is intended to produce a clean ferro-nickel concentrate that would be suitable as direct feed for stainless steel producers. Preliminary test work has shown that rougher concentrate can be cleaned using flotation. The initial stage of flotation would depress alloy mineralization and produce a sulphide concentrate that would be pumped to the nickel sulphide circuit. A subsequent stage of flotation would then separate ferro-nickel concentrate from gangue. Cleaning losses (5%) and the ultimate grade of cleaner ferro-nickel concentrate (nominally 25%) have been established based on a limited amount of tests. Further work during the next stage of project development will be performed to better quantify these values Testwork Results The metallurgical performance for the three styles of mineralization (sulphide, alloy and mixed) using a common flowsheet was estimated by Mineral Solutions using the results from the Phase 2 metallurgical testwork program. These estimates, based on the standard simplified flowsheet presented in Figure 16.1, are summarized in Table

102 Figure 16.1 Simplified Metallurgical Flow Diagram Defibering Dynamic Classification Grinding and De-sliming Stage 1 Mag. Sep. Rougher Flash Float Flotation Conc. Fluff Float Tailings Slime Float Tailings Mag. Sep. Cleaner and Re grind Tailings Flotation Conc. Flotation Conc. Final Mag. Conc. Grinding and De-sliming Stage 2 Tailings Rougher and Scavenger Flotation Cleaner Flotation Flotation Conc. Tailings Tailings Table 16.2 Average Composite Sample Parameters and Estimated Nickel Recoveries Description Units Style of Mineralization Sulphide Alloy Mixed Number of test results Average Sample Grades Average Ni Deportment Ni Recovery to Rougher Concentrate Minimum % Ni Maximum % Ni Average % Ni Cut-off % Ni Ni in Pentlandite % of Total Ni Ni in Heazlewoodite % of Total Ni Ni in Awaruite % of Total Ni Ni in Silicates % of Total Ni Minimum 1 % of Contained Ni Maximum 1 % of Contained Ni Average 1 % of Contained Ni Adjusted Recovery 2 % of Contained Ni Recovery to rougher concentrate, excludes contribution from fibre and slimes scavenger circuits. 2 Adjustment includes contribution from fibre and slimes scavenger circuits which increases overall recovery to rougher concentrate by a minimum of 6% (alloy and mixed mineralization) to a maximum of 8% (sulphide mineralization). These estimates are based on recovering approximately 50% of the contained Ni reporting to these scavenging circuits. 90

103 16.4 RECOVERY EQUATIONS Recovery equations that are used in the PA were developed by David Penswick on behalf of Royal Nickel. These equations used the available testwork results on composite samples for which data on mineral deportment was available. Table 16.3 summarizes results of the metallurgical tests completed on the 32 composite samples. This table includes the grades of various economic nickel-bearing minerals available; including awaruite (Aw), pentlandite (Pn) and heazelwoodite (Hz). The magnetite (Mag) content was also estimated for each sample. The domain sampling program used to prepare the composites identified in Table 16.3 was conducted for the purpose of providing samples representative of the metallurgical response variability of the Dumont deposit. Sampling was designed to provide full representation of material within the mineralized envelope, down to the 0.20% nickel cut-off grade. Samples were collected in drill holes distributed to be spatially representative both along strike, and across dip (stratigraphy) of the deposit (Figure 16.2). These holes were drilled solely for the purpose of metallurgical sampling and are twins of previously drilled holes, except in the case of 09-RN-197 which was also a resource definition hole. Continuous domain samples were composited along the length of the holes as shown in Figure Each of the composites defined a homogeneous domain as characterized by nickel grade, nickel deportment, mineralization grain size and alteration. Figure 16.2 Locations of Domain Composite Samples Figure supplied by Royal Nickel Corporation. 91

104 Figure 16.3 Cross-Section Showing a Typical Downhole Distribution of Domain Composite Samples 0 50 Metres Samples RNC-184A to RNC-184H from Drill Hole 09-RN-184. Viewing Direction is Northwest (Azimuth = 315) Figure supplied by Royal Nickel Corporation. Analysis of the metallurgical results (Table 16.3) revealed that composites prepared from hole 176 yielded anomalously low results which was determined to result from the unique mineral assemblages for the domain in which this hole was located. Consequently, results from this hole were used to predict sulphide metallurgical performance for Domain 3 only, while the remaining samples were used to predict sulphide performance for all domains. Also, only results from samples that met criteria for alloy ore were used in deriving equations for magnetic recovery, resulting in seven samples being used to predict magnetic recovery for Domain 3 and an additional three samples (for a total of 10) used to predict magnetic recovery for the remaining seven domains. It is noted that these results exclude nickel recovery from scavenging the fibre and brucite slimes tails. Regression analysis was used to develop the nickel recovery equations. Each equation was applied to the entire modeled resource for domains 1 through 7 and an average recovery was calculated for each of the three types of mineralization (sulphide, alloy, and mixed). These calculated recoveries were compared to the average recoveries calculated by Mineral Solutions by using results from repeated flowsheet testing of the three types of mineralization (see Figure 16.1 and Table 16.2). The equations selected for use in the study were those that resulted in the lowest variance between the calculated and idealized recoveries. These equations then had adjustment factors applied to cater for the variance between calculated results and the average recoveries estimated by Mineral Solutions. 92

105 Table 16.3 Metallurgical Testwork Results From Domain Composite Samples Sample Grade of Minerals (%) Recovery (%) Notes Hole Composite Hz Pn Aw AzPnAw Mag Ni Mag. Sulphide Total Sulphide Eqn. Magnetic Eqn A Zone 3 only Not used B Zone 3 only Not used C Zone 3 only All zones D Zone 3 only Not used E Zone 3 only Not used F Zone 3 only Not used G Zone 3 only Not used H Zone 3 only All zones A All zones Not used B All zones Not used C All zones Not used D All zones All zones A All zones Not used B All zones Not used C All zones Not used D All zones Not used E All zones Not used F All zones Not used G All zones Not used H All zones Excludes zone A All zones Not used B All zones Not used C All zones Not used D All zones Not used A All zones Excludes zone B All zones Not used C All zones Not used D All zones All zones E All zones Not used F All zones All zones G All zones All zones H All zones Excludes zone 3 These selected equations are as follows: Domain 3 Only: Recovery to Rougher Sulphide Concentrate: o Sulphide (Type 1) Recovery = ln % (Pn + Hz) in ore o Alloy (Type 2) Recovery = ln % (Aw + Pn + Hz) in ore o Mixed (Type 3) Recovery = ln % (Aw + Pn + Hz) in ore Recovery to Rougher Magnetic Concentrate: o Sulphide Recovery = x % magnetite in ore o Alloy Recovery = x % magnetite in ore o Mixed Recovery = x % magnetite in ore 93

106 Domain 1-2 and 4-8: Recovery to Rougher Sulphide Concentrate: o Sulphide Recovery = ln % (Pn + Hz) in ore o Alloy Recovery = ln % (Aw + Pn + Hz) in ore o Mixed Recovery = ln % (Aw + Pn + Hz) in ore Recovery to Rougher Magnetic Concentrate: o Sulphide Recovery = x % magnetite in ore o Alloy Recovery = x % magnetite in ore o Mixed Recovery = x % magnetite in ore The equations also include a provision for additional recovery of nickel from the fibre and slimes removal scavenger circuits. The additional Ni recovery attributable to these scavenger circuits comprised 8% for sulphide mineralization and 6% for alloy and mixed material. These estimates are based on preliminary open circuit flotation testwork and will need to be confirmed by more detailed testing during subsequent phases of process development. In interpreting the metallurgical results the following flotation and magnetic separation cleaner parameters were assumed: Average flotation cleaner concentrate grade 30% Ni. Average magnetic cleaner concentrate grade 25% Ni. Losses in cleaning both concentrate 5% of the Ni contained in the rougher concentrate. These parameters are conceptual, based on the preliminary testwork undertaken so far. They will need to be confirmed by a more rigorous series of locked cycle or continuous tests CONCLUSIONS Since 2008, significant progress was made in improving the metallurgical response of the Dumont mineralization. The initial phase of this work ( ) consisted of laboratory testwork and mineralogy investigations to identify the various characteristics of the mineralized material that affected the metallurgical recovery. The testwork used composite samples that represented the deposit mineralization to focus on maximizing nickel recovery. Unit operations considered included chrysotile defibering, brucite slimes removal, magnetic separation and flotation. This work demonstrated that high nickel concentrate grades can be obtained following cleaning. Based on this optimization, a standard laboratory flowsheet (rougher stages only) was developed to test the three main mineralization types (sulphide, alloy and mixed) in the various domains within the Dumont Property (2009 onward). This procedure allowed the 94

107 characterization of variability of the mineralization. The results from this test program (which evaluated 32 samples, over the three mineralization types) were used to form the basis of the nickel recovery equations used in the preliminary assessment. The average life of mine metallurgical nickel recovery included in the preliminary assessment is 65.5%, comprising 9.9% into a magnetic product and 55.4% into a flotation concentrate. The results from the testwork completed to date were used as a basis for the engineering design and associated costing used for the preliminary assessment. Micon notes that the metallurgical recoveries and design criteria are conceptual in nature and although they are reasonable for use in a scoping study or preliminary assessment, more detailed work will be required to confirm the process design and metallurgical performance as the development of the project moves forward into pre-feasibility and feasibility study level of detail. Richard Gowans P.Eng., President and Principal Metallurgist of Micon International Limited, has reviewed the metallurgical testwork, the interpretation of the testwork results and the procedures used to predict metallurgical recoveries FUTURE WORK In February, 2010, Royal Nickel commissioned Minerals Associates Inc. (Minerals Associates) to design and construct a continuous mini pilot plant (MPP) on the Phase 1 preliminary assessment - Barmac Option at a throughput of kg per hour. This plant was completed in July, 2010 and testing of samples commenced in August, The MPP testwork is being performed to confirm the laboratory metallurgical performance (recoveries, concentrate grades and reagent dosages) for the various mineralization types. Flowsheet optimization work will further investigate both sulphide and magnetic cleaning circuits (concentrate grade and recovery), and optimize reagent and energy costs. A trade-off study to evaluate alternative primary grinding options will also be completed. 95

108 17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES The preliminary assessment presented in this report is based on the mineral resource estimate prepared by Micon in August, 2010 (Lewis and San Martin, August, 2010). Royal Nickel s August, 2010 mineral resource estimate is based on its 2010 exploration program and this estimate, which has been reviewed and audited by Micon, supersedes the previous resource estimate which was contained in the April 5, 2010 Technical Report. With the completion of the 2010 drilling program, Royal Nickel has obtained sufficient data to undertake a mineral resource estimation with a higher degree of confidence, enabling it to report measured resources over a portion of the deposit and to eliminate the use of any historical drilling results. The August, 2010 mineral resource estimate complies with the CIM standards and definitions for mineral resources, as required by NI The following is reproduced from Section 17 of Lewis and San Martin, August, The mineral resource estimate discussed herein is based on the current property database which contains a total of 70,577 m of assay results from 223 drill holes that Royal Nickel has obtained through its 2007 to 2010 drilling programs. The total metreage for the 223 holes available for the resource estimate is 90,212 m. Micon estimated the updated mineral resource based on the geological information and assaying data for the Dumont Property available as of April 22, The effective date of the resource estimate is August 16, For previous Technical Reports, insufficient techno-economic data pertaining to the mining and processing of Dumont mineralization and marketing of resulting concentrates were available to allow a cut-off grade to be estimated quantitatively. The cut-off grade of 0.25% Ni used in previous reports was estimated qualitatively, based on the cost structure and metallurgical performance of similar open pit bulk-tonnage, low-grade operations. In previous reports, only nickel grade and specific gravity were interpolated in the resource block model. For this Technical Report, it has been possible to refine the estimated cut-off grade based on work from the ongoing preliminary assessment. This study was initiated in the first quarter of 2010 and is scheduled for completion by the end of the third quarter. The operational cost structure being outlined in the preliminary assessment, indicates that considerable realizable value is contained in some material with grades below 0.25% nickel, where a high proportion of the nickel is contained in recoverable minerals (pentlandite, heazlewoodite and awaruite). Recognizing that the abundance of nickel in recoverable minerals is of paramount importance to mine planning and plant design, Royal Nickel retained Golder to prepare a resource block model that would incorporate nickel grade and major mineralogical abundances. The resource block model work was completed by Olivier Tavchandjian, P.Geo., and was peer reviewed by Greg Greenough, P.Geo., both of Golder. The current mineral resource estimate will form the basis for the preliminary assessment. 96

109 17.1 MICON UPDATED 2010 RESOURCE ESTIMATE FOR THE DUMONT PROPERTY Micon has audited Royal Nickel s current resource estimate for the Dumont Property using structural, geological and assay information from a total of 223 drill holes totalling 90,212 m, drilled between 2007 and The mineral resource has been categorized into measured, indicated and inferred categories which are compliant with CIM guidelines as required by NI regulations Modelling Methodology Royal Nickel has used the structural (fault) model developed by Itasca Consulting, in conjunction with the definition of geological contacts and grade distribution defined by drilling, to construct several mineralized envelopes corresponding to structural domains. These mineralized envelopes have been modified from those used in the previous resource estimation by expanding them slightly to accommodate the decrease in reporting cut-off grade. These envelopes have been used to spatially constrain the resource block model. The current resource block model interpolates nickel, copper, cobalt, chromium, platinum, palladium and gold grades, specific gravity, and ten factor scores used to calculate the mineral abundances of pentlandite, heazlewoodite, awaruite, olivine, magnetite, serpentine, brucite and coalingite. The abundances of individual minerals have been estimated over the entire deposit through mineralogical modelling completed by Golder, based on the correlation between mineralogical measurements and geochemical and assay data Structural and Geological Modelling Methodology Fault Model Royal Nickel identified a series of faults which appear to cross-cut the Dumont deposit, locally offsetting the mineralized zone and potentially affecting the distribution of the mineral resource. It was therefore decided to engage Itasca Consulting to collaborate on the development of a detailed representation of the faults within the Dumont deposit for use in the resource model. The Itasca Consulting work plan included the following: Review available data and interpretations and develop a strategy for defining the faults with the available mapping, drilling and geophysical data. Refine the lineament interpretation and analysis for the first Vertical Derivative Magnetic shaded relief model and LIDAR Bare Earth model. Attribute the final data set with the source of the interpreted lineaments and correspondence between LIDAR and magnetics. Press all interpreted lineaments to topographic elevation in order that they can be used in the fault model. 97

110 Ground truth lineaments to obtain fault plane information by mapping structures on a select set of existing outcrops, and other potential outcrops within the ultramafic defined from the LIDAR data. This assists in understanding the dip for projecting some of the faults in the 3D model. Mapping the orientation of the net slip lines helps to define the displacement magnitude and shear sense of the faults. Assemble drill hole structural information and lineaments in Datamine to construct a 3D fault model, using plan and section analysis on the drill-intercepted faults and the interpreted lineaments to define the orientation, distribution, connectivity and persistence of the faults. Itasca Consulting defined a total of 18 interpreted faults in its model of the deposit, but only six major faults with significant displacements were used to divide the Dumont deposit into seven separate structural domains for the current resource model, compared to two domains used in the resource models prior to April, Mineralized Envelope Golder and Royal Nickel conducted all of the 3D modelling work primarily using Datamine Studio but also GEMS 6.2 software. Micon has verified and audited the mineralization envelopes using primarily Surpac software. Royal Nickel provided Golder with the mineralized resource envelope described in the previous Technical Report dated April, Royal Nickel and Golder reviewed the previous work and, in light of the new 2010 resource drilling, decided to update the mineralized resource solid. For consistency in the modelling of additional parameters in the geometallurgical work, Golder recommended a constant geological hangingwall contact. The resource solid was updated so that every hangingwall contact honoured the geological dunite contact. Additionally domains D3a, D3b and D3c are now combined into one continuous D3 domain, because of the expansion of D3 toward the footwall. This expansion is a result of the 2010 sectional drilling program described previously in Section 11. The finalized new solid shapes were returned to Royal Nickel for discussion and approval prior to proceeding with the new interpretation. The updated information for the seven domains is shown in Table The overburden surface was constructed using the drill hole data. Based on all of the data currently available, seven separate solids were generated. The seven solids do not overlap each other in space, but all are contiguous and have been constrained on the basis of a 0.2% nickel reporting cut-off grade. The seven solids were constructed on the basis of the available structural model and the increased confidence level in the data set. 98

111 Table 17.1 List of Interpreted Structural Domains for the Dumont Project Domain Bearing Dip Length Volume Percentage Description Name (º) (º) (m) (m 3 ) (%) D1 269º/296º -70º Southeast end domain 1,200 31,930, D2 288º -58º ,711, D3 302º -68º 1, ,158, D4 308º/346º -58º ,936, D5 311º -60º ,434, D6 305º -60º ,578, D7 308º -60º Northwest end domain 2, ,759, Totals 7, ,508, Table supplied by Royal Nickel Corporation. Along the strike direction, the current resource model extends between sections 3600E and 10000E and, due to the differing strike directions of the seven domains, the total length is 7,035 m. The vertical boundaries are defined using the overburden and rock interface as the upper boundary, while the lower boundary is defined by using a variable projected distance of approximately 50 m below the deepest drilling assays above the cut-off grade. The hangingwall and footwall boundaries are projected in the down dip direction (average of - 58º) as defined by the actual assays above the cut-off criterion. Figure 17.1 illustrates the ranges of data used in the modelling for the current mineral resource estimate at the Dumont property. Appendix 6 [see Lewis and San Martin, August, 2010] contains examples of plans and vertical sections showing the block model nickel percentages and drill holes, with nickel percentage histograms the block model and drill holes. The seven 3D resource solids were generated using Datamine Studio software and based on the data from 223 drill holes totalling 90,212 m Geometallurgical (Mineralogical) Modelling Methodology Resource models typically consist of tonnes above cut-off, grade above cut-off and the 3D distribution of tonnes/grade above cut-off for one or more variables of economic interest. Metallurgical recovery, dilution and mineability are modifying factors that ultimately influence profitability. Geometallurgy is an emerging field targeted at integrating these issues by identifying either direct measures or proxies for throughput and recovery from easily collected microscopic data (e.g., Dunham and Vann 2007). By integrating geology, mining operations, mineral processing and metallurgy, geometallurgy aims to improve the fundamental understanding of resource economics. The basis for geometallurgical value modelling is an accurate mineralogical model of the deposit. 99

112 Figure 17.1 Magnetic Intensity Map of the Dumont Project with Interpreted Lineaments Figure supplied by Royal Nickel Corporation. Royal Nickel s metallurgical process development work to date (Section 16) has indicated that nickel deportment (mineralogical mode of occurrence), as well as the abundance of magnetite, are of primary importance to recovered value and consequently to mine planning and plant design. Royal Nickel decided to engage Golder to collaborate on the development of a detailed mineralogical model to quantify and identify the various attributes that contribute to the realizable value of the Dumont resource. The mineralogical modelling work was completed by Olivier Tavchandjian, P.Geo., and was peer reviewed by Greg Greenough, P.Geo., both of Golder. Royal Nickel has used 189 mineralogical mapping samples and their known assay results to develop a predictive model that will estimate nickel deportment in the primary nickel bearing phases: pentlandite, heazlewoodite, awaruite and silicates. This allows the individual blocks in the 3D model to be classified by mineralization type (sulphide, alloy, mixed) as described in Section 9. By creating a mineralogical model, Royal Nickel is now able to estimate metallurgical recoveries based on nickel deportment and metallurgical testwork, as opposed to using nickel grade alone. The improved mineralization type characterization, combined with the spatial modelling of critical physical characteristics that forms the basis of a geometallurgical approach, provides a much improved basis for operational and mineral processing plant design. 100

113 Golder s modelling procedure included the following: High-level data validation of the three main databases and merging of information where more than one type of data was available. The three databases used to develop the mineralogical model are: o Drill hole Geochemical / Assay database: 70,577 m of assay results from 223 drill holes. o Mineralogical Mapping Database (EXPLOMIN TM Qemscan: 189 samples in 27 holes). o Mineralogical Domain Composite Characterization Samples (Qemscan mineralogy on sized samples from 32 metallurgical domain composite samples). Cluster analysis of systematic down-hole measurements of 43 variables and 8 mineralogical abundances in order to reduce the number of meaningful variables. Factor analysis on the selected variables in order to obtain factor scores for the 10 most significant linear combinations of these variables. Multi-linear regressions between the factor scores, the mineralogical abundances as obtained from EXPLOMIN TM results, and between factor scores and mineralogical associations and fraction size distribution from the full Qemscan results (Appendix 7 describes this procedure with an example [see Lewis and San Martin, August, 2010]). Selection and length compositing of factor score and assay data within the seven structural domains. Variogram analysis of factor score and assay composited data for the seven domains. Interpolation (Ordinary Kriging) of nickel, copper, cobalt, chromium, platinum, palladium and gold grades, specific gravity and ten factor scores in 20 m x 20 m x 15 m blocks in the seven domains (plus extrapolation in a potential extension zone down to the 700 m elevation to investigate the value of extending the exploration drilling to a depth of 750 m below surface). Application of multi-linear regressions to reconstitute, in each block, all the mineralogical abundances and associations required for the application of the recovery equations in each structural domain (Appendix 7 [see Lewis and San Martin, August, 2010] describes this procedure with an example). Visual and statistical validation of the models, and smoothing assessment and correction when necessary, for the nickel grade and nickel bearing mineral abundance estimates for each domain. Calculation of nickel deportment in pentlandite, heazlewoodite, awaruite and silicates and mineralization type classification using Royal Nickel s algorithms, described in Section

114 The final mineralogical model of the Dumont project contains all the chemical, physical and mineralogical parameters required to estimate an in-situ recoverable value, when combined with metallurgical recoveries for the various mineralization types as determined by the metallurgical testwork program (as described in Section 16). The resultant recovery model will be used to conduct an open pit optimization for the purpose of the ongoing preliminary assessment. Royal Nickel will continue testing, calibrating and refining the geometallurgical model as more mineralogical mapping samples (EXPLOMIN TM ) become available Resource Block Modelling Methodology The resource solids indicate that the mineralization strikes in a northwest direction at about 315 (Figure 17.1). A block model was created to cover the volume of each structural domain using a common origin point to allow combination of the seven models. The block size was set to 20 m (UTM Easting) x 20 m (UTM Northing ) x 15 m (vertical) in all areas. The block model was not rotated as in the previous Technical Reports. The 20 m x 20 m x 15 m block size was selected as a reasonable approximation of the selective mining unit (SMU) that will be achieved during mining. A volume check of the block model versus the solids used to generate it revealed a very good representation of the volume overall for each area (Table 17.2) Compositing Data Table 17.2 Block Model versus Solid Volume Check Domain Volume (m 3 ) Block model Wireframe D1 31,930,899 31,930,814 D2 34,711,838 34,711,772 D3 151,158, ,158,466 D4 100,935, ,936,096 D5 146,434, ,434,145 D6 63,578,069 63,578,037 D7 199,758, ,759,274 Data supplied by Golder and Associates Ltd. (Warren, 2010). All of the current Royal Nickel raw assay data were selected and captured within the resource solid and then composited using 5.0 m intervals. The vast majority of the samples were originally assayed over a 1.5 m length in all domains. However, the block size was set to 20 m x 20 m x 15 m in order to accommodate the anticipated level of mining selectivity for an open pit operated with large excavating equipment. As a result, a composite length of 5 m was chosen as more appropriate. Compositing was conducted from the top to the bottom of the profile and within the mineral resource solid. Since specific gravity (SG) does not present a significant correlation with chemistry, it was not deemed necessary to weight by density during the compositing process. A verification was conducted to ensure that the total 102

115 sample length before and after compositing remained approximately the same within the mineral envelope. Figure 17.2 illustrates the drilling information used for the mineral resource model. Figure 17.2 Drilling Information used for the Mineral Resource Model Figure supplied by Royal Nickel Corporation. The current Royal Nickel database contains some core intervals from which no samples were taken. A total of 363 intervals, representing m of core were not sampled but lie within the current resource solid limits. The non-sampled core represents 1.5% of the total 53,689 m of sampled core within the resource solid constrains. Table 17.3 summarizes the nonsampled core by domain. Table 17.3 Missing Assays by Structural Domain within the Resource Wireframe Domain Metres of Core Not Sampled Metres of Core Sampled % Not Sampled D D , D , D , D , D , D ,

116 TOTAL , Table supplied by Royal Nickel Corporation. In past reports, the intervals that were not sampled were assigned the lower of the background or the average values from samples to fill non-sampled core intervals. The background values were calculated statistically as the 25% percentile of the population for each domain. Although the domains have a fairly homogeneous nickel grade, Royal Nickel has decided to review the non-sampled intervals and may resample depending on their significance to the deposit model. For the current resource update, the non-sampled core has been treated as absent data. Micon considers this to be appropriate, given the change to the resource estimation methodology and the fact that non-sampled core represents only 1.5% of the total core within the resource solid. Micon recommends that the review of the nonsampled intervals and any potential re-sampling be conducted prior to any further resource update, if the review deems that re-sampling of the core is necessary Rock Density Royal Nickel has specific gravity (SG) determined on all of its pulp samples by pyncnometer. A total of 41,609 samples with SG values were located in the database provided by Royal Nickel, within the current resource solid limits. Where no SG data were available for the samples, the interval has been treated as absent data. The statistical SG results are summarized in the domain statistics provided in Table 17.4 and Table Figure 17.3 through Figure 17.9 are plots of nickel grades versus SG. They demonstrate that the SG is not proportional to the nickel grade, which might have been expected. At Dumont, lower nickel grade samples tend to have higher SGs and vice versa; SG is primarily correlated with degree of serpentinization. The average SG value in the domains ranges from 2.55 to 2.59, and the SG values for most of the samples with higher nickel grades (>0.25% nickel) are lower than the average value. This SG feature at the Dumont Property tends to support the proposed genesis hypothesis which states that the nickel mineralization is genetically related to serpentinization alteration, where the introduction of water during metamorphism results in the hydrolysis of the dunite to form serpentinite, with a corresponding increase in volume of up to 40% and a concomitant decrease of SG. During the serpentinization process, olivine is destroyed and the nickel contained in olivine is available for partition into pre-existing magmatic sulphides and newly formed iron-nickel alloy phases, resulting in enriched nickel grades. 104

117 Figure 17.3 Plot of SG versus Nickel Grade for Domain 1 at the Dumont Property Figure 17.4 Plot of SG versus Nickel Grade for Domain 2 at the Dumont Property 105

118 Figure 17.5 Plot of SG versus Nickel Grade for Domain 3 at the Dumont Property Figure 17.6 Plot of SG versus Nickel Grade for Domain 4 at the Dumont Property 106

119 Figure 17.7 Plot of SG versus Nickel Grade for Domain 5 at the Dumont Property Figure 17.8 Plot of SG versus Nickel Grade for Domain 6 at the Dumont Property 107

120 Figure 17.9 Plot of SG versus Nickel Grade for Domain 7 at the Dumont Property Assay Sample Nickel Grade Statistics The statistical results for the available assays (raw data) for the resource solids, within the 0.2% nickel envelope cut-off grade, were compiled and are summarized in Table 17.4 and Table All samples were composited at equal length intervals of 5.0 m. There is little difference in the statistical results among the raw data set, and the selected data set inside the solids. This is not unexpected for the Dumont Property, since the nickel mineralization is almost homogeneously distributed throughout the ultramafic sill. Table 17.4 Statistics for Results for Nickel Raw Samples from All Domains 1 (D1) to 7 (D7) Description Nickel Nickel Nickel Nickel Nickel Nickel Nickel D1 D2 D3 D4 D5 D6 D7 Number of samples 440 1,295 15,886 8,379 9,792 3,009 2,830 Minimum value Maximum value Mean Median Geometric mean Variance Standard deviation Coefficient of variation

121 Table 17.5 Statistics for Results for Density Raw Samples from All Domains Domain 1 (D1) to 7 (D7) Description DensityD1 Density Density Density Density Density Density D2 D3 D4 D5 D6 D7 Number of samples 440 1,295 15,875 8,368 9,792 3,009 2,830 Minimum value Maximum value Mean Median Geometric mean Variance Standard deviation Coefficient of variation Variography A variogram analysis was conducted on nickel, copper, cobalt, chromium, platinum, palladium and gold grades, specific gravity, and the ten factor scores selected for interpolation, in order to characterize their spatial variability and to establish optimum search strategies in the D3, D4 and D5 structural domain data sets. The variogram analysis was conducted in the XYZ Cartesian coordinate system separately for each one of the seven structural domains. The parameters used in the calculation of the experimental variograms are summarized in Table Table 17.6 Experimental Variogram Parameters Parameters/Domain D1 D2 D3 D4 D5 D6 D7 Lag distance Number of lags Sub-lag distance Number lags to be sub-lagged X-axis N100/0 N115/0 N125/0 N135/0 N135/0 N135/0 N135/0 Y-axis N10/75 N25/75 N35/65 N45/65 N45/65 N45/65 N45/65 Z-axis N190/15 N205/15 N215/25 N225/25 N225/25 N225/25 N225/25 Angular tolerance in horizontal plane 30 o 30 o 30 o 30 o 30 o 30 o 30 o Angular tolerance in vertical plane 30 o 30 o 30 o 30 o 30 o 30 o 30 o Cylindrical search radius 10 o 10 o 10 o 10 o 10 o 10 o 10 o Table supplied by Golder Associates Ltd. (Warren, 2010). The angle rotations were adjusted to the average strike and dip of each domain. The D1, D2, D6 and D7 domains individually do not contain sufficient data to support a variogram analysis. For D1 and D2, D3 was used as the best analog, based on discussions with Royal Nickel geologists and was used D5 was used as the best available analog for D6 and D7. The variogram models adjusted to the experimental values are presented, as an example, for percent nickel in Table 17.7, in order to illustrate the similarities between the seven domains. The full suite of models for all 18 variables and for all domains as provided by Golder is contained in Appendix 8 [see Lewis and San Martin, August, 2010]. Variogram fitting was adjusted after filtering on a minimum number of pairs of 30. Pair-wise relative variograms were used for most chemical elements, as their frequency distribution is usually highly skewed. Normal variograms were used for factor scores, as they have histograms close to a 109

122 normal distribution. In most variogram models, a zonal anisotropy was identified and modelled as a third spherical structure in one direction only, usually in the Z direction, after rotation reflecting the alteration zoning. Table 17.7 Variogram Models Used for Percent Nickel Parameters/Domain D1 D2 D3 D4 D5 D6 D7 Nugget sill range 1 X (m) range 1 Y (m) range 1 Z (m) sill range 2 X (m) range 2 Y (m) range 2 Z (m) sill range 3 X (m) 10,000 10,000 10,000 10,000 10,000 10,000 10,000 range 3 Y (m) 10,000 10,000 10,000 10,000 10,000 10,000 10,000 range 3 Z (m) Table supplied by Golder Associates Ltd.(Warren, 2010) Data Interpolation Strategy Inside the Resource Block Model Nickel, copper, cobalt, chromium, platinum, palladium and gold grades, specific gravity, and the ten factor scores were estimated using Ordinary Kriging (OK). Nearest Neighbour estimates of the same elements provided declustered sample grades for the purpose of validating the block model. Anisotropic searches were performed, using the factor score 1 variogram model as a guide, since this factor represents the largest portion of the overall variability. The search parameters used for each domain are similar and are summarized in Table Table 17.8 Search Parameters used for the Ordinary Kriging Interpretation Parameters 1st search 2nd search 3rd search Search radius in metres D1,D2, D3 80x60x25 x 2 x 2 D4 120x60x30 x 2 x 2 D5, D6, D7 100x60x30 x 2 x 2 Rotation See Table See Table See Table Minimum number of octants with samples Maximum number of samples per octant Minimum number of samples Maximum number of samples Table supplied by Golder Associates Ltd. (Warren, 2010). Since factor score 1 explains the largest portion of the overall variability, its variogram model was used to establish the search ellipsoid dimensions. The optimum search was set in the down-dip direction (Y axis) to 60 m. The corresponding radius of the search ellipsoid in the other two directions is approximately 120 m along strike (X) and 30 m across strike (which is also approximately the down-hole direction (Z)). As a result, the first search ellipsoid was set to 120 x 60 x 30 m after rotation. The minimum number of composites 110

123 found in the first search was set to 20 and the maximum to 32, adding octant constraints of at least 5 octants with information and no more than 4 composites per octant. The second search applies the same criteria, except that the search radius is doubled in all directions. For the third and final search, the search radius is again doubled, with a minimum of 1 composite with no octant constraints Model Validation and Post Processing Validation of the kriged block model consisted of: Visual validation of grade interpolation by a systematic walk-through of sections, comparing the actual drill hole data with the estimated block average values (Appendix 8) [see Lewis and San Martin, August, 2010]. Global statistical comparison for each variable of the mean and variances, in order to verify that the global mean of the kriged model was similar to a declustered mean obtained through a nearest neighbour interpolation, and that the global variance was lower than the sample variance. Smoothing assessment on the percent nickel estimates in order to confirm that the search strategy was based on enough samples and also to identify where a smoothing correction may be necessary. Results of the block model validation are summarized, from Golder, by domain in Appendix 9 [see Lewis and San Martin, August, 2010]. As expected, no global bias was identified and significant smoothing occurred in areas with sparser drilling. Where the percent nickel block variance was found to be materially lower than expected, an indirect lognormal correction (Isaak and Srivastava, 1993) was applied, in order to restore the block variance expected for 20 m x 20 m x 15 m blocks from the variogram model. Ordinary kriging can generate negative interpolation weights and, as a result, may produce negative grade values. This is normal and no attempt was made to correct the weights. Negative values were observed in some of the domains, mostly for g/t gold, as well as for ppm copper in some cases. The negative estimated grades were reset to 0 and verifications were made that the number of corrected values was small and did not have any material impact on the global statistics. From the factor score block estimates, the predictive equations for pentlandite, heazlewoodite, awaruite, olivine, magnetite, serpentine, brucite and coalingite mineral abundances were applied in each domain. After applying these multi-linear regressions, several variables showed a significant number of negative values. These negative values were reset to 0. This correction resulted in some cases in a significant change in the average mineral abundance, usually for pentlandite, heazlewoodite, awaruite and coalingite. In those instances, a top cut was selected for each one of these variables and for each domain, so that the global mineral abundance average would remain unchanged from the original estimates. 111

124 Appendix 10 [see Lewis and San Martin, August, 2010] from Golder summarizes these corrections. The variance of the estimated mineral abundance in each block was compared to the variance in the original mineralogical mapping samples. Although some smoothing is expected when averaging sample values into block estimates, it was found that the smoothing was largely enhanced by the linear regression process, especially for those mineral abundances with a lower correlation coefficient associated with the regressions. A smoothing correction was developed using an indirect log-normal correction for the main minerals of interest, i.e., pentlandite, heazlewoodite, awaruite and olivine. The magnitude of the correction was estimated using the following approach: Find the factor score presenting the highest coefficient of correlation with each mineral abundance. Estimate the percentage reduction from the most closely related factor score variogram in the variance normally expected when estimating 20 m x 20 m x 15 m blocks from 5 m samples. From the 1.5 m sample EXPLOMIN TM actual mineral abundance variances, use the expected variance reduction (adjusted for the difference in sample length) to estimate a theoretical block variance for pentlandite, heazlewoodite, awaruite and olivine in each domain. Apply an indirect log-normal correction to restore the expected variance in mineral abundances for 20 m x 20 m x 15 m blocks. This post-processing step is summarized and illustrated with two practical examples for the awaruite content in Domain 3, in Table Table 17.9 Example Corrections for Over-Smoothing in Awaruite Content for Two Representative Blocks in Domain 3 Smooth Estimate of Awaruite Mean of the Corrected Value (Using Indirect Lognormal Correction) Smoothing Ratio Abundance Deposit Table supplied by Golder Associates Ltd. (Warren, 2010). The last step in the post-processing of the model consisted of calculating the nickel deportment in each mineral using the interpolated total percent nickel content and the estimated abundances of the nickel-bearing minerals. The nickel content in each mineral was provided by Royal Nickel and set to 34.21% in pentlandite, 73.30% in heazlewoodite, 72.43% in awaruite and 0.35% in olivine. A total percent nickel content was estimated from these four minerals using these values. Because the percent nickel in head grade is estimated independently from the mineralogy, the reconstitution of the percent nickel grade from the abundance of these four minerals could exceed the total percent nickel grade. As a result, a 112

125 rescaling of the percent nickel grade in each mineral was conducted where the totals were not consistent. When the percent nickel content estimated from the mineralogy was lower than the percent nickel grade estimated from actual assays, the balance of the percent nickel content was assigned to percent nickel in serpentine. Finally, the percent nickel contained in olivine and serpentine was added as the total percent nickel in silicate minerals Future Work A sound approach has been developed to predict the mineralogy from chemical and physical measurements available from all of the samples of the drill hole database. Royal Nickel will continue the calibration and refinement of the predictive model of the mineralogy as more mineralogical mapping sample results becomes available. Micon has reviewed the process developed by Golder for the mineralogical block model and considers that this approach is applicable for the Dumont deposit. Micon recommends that Royal Nickel continue to use the model in future mineral resource updates as further information and results become available MINERAL RESOURCE CLASSIFICATION The Mineral Resources for the Dumont Property were estimated by Micon following the CIM guidelines. The following definitions were adopted for the categorization of mineral resources. Measured Mineral Resources are defined as those portions of the mineralized area which are drilled on a grid of 50 m by 50 m or less. The blocks were estimated based on detailed and reliable drilling information spaced closely enough to confirm grade continuity. Indicated Mineral Resources are defined as those portions of the mineralized area which are generally drilled on a grid of 100 m by 100 m or less. The blocks were estimated with a minimum of 4 samples. Inferred Mineral Resources are defined as those portions of the mineralized area which are drilled on a grid of 200 m by 200 m or less. While the geological continuity has been interpreted, there is not enough drilling to confirm the continuity of the grade within the inferred mineralization. In the earlier, April, 2010, Technical Report, Micon recommended that Royal Nickel complete a geostatistical analysis of the drilling results on 50 m, 100 m and 200 m centres in order to determine if there is any significant difference in results (Appendix 11 [see Lewis and San Martin, August, 2010]). The initial analysis indicates that the nickel mineralization has very good continuity and exhibits nearly homogeneous nickel grades. In addition, the nickel grades do not vary significantly when comparing 50 m, 100 m and 200 m drill spacing. Therefore, based on the initial analysis, a strong case could be made to move to 100 m spacing for measured mineral resources for the Dumont deposit. However, for the purposes of the current measured mineral resource estimate, the drill spacing continues to be 113

126 set at 50 m. Future work and independent verification of the geostatistical analysis will ultimately determine the final drill spacing at Dumont MINERAL RESOURCE ESTIMATES Economic Discussion and Cut-off Grade The CIM definitions of a Mineral Resource require that there are reasonable prospects for eventual economic extraction. For previous Technical Reports, insufficient technoeconomic data pertaining to the mining and processing of Dumont mineralization and marketing of resulting concentrates were available to allow a cut-off grade to be estimated quantitatively. Accordingly, the cut-off grade of 0.25% Ni was estimated qualitatively, based on the cost structure and metallurgical performance of similar open pit bulk-tonnage, lowgrade operations. For the August, 2010 report, it was possible to refine the estimated cut-off grade based on a set of development parameters for the Dumont deposit as derived in the study then being conducted by Royal Nickel. Key preliminary conclusions from the study were as follows: The primary control on recovery of nickel to concentrate is the nickel contained in recoverable minerals (pentlandite, heazlewoodite and awaruite) and not the overall nickel grade. Testwork has shown that 75% to 85% of nickel contained in these minerals can be recovered, and that these minerals represent 75% to 80% of total contained nickel. The remaining contained nickel is associated with silicate minerals that are currently largely unrecoverable. On a global basis, a recovery of approximately 65% of total contained nickel is forecast. However, as the ratio of recoverable minerals to silicate minerals is variable, recovery can vary from a minimum of less than 50% to over 85%. It will be possible to produce two different concentrates. The bulk of the nickel will be recovered to a sulphide concentrate that would be marketed to independent smelters. The unique mineralogy of this concentrate (predominantly pentlandite and heazlewoodite with no pyrrhotite) is expected to result in high concentrate grades of 20 to 40% nickel, compared to industry averages of 10 to 20%. The remaining nickel will be recovered to a magnetic ferro-nickel concentrate that contains primarily awaruite, a naturally occurring iron-nickel alloy. This concentrate is expected to grade between 20 to 30% nickel. Terms for smelting and refining the sulphide concentrate are expected to be comparable to competitive terms negotiated by other sulphide producers. Treatment charges, based on per tonne of concentrate treated, will be less on a per pound nickel basis, due to the higher concentrate grade. The ferro-nickel alloy concentrate is expected to pay no treatment or refining charges (TCRCs) as it would be used as direct feed for stainless steel making. Due to the combination of high grade sulphide concentrate and no TCRCs for ferro-nickel alloy concentrate, the weighted average TCRC (per pound nickel basis) for Dumont concentrate is expected to be 10% lower than that paid by producers of typical sulphide concentrate. 114

127 The size of the deposit would permit a high mill throughput, potentially in the range of 75,000 to 100,000 t/d. Economies of scale, coupled with the low cost of electricity in Quebec, are expected to offset the additional costs associated with production of two concentrates, and milling costs are expected to be in-line with industry averages. As is normal for open pit mines, the cut-off grade calculation shown in Table ignores fixed capital costs, as well as the cost of mining and transporting mineralization to the rim of the pit. The costs included in this calculation are: The full cost of re-handle for low-grade mineralization that would initially be stockpiled and fed to the processing plant at the end of mine life. The full cost of milling. The full cost for general and administration (G&A). Variable sustaining capital costs, including mill sustaining capital and the costs associated with increasing tailings dam capacity. Table Dumont Cut-off Grade Calculation Item Units Marginal Nickel Price US$/lb $7.50 TCRC US$/lb $1.20 NSR - pre Royalties US$/lb $6.30 NSR Royalties 1 US$/lb $0.09 NSR - post Royalties US$/lb $6.21 Mining Cost 2 US$/tonne ore $0.35 Milling Cost US$/tonne ore $7.25 G&A Cost US$/tonne ore $0.60 Sustaining Capital 3 US$/tonne ore $0.70 Total Cost $8.90 Payable Nickel 4 lbs/tonne ore 1.4 Ni Payables 5 % of recovered nickel 92.5% Minimum Concentrate Recovery % of contained nickel 40% Contained Nickel lbs/tonne ore 3.8 Head Grade Nickel 6 % nickel 0.17% Resource Grade Nickel % nickel 0.18% Notes: 1. Based on maximum royalty rate of 3.0%; assumes option to buy-down 50% is exercised (net rate = 1.5%). 2. Mining cost for re-handle of low-grade stockpiles. 3. Sustaining capital costs for mill replacements and tailings dam expansion, incurred on a per tonne basis. 4. Minimum payable nickel to cover total costs. 5. Weighted average payables for sulphide and ferro-nickel concentrates. 6. Head grade includes 2% dilution. Table supplied by Royal Nickel Corporation. 115

128 The cut-off calculation used in the preliminary assessment has been based on value (NSR/tonne) rather than grade of total contained nickel. However, the calculated cut-off grade of 0.18% shown in the above calculation is a reasonable proxy for the total inventory of potentially mineable material contained within a pit shell defined using the Lerchs- Grossmann algorithm. The preliminary assessment also applies an elevated cut-off in order to maximize net present value, based on a discount rate of 8 to 10%. The impact on resources of elevating the cut-off grade can be reasonably represented by increasing the calculated cut-off to 0.20% nickel. A cut-off of 0.20% nickel has thus been selected as the basis for the current resource estimate August, 2010 Mineral Resource Estimates Table summarizes the mineral resources at a 0.20% nickel cut-off grade. This table includes the measured, indicated and inferred resources tabulated by resource category and domain for the entire deposit and all domains (D1 to D7), based on the information obtained from the 2007 to 2010 drilling programs. Table summarizes the mineral resources for the core area of the deposit and is partly comparable to the area covered by the previously reported indicated resources and the 1971 historical resource and reserve estimate. A direct comparison with the current estimate is impossible for estimates prepared prior to April, Royal Nickel has conducted a more extensive drilling program outside the historical area which has added to the resource estimate, and the structural model has changed the interpretation used in the models prior to April, 2010, which envisioned a simpler single mineral domain. The tonnages and grades for the Measured, Indicated and Inferred Mineral Resource estimates are summarized Table The estimates in Table have been rounded to reflect the uncertainty inherent in resource estimations. Figure illustrates the location of the 7 domain solids involved in the mineral resource estimate. Figure illustrates the block model for the Dumont project, along with the drill hole and grade distributions above and below the 0.20% nickel cut-off grade. Figure illustrates the block model for the Dumont project, along with the distribution of the mineralization types. Figure illustrates the block model for the Dumont project, along with the distribution of the resource categories. Mineral resources that are not mineral reserves do not have demonstrated economic viability. There are no mineral reserves presently identified on the Dumont Property. The stated mineral resources are not materially affected by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political or other relevant issues, unless stated in this report, to the best knowledge of the authors. There are no known mining, metallurgical, infrastructure or other factors that materially affect this mineral resource estimate, at this time. 116

129 Table Summary of the Dumont Mineral Resources at a 0.20% Nickel Cut-off Grade in all Domain Solids by Category and Structural Domain: Cumulative Category Structural Domain Rock Code Cut-off Nickel % Cum. Tonnes Cum. Avg. Grade Nickel % Contained Nickel (Lbs) Measured Domain 3 3 > ,142, ,160,210 Measured Domain 5 5 > ,537, ,204,393 Total Measured 155,680, ,364,603 Indicated Domain 2 2 > ,860, ,418,521 Indicated Domain 3 3 > ,642, ,686,290,804 Indicated Domain 4 4 > ,057, ,337,016,524 Indicated Domain 5 5 > ,089, ,667,084,562 Indicated Domain 6 6 > ,613, ,960,560 Indicated Domain 7 7 > ,223, ,055,011 Total Indicated 1,003,486, ,966,825,982 Total Measured & Indicated 1,159,166, ,952,190, Inferred Domain 1 1 > ,882, ,747,687 Inferred Domain 2 2 > ,701, ,307,089 Inferred Domain 3 3 > ,917, ,339,023 Inferred Domain 4 4 >0.20 4,232, ,923,579 Inferred Domain 5 5 >0.20 5,920, ,167,295 Inferred Domain 6 6 > ,473, ,132,902 Inferred Domain 7 7 > ,276, ,237,602,714 Total Inferred 581,404, ,198,220,289

130 Table Summary of the Dumont Mineral Resources Between the 5500 and 9100 Section Lines at a 0.20% Nickel Cut-off Grade by Category and Structural Domain: Cumulative Category Structural Rock Cut-off Cum. Avg. Grade Cum. Tonnes Domain Code Nickel% Nickel% Measured Domain 3 3 > ,142, Measured Domain 5 5 > ,537, Total Measured 155,680, Indicated Domain 2 2 > ,860, Indicated Domain 3 3 > ,642, Indicated Domain 4 4 > ,057, Indicated Domain 5 5 > ,089, Indicated Domain 6 6 > ,613, Indicated Domain 7 7 > ,223, Total Indicated 1,003,486, Total Measure & Indicated 1,159,166, Inferred Domain 2 2 >0.20 7,762, Inferred Domain 3 3 > ,917, Inferred Domain 4 4 >0.20 4,232, Inferred Domain 5 5 >0.20 5,920, Inferred Domain 6 6 > ,473, Inferred Domain 7 7 > ,297, Total Inferred 87,604, Table Summary of the Measured, Indicated and Inferred Mineral Resource in the Seven Structural Domain Solids at a Cut-off of 0.20% Nickel (As of August 16, 2010) Area within Deposit Model All Structural Domains All Structural Domains Mineral Resource Category Tonnage Nickel Grade (%) Nickel (tonnes) Nickel (pounds) Measured 155,680, , ,365,000 Indicated 1,003,487, ,707,000 5,966,826,000 All Structural Domains Total Measured and Indicated 1,159,167, ,154,000 6,952,191,000 All Structural Domains Inferred 581,405, ,451,000 3,198,220,

131 Figure Location of the Seven Structural Domain Solids Involved in the Mineral Resource Estimate and their Structural Boundaries (Grey) 119 Figure supplied by Royal Nickel Corporation. Note that in Figure the mineralized envelopes are based on 0.20% nickel cut-off. These envelopes are used to constrain the resource block model.

132 Figure Block Model for the Dumont Project Illustrating the Grade Distribution along with the Drill Hole Distribution 120 Figure supplied by Royal Nickel Corporation.

133 Figure Block Model for the Dumont Project Illustrating the Distribution of the Mineralization Type with the Drill Hole Distribution 121 Figure supplied by Royal Nickel Corporation.

134 122 Figure Block Model for the Dumont Project Illustrating the Resource Categories with the Drill Hole Distribution

135 The Mineral Resource estimate as of August 16, 2010 is compliant with the current CIM standards and definitions required by NI regulations and is, therefore, reportable as a mineral resource by Royal Nickel DUMONT PROPERTY EXPLORATION POTENTIAL There is the potential to find a large amount of additional mineralization on the Dumont Property. This potential includes extensions to the current zones of mineralization in the resource model both at depth and on the northwest and southeast ends of structural domains 1 and 7. Exploration either at depth or on the flanks, may contribute further mineral resources to the Dumont Property. 123

136 18.0 OTHER RELEVANT DATA AND INFORMATION Royal Nickel has prepared a preliminary assessment, or scoping study, for the Dumont project. Participants in the study are listed in Table Table 18.1 Participants in Dumont Preliminary Assessment Activity Lead Organization Project Management A.St. Jean Royal Nickel Geology Exploration and Database A.St. Jean Royal Nickel Mineralogy S. Downing SGS Resource Model O. Tavchandjian Golder Resource Estimate W. Lewis Micon Geotechnical Slopes, Designs M. Garon Genivar Hydrology O. Fala Genivar Mining Concept Selection D. Penswick Independent LG Pit Design A.von Wielligh Prysm Resources Detailed Pit Design and Schedule R. Kear Independent Design Review M. Garon Genivar Processing Flowsheet Concepts and Design R. Salter Mineral Solutions Testwork J. Marois CTMP Engineering Design C. Hardie BBA Tailings Design, Concept R. Ouellet Golder Pumping P. Primeau Paste Tec Environmental Baseline Studies and Testwork D. Blanchet Genivar Water Balance C. Hardie BBA Closure Plan D. Penswick Independent Schedule and Execution Plan D. Penswick Independent Capital Cost Estimate Mining D. Penswick Independent Processing, Infrastructure M. Fitzgibbon BBA Tailings M. Lemieux Golder Operating Cost Estimate Mining D. Penswick Independent Processing C. Hardie BBA Tailings M. Lemieux Golder G&A D. Penswick Indpendent Evaluation and Report D. Penswick Independent The resource model used as the basis for evaluation of both the base case and upside case incorporates the results of drilling completed on or before April 22, 2010, and was documented in Micon s technical report dated August 30, 2010 (Lewis and San Martin, August, 2010). 124

137 Equations used to estimate the concentrator recovery of nickel and the design criteria used to estimate capital and operating costs are based on results from the initial 32 metallurgical domain composite characterization samples tested during Phase 2 of the metallurgical program. The suitability of these equations has been confirmed by the test results from a subsequent 22 samples GEOTECHNICAL WORK Genivar prepared two reports on geotechnical work carried out on the Dumont property: Preliminary Stability Analysis of Slopes (January 2010) Preliminary Slope Evaluation of the Overburden Dumont Project (April 2010) Open-Pit Rock Mechanics Feasible hard-rock slope angles for the open pit were estimated by Genivar, based on three geotechnical holes that were sited so as to intersect the various lithologies that would be encountered in the hanging wall and footwall of the pit. The geotechnical data gathered from these holes were analyzed by Genivar using RocLab Software, which calculates a factor of safety for a given slope angle. The conclusions arrived were that for the hanging wall rocks, an acceptable factor of safety (1.5) would be achieved with a 47 slope angle, while the more competent footwall rocks would allow the same factor of safety to be achieved using a 52 slope Open-Pit Soil Mechanics Overburden at Dumont is composed of a number of different lithologies, which were grouped by Genivar into the following two broad classifications based on similar geotechnical behaviour: Fine-grained material, including clay, silt and organic material. Coarse-grained material, including sand, gravel and boulders. The profile of fine and coarse material within the overall overburden horizon was determined by Genivar, based on logs from existing diamond drill holes. Genivar s analysis showed that fine material is generally confined to the uppermost 30% of the overburden profile. As the average depth of overburden is typically 30 m and the maximum depth seldom exceeds 45 m, for purposes of pit design it was assumed that the initial 15 m bench was composed of fine material and the remaining material to the bedrock contact was coarse material. The following feasible overburden slope angles for the open pit were estimated by Genivar, based on analysis of soil types at similar projects in the Dumont area. Fine-grained material slope angles 3H:1V (18 ). 125

138 Coarse-grained material 2H:1V (27 ) Waste Impoundments Overburden The base case pit shell contains Mt of overburden. Approximately 70% of this material is coarse grained, comprising sand, gravel and boulders. This material is expected to present no challenges to either excavation or impoundment. Investigations conducted to date have shown that the remaining fine-grained material, including clay and organic material, is not overly saturated. As a consequence, it is also assumed that this material will not present significant challenges to excavation and impoundment Waste Rock The base case pit shell contains 944 Mt of waste rock. Geochemical test work has shown that waste rock is not likely to be acid generating and will be a suitable material for construction. Approximately 20% of waste rock is planned to be utilized in construction of the tailings management facility (TMF). The remaining 760 Mt will be stored in two impoundments located to the south and east of the open pit. Both impoundments would be located a minimum of 150 m from the rim of the open pit, to ensure that there is no sloughing of waste into the pit. Impoundments will have slopes of 24 (2.25H:1V), composed of 15 m lifts with face angles equivalent to the angle of repose of approximately 34 (1.5H:1V) with 11 m catchment berms. The impoundments will reach a maximum height of 10 lifts, or 150 m Tailings Management Facility Life of project mill feed of 896 Mt will produce approximately 5 Mt of concentrate and 891 Mt of tailings. The current concept for the TMF includes mixing all three streams of tailings material (chrysotile fibres, brucite slimes and serpentine waste rock) into a single stream. Due to the impact of fibres and slimes, the density of this stream would be relatively low (45% by mass solids), and tailings will occupy a large volume. This has been mitigated by the strategy of accelerated mining, which will allow approximately 30% of tailings to be impounded within the open pit following the completion of mining activities. Geochemical testwork has shown that tailings are not expected to be acid generating and will not require sub-aqueous deposition. Consequently, the TMF will be a conventional terrestrial facility that uses waste rock from the open pit for construction of the impoundment dyke. The TMF will comprise two cells, with the initial cell located to the immediate north of the open pit and having a capacity of 444 Mt, and an expansion cell to the east having a capacity of 174 Mt. The maximum height for both cells will be 41 m. The remaining 273 Mt will be disposed of in the pit. 126

139 18.2 MINE DESIGN Mine design has been based on the following reports and is summarized below: Dumont Nickel Conceptual Study Update (Aker Solutions, August 2008). Conceptual Review Dumont Property (Genivar, November 2008). Dumont Nickel Project Open-Pit Optimization Study (Prysm Resources, June 2010). Dumont Project Preliminary Mine Plans (R. Kear, August 2010). Scoping Study for the Dumont Nickel Ore Project (BBA, July 2010) Overview For large-scale, low-grade deposits such as Dumont, the limits of open-pit mining are generally defined by the cost structure of the operation. The cost structure, in turn, is in large part a function of the scale of operation. The following iterative approach was used to generate the scoping study mine design: The cost structure was initially estimated using a zero-based techno-economic model and a production schedule generated from outputs of the previous conceptual study. NPV Scheduler mine planning software (incorporating a computerized Lerchs- Grossmann (LG) algorithm), was used to generate a series of nested pit shells based on the initial estimate of cost structure. The nested shells were evaluated using the techno-economic model. The shell that generated the maximum NPV was selected as the ultimate limit for the detailed mine design. The selected ultimate shell and nested shells contained within were used to generate detailed mine designs that honoured practical mining constraints. These designs included a high value and low-stripping-ratio starter pit along with five pushbacks into progressively lower-value and higher-stripping-ratio material. The nominal scheduling increment for the entire LOM plan was one year. An algorithm was written that allowed different schedules to be generated by adjusting one or more of the scope degrees of freedom, including milling rate, mining speed and cut-off grade. Schedules were evaluated using the techno-economic model. The schedule that generated optimal NPV while honouring existing constraints for the availability of power from the grid was selected as the base case design. The schedule that generated optimal NPV while utilizing the maximum electrical grid power available to the project on a planning basis was selected as the upside case. 127

140 Techno-Economic Model There are a number of degrees of freedom in the design of the Dumont open pit, each of which can have a significant impact on overall project economics. These include increasing concentrator throughput leading to the economics of scale, increasing mining rate to allow access to higher grade material sooner, increasing the cut-off leading to higher-value output over a shorter life. The techno-economic model allowed the impact on revenues as well as capital and operating costs of different scopes of design to be estimated. The capital cost estimate included the following items: A schedule for purchasing mining fleet, based on the fleet size and average accumulated hours for each unit necessary to achieve the production schedule. The capital cost of constructing the plant and associated infrastructure that was prepared by BBA, based on the process flow sheet defined by Royal Nickel s metallurgical consultants. Cost inputs to the LG algorithm include the following: Mining costs, including the cost of mining rock and overburden at surface, as well as the incremental cost for each successive bench the pit is deepened (these are applied to every tonne of material mined). Processing costs, and other costs that are applied to mill feed only. The LG algorithm considers only operating costs and capital costs that are variable in nature (e.g., the cost of expanding the tailings dam is a function of the tonnes that will be impounded). Fixed capital costs (e.g., the cost of mill construction) are not included. Table 18.2 summarizes outputs from the initial iteration of the techno-economic model that were used to generate the LG pit shells. Table 18.2 Cost Inputs to the LG Algorithm Area Item Units Value Surface Rock Mining Cost C$/t rock 1.15 Mining Costs Overburden Mining Cost C$/t overburden 1.65 Mining Cost Increment (rock or o/b) C$/15m bench 0.04 Total C$/t ore Milling Cost C$/t ore 9.00 Processing Costs Site G&A C$/t ore 0.65 Tailings Dam C$/t ore 0.50 Mill Sustaining Capex C$/t ore

141 Note that the marginal cut-off value of C$10.40/t was selected based on total processing-area marginal costs. All blocks in the NSR model with values less than C$10.40/t were classified as waste LG Ultimate Pit Shell The other key input to the LG algorithm was the open-pit slope design. The following slopes angles were assumed: The uppermost 15 m bench in overburden (assumed to be fine grained) = 18 Remaining overburden to the bedrock contact (coarse grained) = 27 Hanging wall rocks = 47 Footwall rocks = 52 Nested shells were generated by applying a revenue factor to the NSR value in each block. No value was attributed to the inferred resource within the open pit: for the purpose of the preliminary assessment it was treated as waste rock. The Dumont LG design was generated by Prysm Resources. This design included revenue factor increments of 1%. The lowest factor able to generate a pit shell was 38% (i.e., 38% of a $7.50/lb base price = $2.85/lb). After an analysis of the results, the various shells were clustered into the six phases that are summarized in Table It can be seen that for each successive phase, the average NSR value per tonne decreases, while the average stripping ratio increases. Table 18.3 Nested Pit Shell Phases Phase Revenue 000 tonnes Stripping Grade NSR Factor Mill Feed Waste Rock Overburden Ratio % Ni C$/t 1 56% 476, , , % $ % 400, ,319 45, % $ % 68,205 87,685 6, % $ % 56,967 99,318 5, % $ % 56,257 86,360 5, % $ % 44,587 86,825 5, % $17.11 The six phases were evaluated using the second iteration of the techno-economic model. This iteration allowed the individual phases to be scheduled sequentially, with all material contained in Phase 1 completed before Phase 2 commences. While this progression is not completely practical (in reality, the mining of Phase 2 would start well before the completion of Phase 1), it does allow the economics of pushbacks to be tested. The evaluation revealed that for production rates of 50,000 to 75,000 t/d, the maximum NPV was achieved with Phase 4. For 100,000 t/d, a slight improvement in NPV was achieved by including Phases 5 and 6. However, as the second iteration of the techno-economic model did not include a detailed derivation of costs associated with the TMF, the potential increase in costs associated with a larger facility was considered to outweigh the non-material improvement in 129

142 NPV resulting from inclusion of those phases in the design. Figure 18.1 shows the results of evaluation of the nested pit shells. Figure 18.1 Evaluation of Nested Shells The Phase 4 LG shell (generated with an 81% revenue factor) was thus selected as the basis for the detailed mine design (Figure 18.2). Figure 18.2 LG Ultimate Pit Shell 130

143 Detailed Mine Design The LG shell represents the optimal theoretical design, and while the final walls honour the imposed slope constraints, it cannot be considered a practical design as no provision is made for ramps. The practical design was generated by Robin Kear, using proprietary software. The nested pits contained within the ultimate pit shell were used to guide the development of a practical design that included a high-grade starter pit and five subsequent pushbacks. To confirm the feasibility of the planned mining sequence (i.e, to ensure there was ramp access to all benches for the entire LOM), interim face positions were developed on a nominal increment of 1 year. The end of life pit is presented in Figure Figure 18.3 End of Mine Life Pit Faces Base Case Design The base case mining schedule is illustrated in Figure

144 Figure 18.4 Base Case Mining Schedule Key elements of the base case include: Pre-stripping for 18 months prior to the start-up of the concentrator. A total of 93 Mt will be mined during the pre-strip, including: 50 Mt of overburden. 36 Mt of waste rock. Of this, 13 Mt will be used to construct the initial 10 m lift for the first cell of the tailings dam, while the remainder would be impounded in the waste rock dump. 7 Mt of mill feed grading 0.264% Ni, which will be stockpiled and used for commissioning the mill. A mining speed factor of 1.3, which results in mill feed being released faster than is required to satisfy the steady-state concentrator throughput of 80,000 t/d. Excess lower-value mill feed would be stockpiled during the 20-year LOM. After the pit is depleted in year 21, the low-grade stockpile will be re-handled to the mill over the remaining 11 years life of project. Figure 18.5 shows the base case schedule for mill feed. 132

145 Figure 18.5 Base Case Schedule for Mill Feed The strategy of stockpiling lower-value material allows the value of material treated during the initial years to be maximized. As a result, annual output averages 130 million pounds of nickel contained in concentrate during years 1 to 20, compared with 86 million pounds on average when lower-value stockpiles are treated during years 21 to 31 (see Figure 18.6, Table 18.4 and Table The strategy of accelerated mining has the additional advantage of creating a void (i.e., the mined-out open pit), which would accommodate approximately 30% of the tailings produced, thus reducing the surface footprint of operations. Figure 18.6 Base Case Nickel Production 133

146 Table 18.4 Base Case Detailed Production Schedule (Years 1 15) Mining Units LOM 1st 15 yrs Yr -2 Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Yr 11 Yr 12 Yr 13 Yr 14 Yr 15 Mill Feed 000 t 887, ,023-6,672 45,469 45,599 45,096 45,276 45,131 44,570 43,937 42,939 43,401 44,936 44,815 44,580 44,601 44,013 44,605 Waste Rock 000 t 952, ,820-36,025 34,012 33,881 34,378 34,206 34,365 34,903 35,558 36,540 36,079 34,532 34,659 49,914 72,770 75,786 75,170 Overburden 000 t 171, ,240 16,667 33,333 29,069 31,408 26,406 1,605 6, ,053 10,547 1, ,852 2,146 1,050 5,241 Total 000 t 2,011,319 1,293,084 16,667 76, , , ,879 81,087 86,273 80,019 80,548 90,026 81,253 79,526 79,474 97, , , ,017 Stripping Ratio Milling Units LOM 1st 15 yrs Yr -2 Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Yr 11 Yr 12 Yr 13 Yr 14 Yr 15 Ore Milled 000 t 896, , ,915 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 Grade Ni % Ni Contained Ni 000 lbs 5,418,298 2,414, , , , , , , , , , , , , , , ,620 Ni Recovery: % to Ni-Sulphide Conc 000 lbs 3,015,467 1,413, , , , , , , , ,597 99, , , , , , ,156 to Fe-Ni Conc 000 lbs 535, , ,255 16,873 15,733 18,659 17,351 19,056 19,475 21,845 21,184 19,750 19,026 14,045 11,984 10,343 10,769 Recovered Cobalt 000 lbs 104,379 43, ,026 3,423 3,473 3,421 3,475 3,374 3,306 3,260 3,280 3,355 3,423 3,492 3,460 3,304 3, Stockpile Units LOM 1st 15 yrs Yr -2 Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Yr 11 Yr 12 Yr 13 Yr 14 Yr 15 Tonnage - Beginning 000 t - - 6,739 26,748 43,603 59,949 76,478 92, , , , , , , , , ,832 of Year Additions 000 t 336, ,649-6,739 24,931 16,855 16,347 16,528 16,383 15,816 15,177 14,169 14,635 16,186 16,701 17,337 15,847 15,253 15,851 Subtractions 000 t 336,157 7, , , Tonnage - End of Year 000 t - 6,739 26,748 43,603 59,949 76,478 92, , , , , , , , , , ,683 Treatment Units LOM 1st 15 yrs Yr -2 Yr -1 Yr 1 Yr 2 Yr 3 Yr 4 Yr 5 Yr 6 Yr 7 Yr 8 Yr 9 Yr 10 Yr 11 Yr 12 Yr 13 Yr 14 Yr 15 Ni-Sulphide 000 t 3,710 1, Concentrate Payable Ni in Sulphide 000 lbs 2,804,385 1,314, , , , , , , ,766 93,555 92,735 98,339 98, , , , ,115 Payable Cobalt 000 lbs 43,550 18, ,295 1,487 1,528 1,460 1,502 1,441 1,403 1,339 1,352 1,413 1,450 1,552 1,567 1,525 1,553 Fe-Ni Concentrate 000 t Payable Ferro-Nickel 000 lbs 481, , ,529 15,186 14,160 16,793 15,616 17,150 17,528 19,661 19,066 17,775 17,123 12,641 10,786 9,309 9,692 Payable Cobalt 000 lbs Overall Payables %contained

147 Table 18.5 Base Case Detailed Production Schedule (Years 16+) Mining Units Yr 16+ Yr 16 Yr 17 Yr 18 Yr 19 Yr 20 Yr 21 Yr 22 Yr 23 Yr 24 Yr 25 Yr 26 Yr 27 Yr 28 Yr 29 Yr 30 Yr 31 Mill Feed 000 t 211,524 42,977 43,618 43,746 44,155 37, Waste Rock 000 t 259,646 76,785 76,152 71,045 32,833 2, Overburden 000 t 1,199 1, Total 000 t 472, , , ,791 76,988 39, Stripping Ratio 1.23 Milling Units Yr 16+ Yr 16 Yr 17 Yr 18 Yr 19 Yr 20 Yr 21 Yr 22 Yr 23 Yr 24 Yr 25 Yr 26 Yr 27 Yr 28 Yr 29 Yr 30 Yr 31 Ore Milled 000 t 461,322 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 23,322 Grade Ni % Ni Contained Ni 000 lb 2,626, , , , , , , , , , , , , , , , ,871 Ni Recovery: % to Ni-Sulphide Conc 000 lb 1,348, , , , , ,299 88,883 83,087 71,085 60,805 60,805 60,805 60,805 60,805 60,805 60,805 38,828 to Fe-Ni Conc 000 lb 282,167 16,679 10,559 8,190 7,799 5,345 18,486 20,736 21,600 22,340 22,340 22,340 22,340 22,340 22,340 22,340 16,396 Recovered Cobalt 000 lb 53,942 3,341 3,545 3,543 3,348 3,437 3,339 3,344 3,384 3,418 3,418 3,418 3,418 3,418 3,418 3,418 2, Stockpile Units Yr 16+ Yr 16 Yr 17 Yr 18 Yr 19 Yr 20 Yr 21 Yr 22 Yr 23 Yr 24 Yr 25 Yr 26 Yr 27 Yr 28 Yr 29 Yr 30 Yr 31 Tonnage - Beginning 000 t 247, , , , , , , , , , , , ,922 81,722 52,522 23,322 of Year Additions 000 t 81,404 20,703 16,836 15,144 15,557 13, Subtractions 000 t 329,087 6,496 1, ,967 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 29,200 23,322 Tonnage - End of Year 000 t 261, , , , , , , , , , , ,922 81,722 52,522 23,322 - Treatment Units Yr 16+ Yr 16 Yr 17 Yr 18 Yr 19 Yr 20 Yr 21 Yr 22 Yr 23 Yr 24 Yr 25 Yr 26 Yr 27 Yr 28 Yr 29 Yr 30 Yr 31 Ni-Sulphide Concentrate 000 t 1, Payable Ni in Sulphide 000 lb 1,253,670 95, , , , ,778 82,662 77,271 66,109 56,548 56,548 56,548 56,548 56,548 56,548 56,548 36,110 Payable Cobalt 000 lb 21,680 1,429 1,629 1,668 1,578 1,660 1,378 1,338 1,291 1,250 1,250 1,250 1,250 1,250 1,250 1, Fe-Ni Concentrate 000 t Payable Ferro-Nickel 000 lb 253,950 15,011 9,503 7,371 7,019 4,810 16,638 18,662 19,440 20,106 20,106 20,106 20,106 20,106 20,106 20,106 14,757 Payable Cobalt 000 lb Overall Payables % contained

148 18.3 MINING Overburden Mining Logging of diamond drill holes has showed that while there is some fine-grained clay and organic horizons within the overburden, material is predominantly composed of coarsergrained sand, gravel and boulders. Given this lithology, the overburden process design included the following key assumptions: Material will not require drilling and blasting. Material is sufficiently competent to permit use of conventional rigid-body trucks but will not support the 360-t class trucks planned for mining rock. It will be feasible to strip overburden continuously, 12 months of the year. Overburden will be impounded within a single facility located 150 m from the pit rim, with no segregation of fine- and coarse-grained material planned. The impoundment would be constructed in 5 m lifts, and would include a 5 m safety berm between lifts in order to achieve an overall slope angle of 22 (2.5H:1V). Overburden will subsequently be reclaimed from this facility, using the same fleet of equipment to be used for stripping overburden, in order to rehabilitate the terrestrial TMF and waste rock dumps. Covering all exposed faces with a layer measuring 0.4 m thick would consume approximately 15 Mt of the 171 Mt of overburden Rock Mining The mining of rock will be performed using the largest class of equipment that is commercially available. Rock will be mined on 15 m benches using the following primary equipment: Drilling with rotary blast hole drills (270 mm diameter holes). Loading with electrically powered rope shovels (55 m 3 capacity bucket). Re-handling with a large FEL (bucket capacity of 43 m 3 ). Hauling with 360-t payload-class haul trucks. Approximately 197 Mt, or 21% of waste rock, would be used for constructing the TMF. Material used in the initial 10 m lift will be delivered to the TMF using overburden haul trucks. This will entail the dumping of all waste rock used for dam construction at an appropriate location on surface, and then re-handling with an overburden FEL onto an overburden haul truck. After the initial 10 m lift was constructed, it will be possible to drive on the dyke with 360-t rock trucks. This will obviate the re-handling of the material (though it is assumed that 30% of the construction material will still require re-handling due to congestion or other operational issues). The remaining waste rock will be impounded in two facilities located 150 m from the rim of the pit. Rock would be dumped on 15 m lifts. The final slope of the dump would include an 11 m safety berm between lifts in order to achieve an overall slope angle of 24 (2.25H:1V). 136

149 Equipment Maintenance The scoping study assumed that mining plant and mobile equipment would be maintained inhouse by Royal Nickel personnel Mining Fleet Selection Fleet sizes were based on the following assumptions: The mine will operate 24 h/d, 355 d/y. The remaining 10 days will be lost due to a combination of holidays and inclement weather. The mechanical availability of equipment will average 87.5% over its economic life (reflecting initial availability of >90% dropping to 80% 85% in later years). The utilization of equipment will average 80%. An efficiency factor of 90% was applied to utilized time. These factors resulted in equipment averaging 5,964 engine hours per annum (the basis for calculating equipment running costs) and 5,368 effective work hours per annum (the basis for calculating the number of units required). Table 18.6 summarizes the main units of the mining fleet for both overburden and rock operations. In addition to purchase of the new units identified here, provision has been made in the sustaining capital estimate for rebuilding some of this equipment. Table 18.6 Mining Fleet Base Case Overburden Rock Size Example No. Reqd Size Example No. Reqd Unit Init. Sust. Init. Sust. Rotary Drill n/a mm Ø P&H XP Production FEL 13 m 3 Cat m 3 LeTorneau L Rope Shovel n/a - 55 m 3 P&H 4100 XPC Haul Truck 100 t Cat t Cat Track Dozer 13 m 3 Cat D m 3 Cat D Rubber Tyre Dozer 8 m 3 Cat m 3 Cat Grader 16-ft blade Cat 16M ft blade Cat 24M Water Tanker 40 t Cat t Cat One additional unit is required for the Upside Case. - One less unit is required for the Upside Case Support Equipment Open-pit haul roads will be maintained with a fleet of support equipment. 137

150 18.4 PROCESSING The flowsheet for the concentrator is shown in Figure Figure 18.7 Processing Flowsheet The key elements of the flowsheet are: Four-stage crushing, with Vertical Shaft Impact (VSI) crushers used for tertiary and quaternary crushing. Ore drying, to ensure moisture content in VSI feed does not exceed 2%. De-fibering of crushed mill feed using air classifiers. Grinding by ball mill, followed by de-sliming. Scavenger circuits to recover approximately 50% of nickel associated with the fibres and slimes (or approximately 5% of total contained nickel). Magnetic separation, with the magnetic concentrate further separated into ferro-nickel and nickel-sulphide concentrates by flotation. 138

151 Flotation of the magnetic separation tails to recover the bulk of nickel-sulphide concentrate, which will be combined with concentrates from the scavenger and magnetic circuits. Cleaning of the separate ferro-nickel and nickel-sulphide concentrates. The fibre, slimes and rock tails will be combined into a single tailings product for disposal. For the initial 20-year period, when the open pit will be active, the tailings product will be pumped to a tailings management facility (TMF) that will have an ultimate capacity of 618 Mt. For the remaining 11 years of project life, tailings produced from the treatment of lower-grade stockpiles will be pumped into the mined-out open pit Design Criteria The concentrator design was developed by BBA, based on the conceptual flowsheet provided by Mineral Solutions and the following design criteria: The plant will treat 80,000 t/d, with feed grading an average of 0.27% Ni over the life of mine. The plant will operate continuously, 24 h/d, 365 d/y. Conventional gyratory and cone crushers used for primary and secondary crushing will have availability averaging 75%. VSI crushers used for tertiary and quaternary crushing will have average availability of 85%. Grinding and flotation area equipment availability will be 95%. Dryers have been sized to treating 32% of the total feed, and reducing the moisture content from 4% to 2%. Equipment sizes are based on an assumed weight recovery to cleaner concentrate of approximately 2.0%. (The actual weight recovery is forecast at 0.5%). The average Bond ball work index is 21.3 kwh/t Primary and Secondary Crushing Run-of-mine mill feed will be delivered by truck into a 600-t capacity ore dump hopper. From the hopper, mill feed will be crushed to a P 80 of 140 mm by a single 60-in by 110-in gyratory crusher. Crusher discharge will be screened to remove fines (less than 25.4 mm) that will bypass the secondary crushing circuit. The coarse fraction will then be conveyed to one of four surge bins feeding the secondary crushing circuit. The secondary cone crushers will crush the coarse material to a suitable size for the drying process, with a P 80 of 45 mm. The crushing and screening facilities will be equipped with a dust extraction system. Dust collected by the system will be fed to the ore dryers. 139

152 Drying In order for the de-fibering process to be effective, feed material must contain less than 1% moisture. Moisture is primarily associated with the primary screen undersize material which represents approximately 32% of total run-of-mine feed. This size fraction is assumed to have a moisture content of 4%, which will be reduced to 2% by two dryers, 5.12 m in diameter and m in length. After drying, the moisture content will be further reduced by evaporation within the remaining stages of the crushing circuit. Discharge from the dryers will be conveyed to a covered stockpile with a live capacity of approximately 25,000 t, which is equivalent to 8 hours of feed. The preliminary assessment design assumes that dryers would burn fuel oil. During the prefeasibility study, the potential benefits of using natural gas should be assessed Tertiary and Quaternary Crushing Discharge from the secondary crushers and ore dryers will be combined at the secondary screening plant. Double deck screens will divide material into three size fractions: Oversize (>30 mm) will be re-crushed in an open circuit cone crusher, with product fed to the tertiary crushers. Intermediate (>6 mm and <30 mm) will be fed to the tertiary crushers. Fines (<6 mm) will bypass the tertiary circuit and be fed directly to the quaternary crushers. Vertical Shaft Impact (VSI) type crushers will be used for both tertiary and quaternary crushing circuits. With a VSI crusher, material is introduced into a high-speed rotating impeller, and the crushing action is autogenous, as a result of rock-on-rock impact. Tertiary crusher discharge will have a P 80 of 6 mm and will be combined with the fine fraction from the secondary screens to feed nine quaternary crushers. These will reduce the P 80 to 840 µm. Quaternary crusher discharge will be treated through air classifiers to remove fibrous material. A dust collecting system will be provided for all screens and crushers Grinding Grinding will be conducted at low density given the high viscosity associated with serpentine mineralization. The circuit will comprise three parallel ball mills, 6.1 m diameter by 7.9 m long, driven by 7.7 MW motors, operating in closed circuit with a circulating load of 250%. 140

153 Ferro-Nickel Circuit Discharge from the grinding circuit will be de-slimed to remove ultrafine brucite slimes. Underflow from the de-sliming cyclones will be separated into ferro-nickel and nickel sulphide streams using a low-intensity magnetic separator (LIMS), based on the strong magnetic characteristics of the ferro-nickel mineral (awaruite) compared with the sulphide minerals (pentlandite and heazlewoodite) which are both weakly magnetic. Given the magnetic properties of sulphide minerals, however, the LIMS rougher concentrate will contain a significant amount of sulphide minerals. These will be separated using flotation. Ferro-nickel concentrate will also be upgraded using rougher flotation followed by a single stage of cleaning. Recovery of nickel to cleaner ferro-nickel concentrate will average approximately 10% of total contained nickel, or 15% of nickel recovered. For the typical grade of mill feed that will be processed, the nominal grade of ferro-nickel concentrate is 25% Ni Nickel Sulphide Circuit Tailings from the LIMS will be pumped to the rougher flotation circuit. Recovery of nickel to cleaner nickel-sulphide will average approximately 56% of total contained nickel, or 85% of the nickel recovered to concentrate. The nominal grade of nickel sulphide concentrate will be 35% Ni Scavengers Depending on the style of mineralization, an estimated 6% to 10% of total contained nickel will be associated with the fibres removed during air classification after the quaternary crushing stage. A further 4% to 5% of total contained nickel will be associated with slimes that will be removed prior to the magnetic separation. Recovery of nickel to the fibre scavengers is expected to range between 33% and 50%, depending on the style of mineralization. For the slimes scavenger, recovery is expected to range between 40% and 50%. Scavenger concentrate will be introduced to the nickelsulphide stream at the cleaning stage Concentrate Dewatering and Load-Out Slurry density of both cleaner concentrates will be increased from 19% to 65% using 24 m diameter thickeners. Thereafter, concentrates will be filtered to further reduce moisture to approximately 6%. From the filter press, concentrates will be conveyed from the mill site to a load-out facility close to the existing Canadian National Railway Company (CN) line. 141

154 Tailings Thickening Tailings from the fibre, slimes and flotation streams will be combined into a single product, which will have a relatively low density of 45% solids after thickening. During operation of the open pit, underflow from the tailings thickener will be pumped to the TMF using positive displacement pumps. After open-pit operations are complete, tailings generated from the treatment of lower-grade stockpiles will be pumped into the mined-out pit using lower-cost centrifugal pumps Product Quality Nickel Sulphide Concentrate Nickel sulphide concentrate is anticipated to contain average approximately 35% Ni over the life of the operation, with 1% Co, 2.5 g/t precious metals (platinum, palladium and gold), and 7-10% MgO. The concentrate grade is forecast to vary as a function of grade of ore treated and recovery of nickel to nickel sulphide concentrate. In the initial 20 years of operation, when the cut-off grade for milling is elevated and lower-grade material is stockpiled, concentrate grades of up to 40% Ni are expected. After the pit is depleted and lower-grade stockpiles are reclaimed, the concentrate grade is expected to fall to 20% to 25% Ni. No metallurgical test work has been performed to estimate the percentage of platinum, palladium and gold that would be recovered to concentrate. Assuming recovery of these three metals is 50%, the nickel sulphide concentrate would grade approximately 2.5 g/t precious metals. The nickel sulphide concentrate will contain approximately 65% 75% by mass nickel minerals (predominantly pentlandite and heazlewoodite, with lesser amounts of awaruite). The remaining mass will include magnetite but will be dominated by serpentine. Serpentine contains varying levels of magnesium, depending on the degree of metamorphic alteration, with the range typically being 30% to 40% MgO. As a result, the likely range for MgO content in the nickel sulphide concentrate is expected to be between 7% and 10% Ferro-Nickel Concentrate The ferro-nickel concentrate is expected to contain an average of 25% Ni over the life of the operation, with iron and some cobalt, chromium and copper. The grade of nickel in the ferro-nickel concentrate is generally inversely correlated to that of the nickel sulphide concentrate. During the early years of operation, when the sulphide grades in mill feed will be high, the awaruite content in mill feed will be relatively low and the grade of ferro-nickel concentrates is expected to be approximately 20%. The low-grade 142

155 stockpiles will have a higher proportion of awaruite mineralization, and concentrates are expected to grade approximately 30% Ni in the later years of production INFRASTRUCTURE Site Layout The total area of all concentrator facilities would be approximately 22,000 m 2, occupying a footprint of 300 m by 200 m. Figure 18.8 illustrates the proposed site layout at the end of open-pit operations. Figure 18.8 Site Layout at the End of Open Pit Operations Key elements include: The open pit, from which 2,011 Mt of mill feed and waste will have been excavated. The ultimate pit will measure 3.7 km by 1.1 km and will be approximately 650 m deep. 143

156 A temporary stockpile of lower-grade mineralization. At the end of open-pit operations, this stockpile will contain 320 Mt grading 0.24% Ni, sufficient for 11 years of mill feed at 80,000 t/d. Two waste rock storage facilities that will contain a total of 760 Mt. A TMF which, at the end of open-pit operations, will contain 618 Mt of tailings, while the impoundment dykes will contain 197 Mt of waste rock. An overburden storage facility, containing 171 Mt. The main administration buildings, warehouse, fuel farm and mine workshops. These are located close to the existing highway and CNR rail line. The mill, which is located close to the TMF in order to minimize pumping costs. Load-out facility located on a railway spur, close to the administration building. At the end of project life, the open pit will have been partially filled with tailings produced over the final 11 years life of the operation, the temporary low-grade stockpile will have been reclaimed. Approximately 15 Mt will have been removed from the overburden stockpile and used to rehabilitate waste rock storage facilities and the TMF Site Infrastructure The following site infrastructure will be required to support mining and processing operations: Administration building. Concentrate warehouse. Fuel farm, with a facility to store diesel (mainly for the open pit) and fuel oil (for the concentrator driers) separately. Total capacity of the fuel farm will be approximately 10 days average consumption, or 2 ML of diesel and 3 ML of fuel oil. Mine workshop based on the maximum fleet size of t haul trucks, and associated support equipment. The other major production equipment (shovels and drills) would be maintained in the pit, with components changed out and repaired in the workshop. Electrical sub-station at 120 MW and associated reticulation system. Sewage treatment and a landfill site. 144

157 Off-Site Infrastructure Off-site infrastructure comprises an electrical power line to connect with the grid and water pumping station. The power line will be approximately 28 km long and will connect the site to the grid near Amos. The water pumping station will be located on the Villemontel River, immediately south of the Dumont property and approximately 3 km from the mill Logistics The scale of operation at Dumont would result in significant consumption of items such as diesel, fuel oil, explosives, grinding media and reagents. The project will also produce significant quantities of concentrate requiring transport to off-takers. It is planned to transport the majority of items by rail, in order to minimize the impact on local road traffic MANPOWER Average manpower requirements will vary through the life of the operation and, in particular, will be reduced over the final 11 years when mining has ceased. The average complement of full-time employees over the life of project for the base case will be 515 persons. During the construction phase, it is estimated that employment will peak at 1, ENVIRONMENTAL STUDIES Environmental studies are well advanced, with the following work having been completed: Three phases of environmental baseline studies were completed during the period in order to establish the pre-development environmental condition of the property and identify potential areas of impact. A preliminary geochemistry study on a representative sample of mineralization, waste and potential tailings from the Dumont deposit to determine acid rock drainage and leaching characteristics. Dumont s tailings are composed of a rock type (serpentine) that has demonstrated the ability to passively fix carbon dioxide. A carbon sequestration study is ongoing to assess the potential for Dumont s tailings to offset the carbon footprint of the mining operation. Future studies are planned and include: Further geochemical analysis to fully understand and predict the behaviour of tailings, with a particular focus on the potential for metal leaching. 145

158 Construction of an experimental in-situ tailings cell to quantify the potential for carbon sequestration under operating conditions. Hydrological studies to quantify the impact of proposed operations on the local water table and a nearby aquifer-bearing esker. Quantifying the impact of mining operations on existing wetlands and fish habitats, and identifying opportunities for mitigation. Characterization of the soils in the area that would be impacted by operations. Once the project scope is finalized during the pre-feasibility study, a Project Notice will be submitted to the Quebec Ministère du Dèveloppement durable, de l Environnement et des Parcs (MDDEP), or Ministry of Sustainable Development, Environment and Parks. MDDEP will accordingly advise on the scope and requirements of an environmental impact study (EIS). The project scope is such that this study would be assessed jointly at the provincial and federal levels under the Canada-Quebec Cooperation Agreement. It is expected this assessment could take up to two years from the time of submission of the Project Notice before the granting of a Certificate of Authorization to commence construction. The assessment period would run in parallel with the feasibility study and detailed engineering and the overall impact on the project s critical path would thus be minimal Environmental Management Plan Baseline studies during identified the following key environmental issues: The scale of operation, including its required operational footprint relative to the available surface area. The presence of chrysotile fibres in mill feed. The water balance. The impact of operations on the local water table and a nearby aquifer-bearing esker. The impact of operations on local communities. 146

159 Scale of Operation The Dumont project is based on a large low-grade deposit that requires economies of scale in order to be economically viable. The surface area available for operations at Dumont is limited by the boundary between the Arctic and St. Lawrence watersheds and two northsouth running eskers located to the east and west of the property, respectively 10 km and 6 km from the nearest edge of the proposed pit. As a result, the base case design includes the following features that minimize the operational footprint: Accelerated mining of the open pit will allow approximately 30% of the total tailings produced to be impounded within the mined-out pit shell. The ultimate height of rock dumps (150 m) minimizes the surface footprint of these impoundments. Some overburden removed in the course of stripping will be used to rehabilitate other impoundments, including the waste rock dumps and the TMF, thus reducing the extent of the overburden impoundment Fibrous Minerals As is common for ultramafic deposits, mineralization at Dumont contains chrysotile, which will release fibres during processing operations. The current processing concept is based on four-stage crushing to a particle size of less than 1.0 mm, in order to extract dry fibres. The crushing and screening operations will be equipped with dust extraction systems to ensure personnel are not exposed to any fibres. Chrysotile will be mixed with one or more of the other tailings products to ensure there are no fugitive airborne emissions from the TMF Water Balance The water balance developed for the scoping study indicates that an average of at least 95% of process water requirements could be met through the re-use of process water (from the TMF) or from inflows to the pit that would be captured in a sump. The remainder, averaging 337 m 3 /h (or 93 L/s), will be drawn from the Villemontel River Water Table The impact of operations at Dumont on the local water table is expected to be minimal, given that a preliminary environmental geochemistry study, including acid-base accounting, leaching and humidity cell testing, indicates that neither mineralization nor waste rock are expected to generate acid. 147

160 While there is no legislation or guideline regarding operating in proximity to an esker, operating infrastructure (including operating facilities and waste impoundments) has been designed to come no closer than 1 km from either esker. The open pit is located sufficiently far from the nearest esker (5.5 km at its closest point) that the draw-down cone resulting from the pit is not expected to reach the esker Local Communities The creation of well-paying industrial positions at a long-life operation is expected to be of net benefit to local communities. Potential negative impacts that the project may have on local communities, include increased traffic, which will be mitigated by maximising the use of rail for delivering consumable items and shipping concentrate and the impact on the local landscape as well as noise and dust generated by operations. These impacts will be mitigated by maintaining an adequate distance between operating infrastructure and populated areas Waste Impoundments Studies to date indicate the waste products at Dumont would be benign, with no acid generated. Thus the TMF will not require sub-aqueous disposal and nor will it be necessary to line the TMF with an impervious membrane. Furthermore, clay in the underlying overburden, where present, will essentially act as a low permeability membrane. Royal Nickel has initiated a study to determine the potential for the Dumont tailings to leach metals under neutral basic ph conditions in order to fully predict tailings behaviour Carbon Sequestration The tailings are predominantly composed of serpentine, which is a rock type with the demonstrated ability to passively fix carbon dioxide. Preliminary work has shown that if only 10% of the sequestration potential of the tailings is realized, the carbon footprint of the Dumont operations would be completely offset. A carbon sequestration study is ongoing. This study entails construction of an experimental in-situ tailings cell, to quantify the potential for carbon sequestration under operating conditions ENVIRONMENTAL PERMITTING Two pieces of legislation control the environmental assessment and granting of operating licences for mining operations in Quebec: 148

161 Provincial Environmental Quality Act (Loi sur la qualité de l environnement) (L.R.Q., c. Q-2). The Federal Canadian Environmental Assessment Act (CEAA). All projects are subjected to the provincial legislation. All projects are also initially screened to determine if the scope of the project triggers any federal legislation. Current federal legislation identifies four triggers that would necessitate federal assessment of the project of which the location of the Dumont project on land administered by the federal government is applicable. The scope of the project will result in permits being required from a number of federal departments, including: Department of Fisheries and Oceans Canada (DFO) as baseline studies have identified several species of fish inhabiting wetlands within the footprint of disturbance. Natural Resources Canada (NRCan) as storage and manufacture of explosives requires a licence from this agency. The scale and impact of the Dumont Project on local hydrology may trigger a more extensive assessment based on federal requirements and it has been assumed that the proposed project will require, at a minimum, a Comprehensive Study Assessment. The permitting process is initiated with the submission of a Project Notice to the Quebec Ministère du Dèveloppement durable, de l Environnement et des Parcs (MDDEP). The Project Notice describes the scope of the project and provides a summary of likely environmental impacts. The Project Notice should be based on the PFS design, to ensure there are no major scope changes that would later require the updating and re-submission of the Project Notice. The Project Notice is assessed jointly at the provincial and federal levels. Instruction is then given by the MDDEP on the scope and requirement of an environmental impact assessment and the nature of public consultation required to obtain a Certificate of Authorization MARKETING The preliminary assessment is based on the production of a conventional nickel sulphide concentrate and a ferro-nickel concentrate. As yet, no off-take agreements have been entered into; hence the preliminary assessment is based on assumed commercial terms. The nickel sulphide concentrate will grade 35% Ni and 1% Co. The MgO content of this concentrate is expected to be between 7% and 10%, which is in line with the MgO content in concentrates produced by other ultramafic operations. The nickel sulphide concentrate is 149

162 also expected to contain potentially economic concentrations of platinum group metals and gold. The ferro-nickel concentrate will grade 25% Ni with some cobalt, chromium and copper values in addition to iron. It is anticipated that the nickel sulphide concentrate will be treated conventionally, at smelters and refineries, to yield pure nickel. Assumptions regarding commercial terms for this concentrate have been based on benchmark rates and include: Percentage payables. Base price. Price participation escalator. Penalty for magnesium (MgO). Payment for some of the contained cobalt. It is intended that the ferro-nickel concentrate will be marketed directly to stainless steel producers, which are the primary end users of nickel. There are fewer available benchmarks for commercial terms, and assumptions used include: A conservative estimate regarding percentage payables for contained nickel. No payment for iron, cobalt, chromium or copper Nickel Sulphide Concentrate There are currently 11 nickel smelters globally, while a twelfth unit that will also treat sulphide concentrates, the Vale facility in Newfoundland, is under construction. Of these facilities, the operations of Xstrata and Vale in Canada, the Boliden facility in Finland, and Jinchaun s operations in China will be of greatest interest. Given the high nickel content of Dumont concentrate, it is potentially saleable for direct refining or direct shipment to stainless steel producers. The Xstrata smelter located in Falconbridge, Sudbury, currently treats concentrates produced by Xstrata s operations located in the Sudbury basin (the bulk coming from the new Nickel Rim South mine) and in Quebec (Raglan), as well as from third parties. The smelter uses electric furnace technology, which is suitable for treating concentrates containing elevated levels of MgO. The smelter has an estimated full capacity of 76,000 t/y. Xstrata currently purchases and tolls feeds from many sources to fully utilize the processing capacity. Matte produced by the Falconbridge smelter is shipped to the Nikkelverk refinery in Norway. Overall cobalt recovery through the smelter and refinery is approximately 70%. The main smelter of Vale is located at Copper Cliff, Ontario. Flash smelting technology, is less suitable for treating concentrates containing elevated levels of MgO. However, concentrates from Dumont could be blended with Sudbury-based mine feed that is low in 150

163 MgO as well as other purchased or tolled feeds to take advantage of the relatively high nickel content while maintaining the overall MgO limits of the smelter. The Vale smelter at Thompson, Manitoba smelter employs electric furnace technology and treats concentrates with various levels of MgO from the Thompson area. However, the ability of Thompson to treat Dumont concentrate may be constrained by MgO content. Vale is currently constructing a hydro-metallurgical facility in Newfoundland to treat concentrate from Voisey s Bay. The design criteria are understood to be based on the mineralogy of the Ovoid open pit, which has extremely low levels of MgO and other contaminants and is unlikely to be able to treat Dumont concentrate with its current configuration. At a later date, however, if concentrates from outside Newfoundland and Labrador are treated, the facility may be modified to accommodate concentrates with more impurities, including MgO. Boliden operates the Harjavalta flash smelter in Finland and output is refined at the adjacent Harjavalta refinery, owned by Norilsk. The Harjavalta smelter has a capacity of approximately 45,000 t/y of contained nickel depending on the concentrate grade processed. The smelter can accommodate some quantity of magnesium-bearing concentrates. The Harjavalta refinery has a capacity of approximately 65,000 t/y and is beginning to receive direct intermediate feeds from Talvivaara. The complex achieves high recoveries for cobalt. Jinchuan operates an integrated smelting and refining facility in Gansu Province, China. The smelter currently has a capacity of around 120,000 t/y contained nickel, while the refinery has a capacity of some 150,000 t/y contained nickel. Over 40% of the concentrate feed to the Jinchuan smelter currently comes from third party sources. It is understood that the Jinchuan facilities have the capability to take MgO-bearing feeds and will continue to need third party concentrates in order to maintain capacity utilization Ferro-nickel Concentrate The ferro-nickel concentrate will be marketed directly to stainless steel producers, the primary end users of nickel. There are a large number of stainless steel producers using a wide range of various primary and scrap materials. Potential consumers include: Acerinox (with operations in Spain, the United States and South Africa). Allegheny Stainless in the United States. Thyssen Krupp in Germany and Italy. Arcelor Mittal in Belgium and France. POSCO in Korea and China. Tang Eng and Walsin in Taiwan. YUSCO in Korea and China. TISCO, Baosteel, and others in China. 151

164 Commercial Terms The commercial assumptions for nickel-sulphide concentrate were based on current benchmarks and include the following: Transportation charges of $115/t, to cover rail transport to Quebec City (approximately 500 km) followed by ocean transport to Europe (Boliden), Russia (Norilsk) or China (Jinchuan). In the event that concentrate is treated in Sudbury (approximately 500 km by rail), transportation charges would be significantly lower. Base TC (treatment charge) for smelting of $125/t, with a $25/t penalty for MgO content. Base RC (refining charge) of $0.70/lb, for a base nickel price of $5.00/lb. The scoping study evaluation was based on a long-term Ni price of $7.50/lb. Price participation escalator has been assumed to be 10% of the excess over a base price of $5.00/lb Ni. This results in a price participation charge of $0.25/lb. Payment for 93% of nickel in concentrate. Payment for 50% of contained cobalt in concentrate. Refining charges for payable cobalt were assumed to be $3.00/lb. Since no metallurgical work has been performed to determine the grade of PGEs and other potentially economic by-products in concentrate, no payment has been assumed for these. The commercial assumptions for ferro-nickel concentrate include the following: Transportation charges of $115/t to cover rail transport to Quebec City (approximately 500 km) followed by ocean transport to stainless steel producers located in east Asia (China, Taiwan or Korea), Europe or the United States. No charges for treatment and refining, since the ferro-nickel concentrate would not require upgrading before end use in steelmaking. Payment for 90% of contained nickel in concentrate based on the equivalent stainless steel scrap value of feed to the process. It has been assumed that no payment would be received for iron, chromium or other potentially valuable elements contained in the concentrate SCHEDULE The project development schedule indicates that initial commercial production will take place in the second half of 2015, with full production being achieved in the first quarter of See Figure

165 Figure 18.9 Project Development Schedule Critical path items are shown in red. The schedule assumes that the Project Notice will be submitted to MDDEP as soon as the pre-feasibility design has been completed. A total of 24 months has been allowed from submission of the Project Notice to receipt of the Certificate of Authorization to proceed with construction The timeline to receipt of licences for operation can be increased by up to three months (to 27 months overall) without impacting on the critical path. Re-zoning of claims currently classified as agricultural lands will take place in parallel with the permitting process. The impact to overall project schedule resulting from the permitting process will be minimised by initiating procurement of long lead-time items and completing much of the detailed engineering design prior to receipt of the Certificate of Authorization. The key equipment items for process plant, with approximate lead times for delivery, are the ball mills (58 weeks), primary crusher (53 weeks) and transformer (43 weeks). The schedule currently includes an allowance of 18 months for detailed engineering prior to the start of plant construction at site. This activity is on the overall critical path. 153

166 18.11 CAPITAL COST ESTIMATE Summary The capital cost estimate summarized in Table 18.7 is expressed in real January, 2010 terms and assumes a long-term exchange rate of US$0.90/C$1.00. Details of the various estimate areas are provided in the following sections. US$ millions Table 18.7 Capital Cost Summary Base Case (80,000 t/d) Upside Case (100,000 t/d) Initial Capital Mine $448 $457 Process Plant $709 $859 Tailings Dam $124 $138 Infrastructure $131 $152 Indirects $242 $274 Contingency $369 $424 Sub-Total $2,023 $2,304 Sustaining Capital Mine Fleet $331 $354 Mill $367 $361 Tailings Dam $159 $153 Closure $100 $100 Contingency $181 $182 Sub-Total $1,139 $1,150 Total Capital $3,162 $3, Mining Capital The cost estimate for the mining fleet was generated from first principles, using the mine production plan to estimate the fleet requirements. The cost of each unit was taken from a database of equipment parameters provided by various original equipment manufacturers (OEMs). Overburden and waste rock pre-stripping costs were also generated from first principles, using the operating cost model. Accordingly, the estimate is considered to be of a higher level of accuracy than is normally required for a preliminary assessment. For the upside case, the initial mining fleet is not materially different from the base case, and the initial capital estimate is 2% ($8.5 million) higher. This reflects the similarity in open-pit mining rates for each case the mining rates for overburden are identical, while the mining rate for rock is <10% higher than in the base case. The main difference between the two schedules is in the stockpiling of lower-grade material; with the higher processing rate for the upside case, the lower-grade stockpile reaches a maximum size of 250 Mt, or 70 Mt less than 154

167 the base case. Sustaining capital increases by 7% ($23.2 million), including replacement of a rope shovel and blast-hole drill in the upside case. Table 18.8 shows the mining capital cost estimate for the base case Processing Capital Table 18.8 Mining Capital Cost Summary Base Case US$ thousands Initial Sustaining Rotary Drill 20,674 10,337 Blast Truck OB FEL 11,971 - Tailings FEL1 4,788 - Tailings FEL2 8,745 - Shovel 68,606 40,020 OB truck 32, Tailings Truck1 1,269 - Tailings Truck2 5,642 - Bulk Truck 95, ,444 OB TD 2,044 - TD 2,773 4,160 Tailings TD 1,022 1,022 OB RT 2,114 - RT 3,926 5,889 Tailings RT 1,057 1,057 OB Grader 1,584 - Grader 5,172 6,896 Tailings Grader OB Tanker Tanker 4,938 1,235 Tailings Tanker FEL 8,745 26,234 Service Truck 1,559 1,871 Pick-ups 832 1,538 Extra 1,269 - Ancillary Equipment 7,217 8,079 Sub-Total Mining Fleet 297, ,255 Overburden Pre-Stripping 71,405 - Rock Pre-Stripping 79,810 - Total 448, ,255 The capital estimate (±40%) for the process plant was generated by BBA, based on the conceptual flowsheet described earlier. BBA had initially generated an estimate for a 75,000 t/d flowsheet. In order to estimate costs for 80,000 t/d (the base case) and 100,000 t/d (the upside case), factors were developed to account for the change in the number and/or size of individual pieces of equipment (and associated infrastructure) in order to achieve a different 155

168 production rate. As indirect costs include a greater percentage of fixed costs than direct costs, different factors were used to adjust the direct and indirect cost line items. The resulting estimates for the base and upside cases are given in Table Table 18.9 Process Plant Capital Cost Summary US$ thousands Base Case Upside Case Concrete Works 89, ,419 Structural Works 60,839 73,936 Architectural Works 30,985 37,655 Mechanical Equipment 299, ,934 Building Mechanical 22,955 27,897 Piping 62,590 76,064 Electrical 80,529 97,866 Automation 33,900 41,197 Capitalized Operating Costs 28,420 34,588 Total Process Plant Capital 709, , Tailings Management Facility The capital cost of tailings pumps and associated infrastructure was estimated by Golder PasteTec and assumes that because of the anticipated viscous nature of the material, the volume of slurry to transport and the distances involved in discharging tailings along the entire perimeter of the tailings dam, piston diaphragm pumps would be required. In the event that future rheological testing of the thickened tailings demonstrates that the material could be pumped using centrifugal pumps, a capital saving in the order of $40 million may be possible. Costs for sedimentation ponds and site preparation including a catchment trench at the toe of the dam were estimated by Golder. Costs for earthworks required in constructing the starter lift of the impoundment dyke were estimated using the same zero-based model used for estimating mine operating and capital costs, and account for: The tonnage of material used in dam construction and 1-way haulage profiles from the rim of the pit The size of equipment used (100 t class haul trucks) Support equipment (e.g., track dozers, rubber tire dozers, graders and water carts) Maintenance of mining equipment The estimated cost (Table 18.10) does not include the capital cost of incremental mine fleet purchases (i.e., the additional loading and hauling necessitate the purchase of 2 FEL and 11 trucks), as these have been included in the overburden fleet provided in the mining capital estimate; following construction of the starter lift, this equipment would be available for overburden mining. 156

169 Table Tailings Management Capital Cost Summary US$ thousands Base Case Upside Case Pumps 82,633 95,913 Site Preparation 13,639 13,639 Sedimentation Ponds 4,500 4,500 Construction 23,408 23,726 Total Tailings Management Capital 124, , Infrastructure Capital Infrastructural capital (Table 18.11) was estimated by BBA, based on its database of costs for projects of a similar scope. The level of accuracy of the estimates is consistent with the requirements of a preliminary assessment Indirect Capital Table Infrastructure Capital Cost Summary US$ thousands Base Case Upside Case Site Preparation 35,076 42,628 Truck Shop 7,762 11,643 Wash Bay 1,641 1,992 Road Stone Crusher 9,000 9,000 Warehouse 1,511 1,881 Admin Building 2,096 3,145 Load Out 10,125 10,125 Conveyor to Load Out 15,750 15,750 Tank Farm 11,250 13,500 Protective Domes 2,967 3,560 TMF Pump Infrastructure 13,500 16,200 Explosives Plant 8,251 10,084 Water Treatment Plant 3,150 3,780 Electrical Sub & Distribution 9,000 9,000 Total Infrastructure Capital 131, ,288 The majority of indirect costs (Table 18.12) were estimated by BBA, based on factors applied to various direct costs. This methodology is appropriate for a preliminary assessment, to generate an estimate of indirect costs. Capitalized G&A costs were generated from the zero-based model, taking account of the timing for recruitment of the management team. 157

170 Table Indirect Capital Cost Summary US$ thousands Base Case Upside Case Owner's Costs 53,560 61,573 EPCM 66,233 76,141 Temporary Facilities 34,460 39,615 Temporary Operation 30,400 34,948 Mobile Equipment 3,364 3,867 Freight + Vendor's Reps 30,400 34,948 Capitalized G&A 23,173 23,173 Total Indirect Capital 241, , Initial Capital Contingency Two levels of contingency were assumed: for the mine fleet and pre-stripping, where quantities are well defined based on the engineered mine design, a contingency of 15% was applied; for all other items, a contingency of 25% was applied. The breakdown of the contingency applied to initial capital is given in Table Table Initial Capital Cost Contingency US$ thousands Base Case Upside Case Mining Fleet 44,578 45,527 Pre-Stripping 22,682 23,012 Process Plant 201, ,436 Other 41A 87, ,845 Other (Capitalized Operating Cost) 12,898 14,440 Total Initial Capital Contingency 368, , Sustaining Capital The estimate of sustaining capital provided in Table includes the following items: Replacement and rebuilding of the mine fleet, the schedule of which was based on the mining production schedule. Replacement of major components within the mill, as estimated by BBA, based on the initial cost of construction. Note that the upside case cost is less than the base case cost due to the reduced life of this scenario (25 years vs. 31 years). Costs associated with progressively expanding the TMF, which were estimated in the same manner as the initial capital estimate. Note that the bulk of material used to construct dam extensions would be transported using 360 t trucks travelling on the starter lift that had been constructed using smaller units, and thus has a lower associated cost of transportation. 158

171 Closure costs, which make provision for the following: o Removal of all permanent structures, estimated to cost $45 million. o Rehabilitating the TMF and waste rock dumps with stockpiled overburden, the cost of which was based on the estimated cost of stripping overburden (as the same units would be used). Costs equated to C$0.025/tonne of waste impounded, or a total of $42 million. o A provision for ongoing treatment of water. The combined annual cost of water treatment plant operations and maintenance was estimated to be approximately $1.35 million, based on costs at similar operations. The NPV 10% (at closure) of incurring this cost in perpetuity was $13 million. Under Quebec s current Mining Act, a reclamation bond must be posted to cover 70% of the cost to reclaim areas of accumulation, soil stabilization, securing openings and pillars, constructing water treatment plants and reclaiming roads. Proposed revisions to the Mining Act, as included in Bill 79, will increase the financial guarantee from 70% to 100% of the estimated reclamation cost and include more activities than those presently covered. Accordingly, in the cash flow model, closure costs are split into (i) bond payments, incurred over the first five years of operation (amounting to 100% of the costs for which bonds are required), and (ii) decommissioning and monitoring costs, reflected as a lump sum at the end of the mine life. Table Sustaining Capital Cost Summary US$ thousands Base Case Upside Case Mine Fleet 331, ,509 Mill 367, ,048 Tailings Dam 159, ,626 Closure 99,986 99,986 Contingency 181, ,595 Sustaining Capital Cost 1,139,054 1,149, Working Capital Provision has been made for changes in working capital, with the average stores holding assumed to be 1 month s worth of all consumable items. The investment in working capital peaks at C$32 million for the base case and C$37 million for the upside case. As presented, the cash flow model takes full account of the revenue generated from production in each period. In practice, this will be difficult to achieve, given that some metal will be retained as work in progress and as finished product in transit to the off-taker. Furthermore, no provision is made for accounts receivable. As project development proceeds, it is recommended that these items be provided for in any future evaluation of the project. 159

172 18.12 OPERATING COST ESTIMATE Summary Operating costs were estimated in the following manner: Open pit mining costs were estimated from first principles, based on the mine schedule, performance parameters for mining equipment as recommended by OEMs, and the current cost of commodities and labour rates. This estimate is considered to be of a higher level of accuracy than is normally required for a preliminary assessment. Processing costs were estimated by BBA, based on rates of consumption for reagents and other consumables determined from metallurgical test work, and a labour structure that is appropriate for the proposed flowsheet. Operating costs for the TMF were estimated by Golder, based on the design of the facility, including pumping distances, and the associated number of pumps required. General and administrative costs were estimated from first principles, based on the level of support required for the operation. Costs for treatment and refining of concentrates were based on the assumed commercial terms described earlier, and the scheduled production of concentrates. A summary of operating costs is provided in Table Table Estimated Site Operating Cost Summary Area Units Base Case (80,000 t/d) Upside Case (100,000 t/d) Mining US$/t mined US$/t treated Processing US$/t treated G&A US$/t treated Sub-Total Site Costs US$/t treated US$/lb Ni TC/RCs US$/lb Ni By-Product Credit US$/lb Ni (0.16) (0.16) Net C1 Cash Costs + US$/lb Ni C1 costs include mining, processing, site administration and refining, net of by product credits Labour Labour costs were estimated based on the organizational structure developed for each area and rates of pay estimated from existing mining operations in the Quebec/Ontario region. These estimates represent the total cost to the company including benefits such as overtime, pension and insurance. 160

173 Energy The operating cost estimate includes assumed unit prices for energy of C$0.05/kWh, which includes: The existing base rate for Quebec (approximately C$0.04/kWh). A provision for Hydro-Quebec to recover its investment in extending power lines to the site. The proposed tariff increases discussed in the 2010 provincial budget. Long-term prices for diesel and fuel oil of US$0.80/L and US$0.54/L, which are based on a long-term oil price forecast of US$80/bbl Mining Operating Costs Operating costs for the open-pit mine were estimated from first principles using a zero-based model and the production schedule. Resulting estimates of LOM average costs for the upside case ($1.52/t ex-pit) are approximately 3% lower than the base case ($1.57/t ex-pit). The unit cost of ex-pit material varies over the LOM as a function of the average length of haul. In the early years of operation, when the pit is relatively shallow so one-way haulage distances average less than 2 km and costs range from $1.20/t to $1.30/t. By the end of mine life, one-way haulage distances will exceed 8 km, resulting in unit costs exceeding $2.00/t. Figure shows how the average cost estimate varies with haulage distance over the LOM period. Figure LOM Mining Unit Costs and One-Way Haulage Distance 161