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An integrated underground mining and processing system for massive sulphide ores Bamber, Andrew Sherliker 2004

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A N INTEGRATED UNDERGROUND MINING A N D PROCESSING S Y S T E M FOR MASSIVE SULPHIDE ORES by A N D R E W SHERLIKER B A M B E R B.Sc. Eng (Mechanical) University of Cape Town 1993 Registered Professional Engineer (Engineering Council of South Africa) 1999 THESIS SUBMITTED IN PARTIAL FULFILMENT OF THE REQUIREMENTS FOR THE DEGREE OF M A S T E R OF APPLIED SCIENCE in THE F A C U L T Y OF G R A D U A T E STUDIES Department of Mining Engineering UNIVERSITY OF BRITISH C O L U M B I A November 2004 © Andrew Sherliker Bamber, 2004 A B S T R A C T A conceptual framework for the consideration of underground processing to improve the economics and environmental performance of underground hard rock mining has been developed from a previous phase of research at UBC. This research included identifying the opportunity for and benefits of underground waste rejection and disposal. Several enabling mineral processing technologies including electronic sorting were shortlisted for evaluation. This project represents the second phase of research into underground processing with INCO at UBC. The motivation for underground pre-concentration is discussed, and precedents for the concept are presented. Liberation criteria for evaluation of the amenability of an ore to underground pre-concentration have been identified. Fundamental design criteria and supporting technologies for underground pre-concentration are discussed, and the framework of an idealized integrated underground mining and processing system is presented. The work is considered both unique and valuable in that the contribution of the research is a feasible and practical new mining system that has been developed from previously vague and disparate concepts through the application of a combination of research and engineering skills. Using the conceptual framework thus developed, a case study into underground pre-concentration has been undertaken for the Main (MOB) and 153 Orebodies (153) at INCO's McCreedy East Mine in Sudbury, Ontario. Fieldwork and sampling, mineralogical evaluation, conceptual process design and costing have been completed, leading to the development of a system design which has been integrated into the existing mining scenario. Mineralogical results indicate that the liberation and separability characteristics of both ores as blasted is good: for the 153 ore, 55% of the tons mined could be rejected underground at a Cu recovery of 97%, resulting in a concentrate grade of 22% from a feed grade of 13% Cu. The system design comprises rejection of a coarse barren oversize at the orepass, transport of ore to the pre-concentrator by haul truck, pre-concentration of the ore through dense media separation and hoisting of the pre-concentrate to surface by hydraulic transport. For the MOB ore, 22% of the tons mined could be rejected underground through sorting at 97% Ni recovery, resulting in a concentrate grade of 2.93% from a feed grade of 2.44% Ni. Pre-concentration is achieved through combined conductivity sorting and magnetic separation. Pre-concentrated MOB ore would be hauled and hoisted to surface conventionally using the existing hoisting system. ii Waste products are coarse and competent and appear to be suitable for use as rockftll or as a source of aggregate for cemented fill. Substantial savings will be achieved through reduced underground haulage, hoisting, surface transport, milling and tailings disposal tonnages. Operating cost savings are of the order of 16% for the MOB and 24% in the case of the 153 orebody. In addition to this, further revenue could be generated through eliminating milling and smelting recovery losses in Cu, N i and precious metals from the 153 ore: the pre-concentrate appears to be of a quality suitable for introduction as feed to the smelter or matte converter. Results from the metallurgical testwork, process design, system layout, and a 30% capital and operating cost estimate are presented. A preliminary financial evaluation over the present life-of-mine, based on a total estimated system cost of CDN$30.8 million indicates an NPV of CDN$134 million, and an IRR of 79% at a discount rate of 11%. A brief evaluation of operational, economic and environmental impacts at the case study mine is also presented. A third phase of research and testwork is recommended comprising the following key objectives: - Bench scale testwork on alternative process technologies including coarse-particle flotation and autogenous grinding and classification - Evaluation of the rheology and pumpability of the pre-concentrate for the purposes of hydraulic hoist design - Research and testwork on the mechanical properties and rheology of a cemented backfill using pre-concentration rejects, and the modeling of the impact of introducing such fill on the geotechnical behaviour of an orebody at depth - Assessment of the long term stability of the process plant excavations in different stress and geotechnical situations - Further investigation of the comminution and liberation behaviour of the ores - Piloting of the conductivity sorting and dense media separation technologies on representative samples of the McCreedy ore - Further case studies on ores of different mineralogy and geotechnical properties Integration of the results of this third phase of research with the results presented in this thesis will provide a comprehensive tool for assessing the feasibility and impact of implementing underground pre-concentration for a range of mining scenarios presently considered uneconomic or technically unfeasible due to geotechnical considerations. iii Table of Contents Section Page Abstract ii Table of Contents iv List of Tables v List of Figures vii Acknowledgements x Chapter 1 - Background to the Research and Conceptual System Design 1.1 Introduction - Challenges in Underground Hardrock 1 Mining 1.2 Impacts of Implementing Underground Pre-concentration 10 on Mining 1.3 Design Criteria for an Integrated Underground Mining and 12 Processing System 1.4 Supporting Technologies for Underground Processing 17 1.4.1 Process Technologies 17 1.4.2 Backfill Technologies 27 1.4.3 Material Handling Technologies 31 1.5 Conclusion 37 Chapter 2 - Case Study: Development of an Integrated Underground Mining and Processing System for INCO's McCreedy East Mine 2.1 Introduction 38 2.2 Background 45 2.2.1 Geology and Mining 45 2.2.2 Milling and Smelting 51 2.3 Sampling and Mineralogical Testwork , 53 2.3.1 153 Ore Evaluation 54 2.3.2 MOB Ore Evaluation 62 2.3.3 Testwork Summary 69 2.4 Integrated Mining and Processing System Design 7i iv 2.4.1 Design Basis 71 2.4.2 Mining and Processing System Description 76 2.4.3 Waste Disposal and Backfill 88 2.4.4 System Integration 91 2.5 Conclusion 97 Chapter 3 - System Impact Evaluation 3.1 Introduction 98 3.2 Mining and Backfill 98 3.3 Underground and Surface Material Handling 100 3.4 Underground Mineral Processing 101 3.5 Surface Operations 102 3.6 Economic Evaluation 105 3.6.1 Cost Estimate 105 3.6.2 Operating Cost Evaluation 109 3.6.3 Financial Evaluation 110 3.7 Conclusion 114 Chapter 4 - Conclusions and Recommendations 115 References 117 Appendices 1. 153 Metallurgical Testwork Report 123 2. MOB Metallurgical Testwork Report 126 3. Previous INCO Testwork Data 129 4. Drawings 141 5. Equipment Lists 146 6. Mass Balances 161 7. Cost Estimate Spreadsheets 166 8. Financial Evaluation Spreadsheets 174 v List of Tables Description Page Table 1 - Metallurgical Performance of Pre-concentration 14 Table 2 - Advantages and Disadvantages of Surface Pre- 16 concentration Table 3 - Coarse Particle Flotation Operating Conditions 23 Table 4 - INCO Ore Sorting Testwork Results 39 Table 5 - Mining Rates, costs and values 47 Table 6 - Costs to Surface, all ores - McCreedy East Mine 51 Table 7-153 ore properties 61 Table 8 - Heavy Liquid Separation Results @ SG 3.0 61 Table 9 - MOB Ore Properties 68 Table 10 - Heavy liquid separation results @ SG 3.0 68 Table 11 - Evaluation of Coarse Particle Separation 71 Technologies Table 12 - Pre-concentration Plant Operating Design Criteria 75 Table 13 - Concentrate Specification Comparison 104 Table 14 - Cost Estimation Factors 106 Table 15 - Capital Cost Estimate Summary 107 Table 16 - Mining rates, grades, costs and values 110 Table 17 - Financial Evaluation Parameters 111 vi List of Figures Description Page Figure 1 - Idealized System Diagram for Underground Hard 4 Rock Mining with Fill Figure 2 - Integrated Underground Mining and Processing 12 System Model Figure 3 - Cost and Revenue Impacts of Pre-concentration 16 Figure 4 - Simplified McCreedy East Mine Section 41 Figure 5 - Combined underground pre-concentration facility 43 for McCreedy East Mine Figure 6 - Generalised McCreedy East Longitudinal Section 46 Figure 7 - McCreedy Mucking Circuit Flow Diagram 48 Figure 8 - Main Orebody Panel 49 Figure 9 - Typical 153 Narrow Vein Heading 50 Figure 10 - 153-4550 muckpile 55 Figure 11 - 153-4550 +53mm ore & waste 55 Figure 12 - 153-4550 +6mm fraction 56 Figure 13- 153 4550 +9mm sinks and floats 57 Figure 14-153 ore size distribution; whole ore fraction, sinks 58 and floats Figure 15-153 grade distribution by size 58 Figure 16- 153 4550 Ore fraction (x 50) showing chalcopyrite 59 with characteristic striations Figure 17 - 153 4550 middlings fraction (x 5) showing sharp 60 contact Figure 18- 153 4550 waste fraction (x 50) showing quartzite 60 embedded in Sudbury Breccia. Figure 19 - MOB round 62 Figure 20 - MOB 3575 +6mm fraction 63 Figure 21 - MOB 3575 +125mm fraction showing 64 chalcopyrite mineralization vii Description Page Figure 22 - MOB ore size distribution: whole ore fraction, 64 sinks and floats Figure 23 - MOB grade distribution by size 65 Figure 24 - MOB ore (x50) showing pentlandite, pyrrhotite, 66 silica and magnetite Figure 25 - MOB contact (x 50) showing dissemination of 67 sulphides into waste and waste into sulphides Figure 26 - MOB waste x 50 showing some disseminated 67 sulphides Figure 27 - Nickel Sulphide Ore Conductivity and Magnetic 72 Responses Figure 2 8 - 153 Mining Process Flow diagram 77 Figure 2 9 - 153 Pre-concentration Plant Process Flow 79 Diagram Figure 3 0 - 153 Concentrate Projected Size Analysis 80 Figure 31 - Concentrate Hydraulic Hoisting 81 Figure 32 - MOB Mining Process Flow Diagram 85 Figure 33 - MOB Pre-concentration Plant 87 Figure 34 - Expected Ore & Waste Fraction Size 89 Distributions Figure 35 - Backfill Preparation System 90 Figure 36 - Thomas Katts BS 907A Concrete pump schematic 91 Figure 37 - Simplified McCreedy section 92 Figure 38 - Location 8/9 - Plan and Elevation 93 Figure 3 9 - 153 Integrated Mining and Processing System 94 Layout Figure 40 - MOB Integrated Mining and Processing System 95 Layout Figure 41 - Simplified INCO Operations Flowsheet 99 viii Figure 42 - Mining and Milling Cost Comparison with and 109 without Pre-concentration Figure 42 - Underground Pre-concentration Capital Cost 113 Impacts ix Acknowledgements The support and guidance of Drs. Bern Klein, Mario Morin and Professor Malcolm Scoble of UBC Mining Engineering is gratefully acknowledged. I would like to extend my thanks to Simon Nickson at INCO Mines Technology for supporting the project, as well as Pascal Hamelin, Len Van Eyck and Steve Ball at McCreedy for their sponsorship and support. My thanks also to the beat geologists and planners: Rick LaCroix, Adrian Robertson, Todd McCracken, Cory Thomas, John MacLean and everyone else at Mines Technical Services; Rick Dean, Greg Krueger and the Logistics crew for showing me around the muck circuit and assisting with the sampling of the ores. Previous work by Mario Paventi and Harvey Buksa of INCO Mines Technology was instrumental in defining the scope and nature of the opportunity at McCreedy East, which is also gratefully acknowledged. x Chapter 1 - Background to Research and Conceptual System Design 1.1 Challenges in Underground Hard Rock Mining Canada's Mining industry has a history of over 100 years, and many national mines can be considered in a mature stage of development. Mining companies are operating in an environment of cyclically decreasing real commodity prices, whilst mining deposits of decreasing grade at increasingly greater depths (Scoble, 1994). In exacerbation of this, real operating costs continue to rise. Canadian mining companies compete globally for capital and markets, and as such are competing with mining operations around the world, many of whom operate in the lowest-cost quartile of the sector. The combination of decreasing commodity prices, increasing operating costs and diminishing in-situ metal grades place enormous pressure on the mining industry to innovate or face extinction (NRCan, 1999). Ore deposits are becoming simultaneously scarcer and deeper. This situation is further exacerbated by increasingly strict environmental legislation, resulting in an increasing cost of meeting statutory requirements. As mining operations increase tonnage to maintain revenue targets in the face of falling grades, there is extreme environmental pressure to utilize less power and water, and to reduce surface and underground environmental impacts, including reducing negative impacts on groundwater, and reducing the physical footprint of mines, particularly surface infrastructure, waste rock dumps, tailings dams and slag heaps (Scoble et al, 2000). The sustainability of mining has been linked to improving environmental and economic performance through innovation (Parsons & Hume, 1997). Despite a tradition of technological conservatism, innovations for improved environmental sustainability in the industry has been demonstrated in the development of technologies such as the ISASmelt Process, Outokumpu's Flash Smelting and Converting technology, and recent advances in flash smelting by INCO at Copper Cliff in Sudbury (Warhurst & Bridge, 1996). The trend towards increasing the mining rate to enhance economics has led to enormous environmental pressure to reduce the surface footprint of mines during the operational phase, as well as 1 increasing the attention paid to post-closure issues in mining (Feasby & Tremblay 1995). Recent papers presented at the International Congress of Mineral Processing Engineers have highlight this trend, outlining a desire to move away from large volume surface mining and towards 'invisible' low-cost mining techniques. Substantial and sustainable innovation is required to realize the goal of invisible mining, including improved directional drilling technology and knowledge of geomechanics as well as in-situ mining techniques such as in-situ leaching (Batterham, 2003). A safe, reliable, low-cost, high-productivity, well-supported mining method for ultra deep-low grade orebodies that is environmentally superior to present high-tonnage techniques has not yet been developed. The successful development of underground pre-concentration is also considered a critical technology in the pursuit of this goal. A decreasing number of economic mineral deposits are found at shallow to medium depth, and most of those shallow deposits that remain are close to sub-economic grade and are required to be mined at high tonnages in order to show a return. For the purposes of this thesis, we can define mining depth as follows: • Shallow <500m • 500 < Medium depth < 1500m • 1500 < Deep < 3000m • Ultra deep >3000m A large proportion of the remaining deposits are higher grade deposits which are presented in situations of either extreme depth, complex structure or poor ground conditions, one or all of which factors make the deposit sub-economic. South Africa is currently mining among the deepest orebodies in the world, and some of the critical issues have been identified in relation to deep-level gold mining in this context (Lloyd et al, 1986; Tamlyn 1994). Key issues with respect to mining at extreme depth are increased rock stress and high virgin rock temperatures. The aspect of heat at depth can be exacerbated by the simultaneous occurrence of high geothermal gradients at the site. High rock stress can be managed by a combination of pre-stressing mining excavations, additional engineered support, the use of backfilling techniques, and the adoption of methods that result in a low post-mining stress such as longwall back-caving. Time-dependent strain behaviour of the rock mass at extreme depth, 2 w h i l e a m e l i o r a t i n g p o t e n t i a l r o c k b u r s t s d u e t o the r e l a x a t i o n o f p o s t - m i n i n g s t resses o v e r t i m e , i s a p r o b l e m f o r c o n v e n t i o n a l e n g i n e e r e d s u p p o r t m e t h o d s a n d c a n o n l y b e a d d r e s s e d t h r o u g h b a c k f i l l i n g o r c a v i n g m e t h o d s ( B o s m a n , 1 9 9 5 ; R e t i e f , 1 9 9 6 ) . T h e m i n i n g o f l o w g r a d e , o r a l t e r n a t i v e l y h i g h l y d i l u t e d o r e at d e p t h r e s u l t s i n e x t r e m e p r e s s u r e o n m i n i n g c o s t s . G r a d e c o n t r o l a n d the m a x i m i z i n g o f the g r a d e o f o r e d e l i v e r e d to s u r f a c e i s o f p a r a m o u n t i m p o r t a n c e . A c o m m o n s o l u t i o n t o i m p r o v i n g the e c o n o m i c s o f m i n i n g l o w g r a d e o r e b o d i e s i s to i n c r e a s e the m i n i n g ra te , w i t h c o n s e q u e n t d e c r e a s e s i n u n i t co s t s a n d a n i n c r e a s e i n c a p i t a l cos t s . I n c r e a s i n g the m i n i n g ra te c a n a l s o e x a c e r b a t e e x i s t i n g r o c k b u r s t i n g p r o b l e m s . A l t e r n a t i v e m e t h o d s f o r the m i n i n g o f n a r r o w - s e a m o r e b o d i e s w h o s e g r a d e i s d e c r e a s e d b y d i l u t i o n w i t h b a r r e n m i n i n g w a s t e i n c l u d e s e l e c t i v e m i n i n g , s e l e c t i v e b l a s t m i n i n g a n d r e sue m i n i n g ( B o c k et a l , 1 9 9 8 ; P i c k e r i n g et a l , 1 9 9 9 ) . H o w e v e r , the a d o p t i o n o f s u c h m e t h o d s c a n l e a d to a d e c r e a s e i n m i n i n g p r o d u c t i v i t y , a n d t h u s i n c r e a s e d c o s t s w h e n c o m p a r e d t o b u l k m i n i n g t e c h n i q u e s . G r o u n d b r e a k i n g t e c h n o l o g y i s r e q u i r e d t o o v e r c o m e the c h a l l e n g e s o f m i n i n g s u c h d e p o s i t s , w h i c h i n c l u d e i n t e r a l i a : • t he m a i n t e n a n c e o f e x c a v a t i o n a n d s tope i n t e g r i t y i n h i g h l y s t r e s s e d a n d p l a s t i c r o c k e n v i r o n m e n t s • the l o g i s t i c s o f h a n d l i n g l a r g e q u a n t i t i e s o f l a b o u r , c o n s u m a b l e s , o r e a n d w a s t e o v e r l a r g e d i s t a n c e s at g rea t d e p t h • the m a i n t e n a n c e o f a d e q u a t e v e n t i l a t i o n , c o o l i n g a n d o t h e r m i n e s e r v i c e s at d e p t h o v e r the ex t en t s o f a l a r g e u n d e r g r o u n d m i n e • the e f f e c t i v e m a n a g e m e n t o f o p e r a t i o n s a n d l a b o u r a c r o s s a h i g h l y d i s t r i b u t e d p r o d u c t i o n s y s t e m • n o n - e x p l o s i v e r o c k - b r e a k i n g a n d c o n t i n u o u s m i n i n g s y s t e m s T h e c o m p o u n d e d p r o b l e m o f e x t r e m e d e p t h a n d l o w g r a d e t h u s r e q u i r e s a d v a n c e d t e c h n i q u e s i n o r d e r to i m p r o v e the e c o n o m i c s o f m i n i n g to the p o i n t o f f e a s i b i l i t y . It h a s b e e n fu r t he r s u g g e s t e d tha t the e c o n o m i c s as w e l l as the e n v i r o n m e n t a l a c c e p t a b i l i t y o f m i n i n g l o w - g r a d e o r e b o d i e s w o u l d b e i m p r o v e d b y r e d u c i n g the q u a n t i t y o f o re to b e t r a n s p o r t e d , h o i s t e d a n d p r o c e s s e d o n s u r f a c e t h r o u g h i m p r o v e d w a s t e r e j e c t i o n ( F e a s b y & T r e m b l a y , 1 9 9 5 ) . T h e t e c h n i c a l a n d e c o n o m i c f e a s i b i l i t y o f the ' i n t e g r a t e d m i n i n g a n d p r o c e s s i n g c o n c e p t h a s b e e n 3 partially examined in a number of studies to date (Lloyd, 1989; Peters et al, 1999). Consideration of underground pre-concentration addresses three of the five critical research areas mentioned above. In order to facilitate the development of this concept, a generic model of an underground hardrock mining system can be defined, and is presented in Figure 1. /"CEMENT. \ WATER MINE SAND FOR FILL CLASSIFIED TAILINGS FOR FILL TAILINGS DISPOSAL PREPARE & DELIVER BACKFILL WASTE ROCK DUMP SURFACE TRANSPORT SURFACE MILL SMELT < METAL SALES UNDERGROUND RESUE WASTE TO FILL DRILL, BLAST, MUCK, BACKFILL LOAD, HAUL, DUMP HAUL/ CONVEY HOIST ORE TO SURFACE (^ OREBODY Figure 1 - Idealized System Diagram for Underground Hard Rock Mining with Fill There are several precedents for the installation of mineral processing technology underground. Mining at extreme depth results in additional requirements for support infrastructure underground: hoisting, ventilation, pumping, material handling systems and comminution systems are common underground, all of which require reliable excavations and additional non-mining labour, power and materials. Excavations and support for process facilities underground is not considered an undue extension of this precedent; compared to these instances, the installation of process plant underground introduces a value-added step which decreases the volume and increases the value of the ore. Increasing depth also results in an exponential increase in the hoisting capacity required, resulting in increased mining cost 4 per ton, which substantially increases the pay-limit required of the ore to be mined at these depths. Reducing the quantity and increasing the value of ore delivered to surface can materially reduce all costs downstream of the reduction. Although examples of underground process plants that have been constructed are not common, there are a number of precedents for the construction of process plants underground. At CODELCO's Andina Mine, a 30 000 tpd copper flotation plant was constructed in a series of near-surface excavations high in the Chilean Andes. The location of the plant was driven by reasons of the extremely harsh climate at that altitude, as well as steep and irregular topography which would have made surface construction impossible (Brewis, 1995). Further to this example, Cameco Corporation have commissioned a 300 tpd automated crushing, milling, and hydraulic hoisting facility underground at their McArthur River operation in Saskatchewan (Edwards, 2004). Operator safety was the primary motivation for this decision, due to the extremely high grade (27% L^Og), and thus radioactivity of the ore. McArthur River have pioneered hydraulic hoisting of the yellowcake slurry thus produced in order to overcome the shortcomings of conventional hoisting in the handling of such high grade uranium. Another example of the application of process technology underground can be found at Kiruna Iron-ore Mine in Sweden, where sorting by Laser-Induced Fluorescence (LIF) is used to separate high-phosphorous from low-phosphorous ore (Kruuka & Briocher, 2002). Similarly, at Stobie Mine near Sudbury, Ontario, LNCO currently employs conductivity sensors to monitor the sulphide grade of ore delivered to surface -. while not strictly pre-concentration, this is a further example of an application of mineral processing technology in the underground environment. The simultaneous handling of valuable ore and barren waste is a problem faced by the majority of mines around the world, regardless of the type or grade of mineral to be exploited. There are four disparate sources of waste arising in underground mining operations: • Hangingwall or footwall waste which is blasted simultaneously with the ore during stoping operations. This is usually referred to as planned dilution. • Footwall or hangingwall waste generated in off-seam development; this waste often falsely reports to the ore stream during the mining cycle. 5 • Sloughing of the hangingwall or sidewall faces during stoping or in the orepass. • Interstitial waste within the ore, either as gangue material separating massive mineralisation, or gangue associated with disseminated mineralisation. It is this last type of waste which is targeted for removal in typical grinding and flotation circuits. Dilution of the head grade when compared to the in-situ ore grade occurs through a loss in high-value fines from the ore, as well as additional unplanned dilution from development waste; furthermore, development waste can be up to 50% of the total mass of material hoisted to surface in a typical mining operation. In the light of this, the opportunity to leave waste underground is attractive, but must be considered simultaneously with suitable mineral processing techniques for the efficient identification of waste, and the integration with a suitable mechanized mining method capable of accepting large quantities of backfill. Due to an approximate 40% increase in the volume of rock subsequent to fragmentation, it is estimated that only 60% of rock by volume can be returned to the mining void, excluding allowances for water, fines and binder for CRP. This estimate excludes development waste, which must by necessity be transported to surface. Reducing the quantity of waste hoisted to surface will increase the capacity of an existing shaft for the hoisting of ore, men and materials. The South African Chamber of Mines Research Organization (COMRO) investigated the potential of an integrated mining and processing system for deep gold mines in some detail between 1978 and 1989 (Lloyd, 1978, 1979; Lloyd et al, 1986). It was concluded that suitable cemented backfill could be generated economically underground using coarse particle flotation on the gold ore, producing 60% rejects at a topsize of 3mm and a recovery of 98%). A relatively coarse waste particle size was desired in order to maximize the strength of the cemented backfill, resulting in an expected uniaxial compressive strength (UCS) of 10 - 20 MPa with the addition of 30% cement by mass. Lloyd concluded that using such an integrated mining and processing system would result in an 85%) improvement in ground control at depth due to superior backfill properties when compared to cemented mill tailings, and a consequential reduction in the requirement for engineered support, reduced ventilation requirements. Furthermore, there would be an increase in hoisting capacity, and an 6 opportunity to hoist the concentrated ore using hydraulic hoisting. The improvement in the quality of cemented backfill and thus improved post-mining support was predicted to lead to a decrease in uncontrolled rockbursts, as well as decreased ventilation requirements post-fill, and potentially increases the safe depth at which mining can occur. There are positive economic impacts of underground pre-concentration in addition to the geotechnical impacts of producing a high quality backfill directly underground. A brief comparison of the economics of producing 50% backfill from underground waste, versus hoisting the waste to surface for a 240 000 tpm gold mine, mining an average grade of 6,6 g/t indicates potential operating and capital cost savings of 40% (Hinde et al, 1986). Further to this work, a conceptual underground pre-concentration study was previously undertaken for an existing base metal operation, grading at 2.5% Cu, 0.5% Zn and 30g/t A u , mining 2Mtpa at a depth of 6000 ft (Peters et al 1999). The mining method was considered suitable for backfill, and the integration of a Dense Media Separation (DMS) plant at the shaft bottom was proposed, rejecting 55% of the feed at a planned recovery of 96%. Capital and operating costs were estimated for the underground pre-concentration plant: the project indicated an operating cost saving of 11%, and a N P V of $6 550 000 per annum with an IRR of 23% at a discount rate of 15%. Further savings in capital and operating costs due to the reduction in required capacity of hoisting plant, surface transport and the surface mill and tailings dam were not taken into account. Further work into mine-mill integration and in particular integrated mining and processing systems has been conducted at U B C between 2000 and the present. A focus of the research to date has been on estimating the impact of improved waste management through the rejection of waste underground on the continuing viability and sustainability of Canadian underground hard rock mining (Klein et al, 2000, 2003). Advanced mining systems under consideration in this body of research include improved 'drill to mil l ' geological information systems, blast fragmentation design and control, minimization of waste dilution through improved selectivity, and mine-mill integration. Further areas of focus include the rejection and storage of waste rock in the underground void, thus minimizing the quantity of rock hoisted to surface, leading to a reduction of the surface footprint of underground mines. Maximizing the 7 underground rejection of waste and the minimization of surface waste disposal are seen as paramount in ensuring the sustainability of Canadian mining. Underground pre-concentration automates the waste sorting process at a location away from the face, and with suitable surge before and after the process, as well as the option to bypass the concentrator if required, can increase the productivity of this process. Due to the increased productivity of the mining method, and the decrease in quantity of ore transported to surface for processing, operating costs can be significantly reduced. Capital requirements are also reduced as downstream facilities such as hoisting, surface processing, and tailings disposal operate at a significantly lower throughput, with a commensurate decrease in capital expenditure. There is however, a notable increase in underground capital expenditure due to the additional process plant, and backfill facility required, as well as a marginal increase in tons mined for the same quantity of metal produced due to the additional recovery penalty. Previous research work at UBC, concluded the following positive impacts of underground pre-concentration (Peters et al,1999; Scoble et al 2000): • Reduction of surface material handling costs, especially for geographically diverse operations • Reduction in backfill unit costs due to an increase in quantity of backfill placed and the close proximity of the source of backfill to the mining void when compared to preparing backfill from surface • Reduction in tailings disposal on surface • Potential lowering of the cutoff grade, and a commensurate increase in the mining reserve • Facilitation of lower unit cost mining methods due to a reduction in mining selectivity, and increase in throughput and a decreased requirement for engineered support due to the availability of cheap, competent backfill • Facilitation of alternative ore material handling to surface such as hydraulic hoisting • Increased environmental compliance and public acceptance of mining through reduced physical surface impacts 8 However, there are numerous technical challenges to the implementation of underground processing, many of which relate principally to the integration of conventional process technology into the underground environment. These challenges include: • The requirement for large, stable, capital-intensive underground excavations: in addition to the excavations required for the process plant itself, additional footwall development must be considered in order to accommodate the inbound ore material handling from the stopes to the pre-concentration plant, and outbound material handling of waste from the pre-concentration plant to the backfill plant, and the transport of the pre-concentrated ore from the plant to surface. Shafts and haulages must also be large enough to accomodate the process equipment during construction. In the context of ultra deep mining, the challenge of constructing large (d>50m) excavations in highly-stressed or plastic rock conditions must thus be overcome. • Inbound and outbound material handling: all production tonnage must pass through the pre-concentration plant. This places restrictions on the number and location of the pre-concentration facility(s) in relation to the producing stopes; locations must be feasible over the life of mine under consideration, and the location of the pre-concentration facilities with respect to the vertical shaft or decline leading to surface must be considered. The bypassing of the pre-concentration plant must be considered to ensure mining productivity is maintained in the face of plant breakdowns. Additional challenges are presented in the logistics of delivering reagents, spares and consumables from surface to the pre-concentration plant. • Plant design, equipment design and layout: The design of the underground processing plant will be substantially affected by the underground environment. In addition to the constraint of space, the underground environment will present conditions of heat, humidity, noise, and dust not experienced on surface. Both the underground pre-concentration plant and the mine ventilation system must be designed to cater for these conditions. Such conditions will affect issues such as materials of construction, bearing, gearbox and lubricant selection, electrical and control system specifications (especially in the case of fiery mines), and operating and maintenance procedures. Layout with regard to the vertical and horizontal relationship between crushing and screening operations, chutes and bins for example must be carefully considered in 9 order to minimize the requirement for excavations. Recirculating loads should similarly be minimized or eliminated in the design, and space- and energy-intensive mineral processes should be considered. It has been suggested that dry processes would be preferred to wet, and the selection of wet processes in the underground environment must take due consideration of the delivery of process water, containment, and the removal of effluent streams from the process. Again, the logistics and safety of utilizing common process reagents in the underground environment must be considered. It should be stressed here that such an integrated underground mining and processing system must also meet the usual design criteria for process plants arising from geology, mineralogy, orebody depth, geometry and orientation, mining method, physical ore properties (density, hardness, abrasion, Bond Work Index, angle of repose), liberation characteristics, and Acid Rock Drainage (ARD) potential. 1.2 Impact of Underground Pre-concentration on Mining Due to the requirement to maximize the positive impact of underground mineral processing by disposing of the maximum possible quantity of waste as backfill, only mining methods which can accommodate backfill will be considered for the integrated system. Mining methods which accommodate the greatest proportion of backfill are preferred. In the light of this the following mining methods are ranked in order of preference for integration with underground pre-concentration: 1. Bulk mining via mechanized cut-and-fill 2. Post-pillar cut-and-fill 3. Drift and fill 4. Room-and-pillar mining 5. Avoca, longhole and longwall stoping with backfill (narrow, steeply dipping) In the consideration of optimal underground mining methods, it is important to take into account the ratio of stoping to development associated with the method. The practicality and 10 benefit of underground pre-concentration would be greatly reduced for mining methods with a low stoping:development ratio, as the bulk of development waste by necessity still requires hoisting to surface. It can be concluded that the optimum integrated mining and processing method for deep underground mines would be a combination of vertical hoisting from the level of the upper orebody horizon, combined with either mechanized overhand cut-and-fill (Pfarr, 1991) or mechanized post-pillar cut-and-fill. Both these methods offer a significantly lower unit mining cost and lower engineered support costs when compared to selective mining methods, such as longwall or drift and fill. Efforts have been made previously to increase the productivity of deep, narrow-vein mining while also minimizing waste dilution. Randfontein Estates in South Africa experimented with thick-seam mechanized gold mining ( ' T M 3 ) in the 1980's. However the resultant decrease in head grade was considered economically unacceptable. A mechanized alternative to the resue mining of steeply-dipping narrow vein gold deposits of the Witwatersrand has been proposed by others (Pickering et al, 1999). A similar mining method has been adopted at Placer's South Deep in some mining sections, with interesting results. The planned grade of ore delivered from the thick seam mechanized sections is some 30% lower than from the conventional longwall mining sections, which can be attributed solely to additional dilution arising from the mining method*. However, the unit cost of the bulk method is 30% lower than the conventional method, and is therefore still competitive. Highly selective or resue mining methods can increase the viability of mining such an ore. Selective blast mining methods (SBM) have been suggested where ore and waste are blasted simultaneously, but are segregated due to a blast design which 'throws' the ore into a different area of the stope (Bock, 1998). Mining methods such as S B M , resue mining and the physical sorting of waste underground increases the availability of waste for backfill with a source close to the stope, reduces the quantity and increases the quality of ore delivered to surface. However, adopting any of the abovementioned methods in conjunction with the principal mining method can increase the delay in face cleaning operations, and hence impact negatively on the productivity of the method. *\PDG Placer Dome operations South Deep\Western Areas Limited - Annual Report 2002 files\assets main.htm 11 1.3 Design Criteria for Underground Pre-concentration The pre-concentration of ore to reduce the quantity and improve the quality of ore is not new. As early as 1936, HMS pre-concentration was implemented at the American Zinc Company in Mascot Tennessee in order to improve the process performance and economics of the operation. In the 195O's it was identified that pre-concentration of ore at the Sullivan Mine in British Columbia could substantially improve the economics of the downstream milling and flotation process. In 1979, Heavy Medium pre-concentration was piloted at Mt Isa Mine in Australia, and eventually a full scale plant was built and commissioned, with significant process and economic improvements in the downstream plant (Munro et al, 1982). Construction costs of the DMS pre-concentration plant were AUS$26m (base 1979), resulting in a capacity increase of 50% to 3.6Mtpa Run-of-Mine (ROM). Metallurgical results are presented in Table 1. There were other benefits to pre-concentration. Mt Isa also observed a reduction of 25% in effective Bond Work Index of the pre-concentrate due to the rejection of the mostly siliceous waste, and facilitation of a finer grind to 87% - 74um in the milling circuit, with attendant improvements in the liberation and recovery of lead and zinc, all of which contributed to the increase in production capacity. Pre-concentration by DMS is common in Bushveld base- and precious metal operations such as Karee and Kroondal Platinum, where due to the narrow (<800mm) and partitioned nature of the UG2 chrome- and platinum bearing seams, up to 30% waste can be included in the mining cut, which decreases the average R O M grade to below the economic cutoff grade in many cases. DMS pre-concentration subsequent to a suitable feed preparation stage is employed on surface in combination with flash flotation for the recovery of PGM's in the fine DMS bypass. The grade of the ore is increased to above the economic cutoff, and tonnage to the grinding and flotation stages is substantially reduced, leading to lower surface plant capital and improved P G M recoveries overall (Lawrence et al, 1996). Several other operations using pre-concentration have been identified in the literature. There is ample evidence to suggest that the introduction of the pre-concentration step underground would bring further economic and environmental benefits to an operation. 12 A diagram of an idealized underground mining and processing system showing the integration of the mineral processing step into the mining cycle is presented in Figure 2. This system model can be compared to the underground mining system presented in Figure 1. CEMENT, SAND PREPARE TAILINGS FOR BACKFILL TAILINGS DISPOSAL WASTE ROCK DUMP SURFACE TRANSPORT SURFACE MILL SMELT METAL SALES PREPARE & DELIVER BACKFILL „ WASTF i m m m UNDERGROUND BY-PASS MILL? REJECT COARSE WASTE DRILL, BLAST, MUCK, BACKFILL LOAD, HAUL, DUMP PRE-CONCENTRATE HOIST CONCENTRATE TO SURFACE (OREBODY HAUL/ CONVEY ngure 2 - Integrated Underground Mining and Processing System Model The economic feasibility of underground pre-concentration will be determined primarily by the proportion of waste that can be rejected and the metallurgical performance of the process technologies employed. Underground pre-concentration must by needs utilize coarse-particle separation technologies. Suitable coarse particle processes include sorting by means of visual, conductivity, radiometric or X-ray fluorescent means, as well as gravity separation and dense-media technologies (Sivamohan & Forssberg, 1991). Coarse-particle flotation has also been suggested previously as a suitable pre-concentration technique. Such processes are common in surface pre-concentration on chrome, platinum and base metal oxide and sulphide ores. The advantages of surface pre-concentration are, by now, well established and the technology is being included in many current plant designs. Mountain States R & D have conducted bench and pilot scale metallurgical testwork on a range of copper ores, demonstrating that in the 13 case of relatively coarse copper porphyry mineralization, pre-concentration by DMS is effective in improving the grade of the copper ore at an acceptable recovery (McCullough et al, 1999). Mineralization at Copper Creek is principally disseminated chalcopyrite, massive chalcopyrite, with minor mineralization of bornite and chalcopyrite in sericitic dolerite, with a liberation size of 13mm. Metallurgical testwork indicated a good potential for HMS and the flowsheet was piloted. Capital costs for the HMS / flotation concentrator were 50% lower than direct grinding and flotation, and operating costs were estimated at $3.50/t vs. $6.50/t for direct flotation. The Phase I design for the Copper Creek project includes for the pre-concentration of 10 000 tpd ore on surface by HMS, prior to grinding and flotation. In Phase II it is proposed to install a HMS plant underground in order to concentrate the ore prior to hoisting (AMT Annual Report, 1997). Rejects from the HMS pre-concentration step are to be crushed and prepared as backfill for the proposed mechanized cut-and-fill mining method. Metallurgical results from the literature reviewed are summarized in Table 1. Table 1: Metallurgical Performance of Pre-concentration Operation Ore type Pre-concentration method Feed grade % %Waste rejected Cone grade Recovery American Zinc Mascot, Tennessee Zn sulphide Gravity concentration /HMS Cone separator 3% Zn 55% 10-12% Zn 96.5% Mt Isa Mines, Aus Pb-Zn sulphide DMS cyclone 2.5% Cu 5.6% Zn 5.9% Pb 33% 3.58% Cu 8.02% Zn 8.45% Pb 96% Masua Mine, Sardinia Pb-Zn sulphide 2-stage Dynawhirlpool 2.0% Pb 2.8% Zn 74.03% 7.32% Pb 8.91% Zn 94.48% 80.65% C.N.R. Laboratory* Sb oxide/ sulphide ore Shaking table/ Dynawhirlpool 3.19% Sb 90.41% 29.62% Sb 89% AMT Copper, Arizona* Copper porphyry DMS 0.75% Cu 75% 2.85% Cu 95% Witwatersrand* Gold conglomerate Coarse flotation 4g/t 60% 6.53 g/t 98% Potash Corp Fluorite/ barite Coarse flotation 9% 28% 11.79% 97% **Sukinda ultramafic belt Nickel laterite Hindered settling, classification 0.8% Ni 27% 1.1% 89% Ferrara & Guarascio, 1980 ** underground pre-concentration testwork ***Mohanty et al, 2000 14 The metallurgical results presented above indicate conclusively that a high degree of waste rejection is possible at a coarse particle size with good metallurgical recoveries on base metal sulphide ores using dense media technology. The technology is high capacity, compact and cheap, especially when compared to grinding and flotation plant of a similar throughput. The additional economic advantages of underground pre-concentration are noteworthy. COMRO's original research into an integrated underground mining and processing system was motivated principally by the need to produce a cheap backfill material with superior mechanical properties underground in order to address the rock mechanics challenges being experienced in deep mining at the time (Lloyd, 1979). Economic considerations for the technology were originally secondary. Several developments have arisen from this work to date, leading to advances in coarse particle flotation, high-energy milling and advances in backfill technology. In the course of the work at COMRO, basic design criteria for the design and construction of mineral processing facilities underground were identified: • Plant should be as compact as possible • Processes should be robust in the face of variations of feed grade and tonnage • Open circuit processes are preferred to closed-circuit • Recovery should be maximized in favour of grade • Waste products should be suitable for backfill The construction of an underground mineral processing plant satisfying these basic design criteria was proposed, comprising crushing, centrifugal milling and coarse flotation, with scavenging of flotation tails by means of a classifying cyclone to maximize recovery. A maximum mass rejection of 60% was targeted in order to optimize backfill requirements, consequently a comparatively low grade of the pre-concentrate was produced at an extremely high recovery. A waste product size of -1mm + 40um was targeted, which was considered suitable for combination with OPC as a paste fill. Capital and operating costs were not examined. Additional design criteria also must be considered in the underground environment. Due to the continuous nature of the process plant, material handling from the pre-concentration plant 15 to the shaft, and from the tip or orepass should also preferably be continuous. In summarizing these criteria, it is suggested that the ideal underground processing plant would target a maximum of 60% waste rejection at a recovery of better than 95% using compact and or energy intensive dry unit processes in open circuit, at particle sizes between 1 and 100mm. Sivamohan and Forssberg (1991) identified a number of the advantages and disadvantages of coarse particle pre-concentration, which are summarized in Table 2. Table 2 - Advantages and Disadvantages of Surface Pre-concentration (after Sivamohan & Forssberg, 1991^ l: Advantages Disadvantages Overall operating cost savings Additional capital and operating cost of pre-concentration Savings in hoisting and surface transport Additional comminution costs to prepare coarse feed Obviation of low productivity mining techniques Recovery penalty across the pre-concentrator Reduction in size of the surface concentrator Not applicable to highly disseminated ores Improvement in the economic cutoff grade through operating and capital cost savings Review of other sources confirms these observations. A comparison of the operating cost and revenue impacts arising from the introduction of DMS prior to a barite/fluorite flotation operation indicates the following impacts on costs and revenues (Schena et al, 1990): Cost and Revenue Impacts of Pre-concentration 25i _ _ _ 1 -J5- JI— • — :I— I— I H _MJM B— Wt- I— Wt- • Revenue va.uem 1 0 r r r I L L L L L L L •cos' I 1 1 1 I I I I I I I • Margin 10 12 14 20 22 24 26 30 32 34 36 % Waste Rejection Figure 3 - Cost and Revenue Impacts of Pre-concentration 16 The savings margin was a maximum at a cut density of 2.75, equating to a mass rejection of 26%. Revenue impacts at this level of waste rejection were estimated at 3%. It can be concluded that there are substantial economic benefits to employing pre-concentration prior to further mineral processing, even when offset against the value of the additional mineral losses. Research into underground processing undertaken by Klein et al (2000) concluded that underground pre-concentration should be implemented as close to the mining face as possible, and at as coarse a particle size as possible. In addition to this it was concluded that dry processes were preferred over wet. We are thus able to rank the available mineral processing technologies (Best Available Technology) on the basis of the size of equipment required, topsize of the feed processed, and whether the process is dry or wet. The constructability and operability of the selected processes in the underground environment must also be taken into account in the final design, as well as the appropriate proximity of the process to the mining face. In the selection of unit processes, observation must also be made of the typical particle size and grade distribution at various stages in the underground mining system, as well as the typical liberation and separability characteristics of the ore, which will be discussed in a later section. Should the appropriate particle size distribution for a particular unit process not occur naturally during blasting, mucking, tipping and hauling operations, comminution steps such as crushing and grinding must also be considered for application in the pre-concentration process. 1.4 Supporting Technologies for Underground Pre-concentration 1.4.1 Process Technologies Technologies for coarse ore separation are numerous, and exploit one or a number of physical properties of the ore. The most common separation technologies and the property exploited are listed in Table 3. Differential hardness of the ore vs. waste may also be exploited as a means to separate out the ore. In addition to these processes, recent advances in coarse particle flotation, particularly flash-flotation as well as the Jameson Cell make flotation an attractive unit process for the underground pre-concentration application. Separation by 17 flotation and froth separation (SIF) has been successful up to a topsize of 3mm in barite and fluorite operations (Leppinen et al, 2000). With further research, similar separation sizes may be possible for sulphide ores. Table 3 - Typical Criteria for Discerning Ore from Waste Property Technology Colour / Reflectivity Optical / Colorimetric Sorting Conductivity Conductivity Sorter Fluorescence X-Ray / Laser / UV Sorting Permeability (n) X-ray Radioactivity Radiometric Sorting Gamma Neutron Activation Magnetism / Paramagnetism Magnetic Separators Temperature Differential microwave heating & sensing Density Gravity Separation Dense Media Separation Size Scalping / screening The amenability of an ore for these separation methods must be determined in each case through suitable process mineralogical techniques, firstly by analyzing the liberation characteristics, and then exposing the liberated ore to the various detection/separation methods at the bench scale. Should an ore show amenability, then further testwork and piloting would be recommended. It is the focus of this thesis to limit the process technologies covered in the development of the pre-concentration flowsheet to the following: • Scalping • Coarse Ore Sorting • Gravity Separation • Dense Media Separation 18 • Coarse Particle Flotation • Crushing and grinding Scalping Coarse particle screening or scalping can be utilized in order to separate ore from waste in cases where there is a clear particle size / grade distribution correlation. Scalping is compact, cheap, efficient and easy to operate in the underground environment. As an example of this, Kroondal Platinum Mine in the South African Bushveld has implemented coarse waste scalping at the tipping points ahead of the section strike belts. Scalped waste is stowed in the mined-out bords either side of the belt. Previous research into underground pre-concentration at McCreedy East identified a clear size/grade relationship in the R O M ore, and it may be feasible to scalp up to 15% of the R O M ore as waste at the tipping points (Buksa & Paventi, 2002). Depending on the liberation characteristics of the ore, it may also be possible to identify a secondary particle size at which a further barren fraction can be identified and scalped out. Under the category of scalping, in the case where a particular size fraction can be identified as sufficiently high-grade, the inverse of this particle size / grade relationship can also be exploited, and the high grade fraction simply screened out. Again, this appears to be the case in the McCreedy ore, where, due to the friability of the high-grade sulphides, the majority of sulphide values are present in the -15mm fraction and can report directly to concentrate. In this case, the feed to the downstream pre-concentration plant is a middlings fraction, with both coarse barren waste, and high-grade fines having been screened out ahead of the pre-concentrator. This is a common occurrence in many ores, and could be exploited in order to achieve cheap and simple pre-concentration at an acceptable metallurgical recovery. This propensity towards a higher metal grade in the fines fraction can be promoted through appropriate comminution steps. 19 Electronic Sorting Electronic sorting encompasses optical, fluorescent, conductivity and gamma-neutron activation detection (GNA) sorting methods. Sorting can be achieved by one or a combination of methods, such as a combination optical / conductivity sorter. Optical sorters exploit the difference in colour or reflectivity of a high grade ore when compared to a low-grade ore. Sulphide ores often exhibit colour, reflectivity or iridescence discrepancies when compared to barren waste. Fluorescent sorters induce fluorescence in a mineral by excitation with either an X-ray, laser or U V energy source. Differential fluorescent responses between gangue and ore can be used to separate high grade from low-grade ore. Conductivity sorters typically measure the conductivity of an ore by eddy current detection -metallic or high-grade sulphide minerals cause a substantial distortion in the eddy current field when compared to low-grade material; alternatively the conductivity of the ore can be measured by direct contact. Gamma neutron bombardment is used in G N A detectors to induce temporary radioactivity in the different elements comprising an ore, and can be used to return an elemental analysis of an ore, which can be used as an acceptance or rejection criteria. The mechanism of sorting can be either mechanical, pneumatic or hydraulic. Sorters may reject either ore or waste, depending on the relative proportion of each; particles are ejected from the ore stream by means of flapper plates, high-pressure air- or water jets, for example the rejection of waste by flappers in the low-grade gold sorting facility at Kambalda Mine in Australia. Whole stream sorters divert the entire ore flow to waste i f a property is considered undesirable, as in the case of the LIF system at Kiruna Mine. Sorter performance is dependent on both the tonnage and particle size distribution of the ore reporting to the sorter. Sorter efficiencies drop markedly as tonnages reach the upper limit of the sorter capacity. Sorters also operate best over a size range of not more than 3:1. Hence the sorter must operate in conjunction with a suitably designed crushing and screening plant in order to prepare correctly sized ore streams at an acceptable throughput for the sorter. With 20 particular reference to photometric sorters, sorting efficiency is further affected by the orientation and spacing of the individual rocks, and the resolution and speed of the optical scanning and analysis system (Schapper, 1975). Principal manufacturers of optical and conductivity sorters include Ore Sorters (Australia), with Optosort and Mogenson based in Germany. Recent developments with Ore Sorters Model 13 and 16 optical sorters indicated good sorting results on 20mm particles at throughputs up to 75tph and at belt speeds up to 4m/s (Arvidson & Reynolds, 1995). Mogenson claim sorting capacities of up to lOOtph for their machines (Mogenson, 2000). INCO Mines Technology conducted electronic sorting tests on synthetic McCreedy ore using a combination of optical and conductivity sensing in a Mogenson sorter, which gave results of up to 77% waste rejection at a recovery of 98%) (Schindler, 2001). Modelling and simulation using these results indicated a saving of 10% in mining costs due to an increase in productivity of the mining suite, and a proportionate savings in haul and hoisting costs. It is proposed to use this result as a basis for the inclusion of a combination of optical and conductivity sorting as potential technologies for the pre-concentrator design. Actual metallurgical performance will be determined through testwork. A more difficult application for electronic sorting is the separation of low- and high grade ore from the same ore stream. Testwork reports on copper porphyry ores from a range of mines in the Montana and Upper Michigan area indicate rejection by electronic sorting of up to 50% by mass at recoveries from 85 - 92% (Miller et al, 1978). Again, while not strictly pre-concentration, the Kiruna Mine in Sweden utilizes LIF sorters underground to reject a high-phosphorous iron ore to the low grade stream (Kruuka & Briocher, 2002). Gravity Separation Gravity separation is a low unit operating- and capital cost technology for the separation of mineral and gangue of different densities at a relatively coarse particle size. In the context of underground pre-concentration both cyclones and continuous centrifugal concentrators are considered suitable units for the separation of liberated minerals subsequent to a suitably designed crushing stage. Shaking tables have been shown to give good metallurgical 21 performance on sulphide ores, but are considered inefficient. While no examples of a gravity circuit operating in the underground environment can be found, the technology is generally attractive as it meets the criteria of compactness and robustness in the face of variable feed conditions as identified earlier in the literature review. C O M R O predicted good recoveries on gold ore using a suitably designed classifying cyclone processing the tails from the coarse flotation stage in order to scavenge the partially liberated free gold which would escape flotation in his conceptual underground processing facility (Lloyd, 1978). Dense Media Separation Dense media separation is the principal process technology used in surface pre-concentration plants. It is effective in removing coarse waste from a high grade ore stream at low operating costs and high metallurgical recoveries. While DMS is considered a relatively low capital and operating cost technology, there are a number of challenges to integrating dense media separation into the underground environment. If the complete installation is considered, including feed preparation, the dense media vessel as well as the media circuit, it is not a particularly compact technology. DMS plants are, however, high capacity, and flexible in terms of feed tonnage and conditions, although variations in process efficiency across the vessel can be experienced when the feed tonnage varies. It is believed that these challenges can be overcome, and benefits in terms of process efficiency and improved mineral recovery would be enjoyed when compared to sorting or gravity concentration in the underground environment. A focus of the design effort would be in simplifying the media preparation and recovery circuit, both to reduce the size of the plant, but also to minimize media losses to the underground environment, as this would make DMS cost-uncompetitive when evaluated against other technologies. As previously mentioned, A M T Copper Creek has proposed the installation of DMS underground in order to pre-concentrate a copper porphyry ore. 22 Coarse Particle Flotation Coarse particle flotation must be considered a viable technology for the processing of minerals underground. Bulk sulphide flotation is effective in separating the sulphide minerals from gangue in complex polymetallic ores. In the underground application, bulk sulphide flotation must by needs be preceded by a coarse waste rejection stage where waste dilution and coarse liberated waste can be removed prior to the process in order to reduce the size of plant required. Conventional flotation is typically undertaken at particle sizes between 10 and 200 um in order to optimize recoveries, although conventional flotation has been utilized at particle sizes up to 600pm (Kallioinen & Niiti, 1997). Flotation cells are operated to produce as high a grade of concentrate as possible at the maximum recovery, within the specified residence time. Flotation at these typical particle sizes would require large scale crushing and grinding plant prior to flotation, making this impractical for consideration in the underground scenario. A high energy flotation cell was developed by Lloyd and Hinde during their research into underground processing at COMRO, and was effective in the flotation of gold ore up to a particle size of 3mm when operated in order to maximize recovery at the expense of concentrate grade (Lloyd et al, 1994). Recoveries of up to 98% were predicted for the flotation cell when operated in combination with a scavenging cyclone recovering coarse gold from the tailings stream, although this is considered an optimistic result. New developments in coarse-particle flotation have demonstrated that using hybrid processes and novel vessel designs, good recoveries in the flotation of apatite can be obtained at particle sizes up to 3mm (Hui & Achmed, 1998; Leppinen et al, 2003). The second development of interest for application in underground coarse particle flotation is the Jameson Cell*. The cell comprises a central downcomer delivering conditioned slurry into a conical separation vessel, somewhat similar to a settling cone. Froth is removed via a peripheral launder at the top, and tails are removed from the apex of the cone. Both coarse and fine particle recoveries are claimed to be improved over conventional flotation cells, due to improved particle-reagent contact prior to introduction in the flotation chamber, a greater http://www.rnineraltechnologies .com/jamesoncell.pdf 23 range in the froth-bubble size distribution and a reduction of short circuiting as well as a short path to the froth launder. A combination of technologies is envisaged for the final system. Separation in Froth (SIF) is effective with coarse particles but is not an efficient process; conventional and flash flotation is efficient, but mass pull to concentrates is limited. Due to the requirement for the removal of only 60% of the gangue by mass in the pre-concentration scenario, a substantial flowrate of slurry to concentrates is required. Short-circuiting of fines to the tails stream will be minimized with a mid-level feed arrangement and increased upward slurry velocity towards the froth launder. At a tailsxoncs mass split of 60:40, slurry velocity to underflow will still be greater than that to the froth launder, and short-circuiting will be minimized. In designing flotation cells for the underground application, the liberation characteristics of the ore are critical in developing the machine, thus established mineralogical techniques must be used to determine the maximum particle size distribution at which flotation can occur for the ore (Morizot et al, 1991). A flotation arrangement comprising coarse (-3mm) slurry feed, introduced high in the cell, with high specific energy and air inputs and a high degree of froth removal is envisaged as a starting point. Underground Comminution The principal objective of comminution is to improve the liberation of the valuable mineral. The separability and liberation characteristics of an ore at various stages of comminution can be determined through mineralogical analysis. The objective of process mineralogical evaluation would be to identify the coarsest possible particle size distribution at which between 40-60%) of the gangue material can be rejected without a material decrease in metal , recovery. This is expected to be achievable at fairly coarse particle sizes for most ores based on Lloyd's pre-concentration results for gold at a topsize of 3mm, which typically requires a fine grind (80% - 75um) for full liberation. The optimization of the comminution / liberation characteristics of the ore can be evaluated using conventional techniques (Mclvor & Finch, 24 1991). With regard to the specific comminution technology to be used in the underground application, several technologies are considered likely to be suitable. a) Coarse-particle comminution The sorting and gravity separation technologies mentioned previously will require feed preparation in the form of crushing and screening. While underground crushing of R O M to improve the material handling characteristics of the ore is common, such comminution followed by a mineral beneficiation stage has not been previously undertaken. Crushers common in the underground environment include double-roll crushers in coal and chrome deposits, and jaw crushers in hard rock applications. LNCO have previously investigated the use of a horizontal-chamber 'Eagle' crusher for in the development of a continuous mining system (Dessurault & Scoble, 2002). For this research, due to the constraints of space, only compact or high-intensity comminution operations will be considered. Crushers which comply with the basic design criteria previously established and which will be considered in the selection of equipment for the underground mineral processing plant include: • Jaw crushers (conventional, horizontal) • Interparticle crushers (Barmac) • Double-roll crushers and High pressure grinding rolls (HPGR) • High speed cone crushers (Metso HP series and Osborn Telsmith HS) Interparticle crushers are attractive due to the high-energy input, and low maintenance requirements due to the crushing action. b) Fine-particle comminution Grinding mills have not been used previously for a process purpose in the underground environment. The McArthur River mill is a 300kW S A G mill operating in open circuit with pre-crushing and thickening of the slurry prior to hydraulic hoisting; the mill is not employed to achieve a process objective. However, the McArthur River design is informative in envisaging a suitable grinding circuit for the underground processing application. Grinding plant thus installed must be as simple, compact, and as mechanically robust and reliable as possible. This suggests a number of additional design criteria to be used in the development 25 of a milling circuit, should milling and bulk sulphide flotation be shown to be an attractive process alternative for the underground pre-concentration facility. Conventional milling in the form of SAG or A G milling must be considered due to the relative cost and availability of this technology. SAG or A G mills could be used in open or closed circuit, preferably at low levels of recirculating load in accordance with the requirement for compactness. A S A G mill grinding 1000 tpd of a typical base metal sulphide ore (15kWh BWI) from an f80 of 40mm to an f80 of 3mm would be approximately 3m x 4.5m effective grinding length (EGL), and rated at 1500 kW, probably requiring a total excavation (including ancilliaries) measuring 20 x 20 x 50 m, which is considered in line with present excavation sizes for non-mining infrastructure in deep-level mining (Hartmann, 1992). Other grinding technologies must also be considered. In the development of a flowsheet for the conceptual integrated underground mining and processing system, COMRO developed a 1MW centrifugal mill in conjunction with Lurgi, with a lm diameter by lm E G L (Lloyd et al, 1982). Due to the additional centripetal forces generated with the eccentric motion of the shell, volumetric efficiencies in centrifugal grinding are much improved over conventional concentrically rotating mill shells. In 1979, the South African Chamber of Mines and Lurgi entered into an agreement to produce the mill, leading to the construction of a 1MW centrifugal mill on surface, capable of autogenously grinding quartzite (BWI 19 kWh/t) from 100% -19mm to 100% -3mm and 35% -75um at a throughput of 60 tph. This is equivalent to the typical performance of a 4m x 5m E G L ball mill. The full-scale mill, however, was only tested on surface, and would require further mechanical development for the underground environment. A further technology which must be considered for underground duty for finer comminution is the vertical stirred mill as in operation at Mt Isa and elsewhere as a concentrate regrind or ultrafine grinding mill. Due to increased grinding efficiencies, the dimensions of a 1MW mill can be reduced to approximately l m x l m x 2m in this configuration. While the Vertimill has been designed for the ultrafine grinding from feeds of between 1mm and lOOuni, it is 26 believed that design criteria for a Vertimill accepting feeds of between 10 and 30mm can be developed (Way, 2004). 1.4.2 Backfill Technologies A brief review of backfill technology, the principal types of backfill, the plant required to prepare backfill, and the function, ideal properties and benefits of backfill will be discussed. Finally, the options for the material handling of the products of pre-concentration to surface will be reviewed. As previously mentioned, a principal advantage of underground processing is the direct production of a possible backfill material underground. Some mining methods, such as sub-level caving and vertical crater retreat, do not require or are unable to accommodate backfill, whereas some methods, such as cut-and-fill mining, rely on a supply of backfill for success. The introduction of backfill into room-and-pillar mining methods can significantly reduce the size of pillar required and thus increase the extraction ratio of the mineral resource, as in post-pillar cut and fil l . Coarse particle mineral processing technologies only are being considered preferable for application in the underground pre-concentration facility. Thus it is reasonable to assume that the waste rejected will be competent, and of a coarse nature, between 80mm and lOOum, and would thus be eminently suitable for use as cemented backfill. There are a number of permutations with regard to the layout of the pre-concentration and backfill plants within the overall mining layout. In the case of a centralized pre-concentration facility, it is logical to assume that the underground backfill preparation facility would also be centralized, and ideally situated in close proximity to the pre-concentration plant. Locating the facility at the lower extent of the orebody would maximize the advantages of pre-concentration, and minimize the cost of material handling to the plant; however this will maximize the cost of returning backfill to the stopes. Situating the plant at the upper extent of the orebody diminishes the material handling advantage gained through pre-concentration of the ore by a factor in proportion to the ratio between the vertical extent of the orebody and the depth of the 27 orebody below surface. However, the cost of returning backfill to the mining void is minimized through the possibility of exploiting a gravity delivery system. This trade-off will have to be optimized on a case-by-case basis. Backfill has been proposed as a viable substitute for engineered support in many instances, and compares favourably with engineered support on the basis of cost. Even i f fill is not required for support, the mining void is used to stow excess mining or development waste to minimize the costs involved in transporting this waste to surface. Fi l l methods are often the only safe and reliable method in situations of extreme depth, high vertical stress gradient, or poor ore- and wall rock competence (Dirige, 1999). Cemented backfill has been successfully substituted entirely to replace engineered support in deep-level stopes at a competitive direct cost, and has been shown to substantially reduce rock stresses, plastic rock deformation and rock bursts when utilized at depths below 2000m (Patchett, 1977; Lloyd 1979). Various types of backfill are employed in the industry: • Rock fill • Cemented rock fill • Sandfill • Hydraulic fill • Classified tailings • Pastefill Sources of backfill are numerous. Coarse development and clearly identified mining waste is utilized as uncemented rockfill, but does not impart a significant degree of roof support. Cemented rockfill (CRF) comprises coarse fill which is consolidated by the application of a dilute sand/cement spray for additional compressive strength. Falconbridge's Strathcona Mine in Sudbury employs CRF comprising -5" rockfill combined with 6.5% water and 6.5% Normal Portland Cement (NPC), although they are experimenting with other fill systems in order to further alleviate rockbursts (Swan et al, 1993). Surface mill tailings is also used as a backfill material, and is typically combined with 2- 5% NPC by mass, generally to reduce slumping and improve pumpability rather than for strength requirements; NPC can easily be 28 replaced by fly-ash, granulated blast furnace cement or some other pozzolanic material. This is sometimes referred to as hydraulic backfill, and is delivered to the stope by means of gravity lines, or pumped to the stope, either from a surface facility, or from a backfill plant located underground. M i l l tailings is not an ideal backfill material as the particle size distribution is too narrow, and contains too little fines (<25% < lOum), in addition to a paucity of coarse material, and in order to generate compressive strengths > 1 MPa a substantially higher ratio of cement addition is required. In addition to this, the ratio of water in the fill can be as high as 55%, leading to slumping and drainage problems in the filled stopes (Blight, 1979). These problems have been largely overcome through the development of high-density 'paste' backfill systems, where a superior fill is prepared with a broader particle size distribution and a decreased water content of between 10 and 25% (Brackebusch, 1994). The particle size distribution of mill tailings is modified by classification, and by the introduction of significant amounts of sand to the classified tailings. Moderate quantities of coarser aggregates up to 25mm can also be added to the dense paste without significantly impairing its pumpability (Cooke et al, 1992). Classified mill tailings are dewatered in a conventional thickener in order to preserve the naturally arising fines, and then filtered to approximately 13% moisture. Paste comprising a controlled mixture of filtered tailings, 15% water, 2-4% cement and some additional aggregates is made up in an automated plant on surface and delivered to the stopes by a combination of concrete pumps and gravity pipes of between 100 - 150mm diameter. Delivery of fill over vertical and horizontal distances of up to 2000m is possible. If cement addition is undertaken on surface, pipelines must be cleared after every fill operation. Cement is sometimes added closer to the stope by induction or a via a venturi system. The cement content of sand- and pastefills is typically between 1-5%, and such fills can achieve in situ strengths of between l-3MPa. Due to the high uniformity co-efficient of the particles in the fill, substantially increasing the cement content does not significantly increase the strength of the fill. Furthermore, increasing the cement content to 10% and above is uneconomic in most backfill situations. 29 Target Gold Mine in the Free State of South Africa has been experimenting with an underground concrete batching plant fed with aggregate prepared from mining waste through prior crushing and screening underground (Mining News, 1995). Typical concrete mixtures comprise a mixture of -19mm crushed aggregate, -1mm sand, cement and water in a pumpable mixture of 6:2:1:1 by mass, and can achieve compressive strengths of typically 20 - 40 MPa. Concrete strength varies with aggregate specification and cement ratios in the mix. Backfill is not required to achieve such strengths, and less competent mix ratios and lower cement contents should result in adequate backfill strength. A mix ratio of 12:2:1:1 is suggested as a starting point. There are thus two potential backfill strategies arising from the integration of a pre-concentration facility into the underground mining environment. Coarse wastes generated by the pre-concentration plant can be transported by either L H D or by conveyor to the backfill stopes. Here, the coarse fill would be combined with a cementing fill prepared from surface mill tailings, water and cement. Compressibility of the fill would be high, and compressive strengths low, and thus would only be suitable for non-supportive fill. Additional engineered support would be required in a high rock-stress environment; furthermore this type of fill would not be suitable for situations where plastic rock deformation is expected. The second alternative would be to prepare the waste from the underground pre-concentration plant as a suitable concrete aggregate, and combine with mill tailings, cement and water in an automated underground fill batching facility. Backfill thus prepared would be pumped and/or gravitated to the stopes as required using a system similar to the paste- or sandfill systems currently in use. The advantage of this would be to exploit the high strength characteristics of the coarse pre-conc waste, as well as maximize the degree of underground waste disposal by minimizing voidage by the addition of surface mill tailings. The pre-concentration plant would not produce a fine waste suitable for the preparation of a high-strength mix, and the alternative is to combine pre-conc waste with sand from surface, which does not maximize the waste disposal or cost-saving potential of this concept. As previously mentioned, waste rejection of only 40 - 60% by mass is targeted in the pre-concentration plant design. 30 McCreedy East presently mines between 1500 and 2500 tpd of sand fill from surface for preparation and introduction as fill for the cut-and-fill sequence. The production of a cheap, high strength backfill material directly underground would be of great advantage from both a cost and rock mechanics perspective. 4.3 Material Handling Technologies On the basis of the mining methods under consideration, a typical mining cycle would include mechanized drilling, multi-face blasting, and mechanized load-haul and dump from the face or draw-point to a tipping point. There are many methods of material handling in use from the tip to surface. A typical outbound ore handling system for a deep-level mine would comprise tipping to an orepass, transfer to a lateral bulk transport system such as haul trucks, tipping to the shaft bottom tip and hoisting to surface in lOt or 20t skips. Both ore and waste are handled in this manner. A n alternative system would be to transport the ore laterally from the L H D tip via belt conveyor and transfer to a central decline conveyor (in the case of a shallower scenario), or to a shaft-bottom tip. It is, therefore, a given that outbound ore transport to the pre-concentration plant would thus be via L H D to a batch haul-truck system or to a continuous system such as a system of section conveyors. It is most likely that the unit operations comprising the pre-concentration facility (scalping, crushing, screening, sorting, gravity concentration etc.) would be distributed at points throughout the entire underground facility, and comprise a combination of distributed (in section) and centralized processes prior to transporting the concentrated ore to surface. The use of a combination of belt conveyors, chutes, bins and feeders within the battery limits of the underground pre-concentration plant is assumed. Wherever possible, mining excavations and raise bores would be utilized for the construction of chutes, bins and tanks in order to minimize the cost of construction. Once ore has been transferred from the discontinuous mining process to the continuous pre-concentration process it would be preferable to utilize a continuous material handling system such as 31 pumping or belt conveyors from that point in the system onwards in order to minimize the substantial provision for surge that is required when converting from batch to continuous processes and vice versa. In the underground scenario, the minimization of vertical height in the plant layout is considered essential in order to minimize cost. The review of material handling technologies pertinent to the consideration of underground pre-concentration can be split into three areas of applicability. Firstly the material handling requirements for ore from the working face to the pre-concentration facility must be reviewed. Secondly the nature and transport of the waste products from the pre-concentration stages back to the mining void must be considered. Alternative Hoisting technologies The implementation of pre-concentration underground presents unique and novel opportunities for the material handling of the pre-concentrated ore to surface. The nominal particle size of the ore would be significantly reduced subsequent to the pre-concentrator, and the mineral or sulphide content will be higher, and thus the Bond- and abrasive work indices will be significantly lower. This suggests that the products of an underground pre-concentration plant would be amenable to alternative hoisting methods such as vertical conveying and hydraulic (pumped) ore hoisting. It has been suggested previously that to convert from a continuous processing system to a batch handling system post the pre-concentrator is inadvisable, however haul trucks and conventional hoisting may still be considered for this task in some instances, especially if there is already an investment made in such equipment. There would be substantial capital and operating cost savings in these cases due to the significantly reduced mass and particle size after pre-concentration. Operating and capital cost savings would be directly proportional to the quantity of waste rejected. In the case of greenfields operations considering pre-concentration, these savings would be maximized. Due to the existence of an existing hoist and shaft at Coleman, the effect of these potential savings is significantly diluted. 32 Continuous material handling systems after the pre-concentration plant would be preferable, and has been shown to offer lower operating costs than comparable batch material handling systems. Underground conveying, in conjunction with crushing has been utilized extensively in both shallow and deep underground mines in order to optimize the material handling characteristics of the ore. Underground conveying, both horizontal and inclined is common. Integrated systems comprising load-haul-dump operations tipping to section conveyors which deliver ore to a central decline conveying system is a standard in shallow room-and-pillar coal, chrome and platinum mines, and such concepts have been used in deep mines as well (Paul, 1990). Inclined conveying on flat or troughed belts at tonnages in excess of 1000 tph is practical up to angles of 20°. With the use of textured or cleated belts this angle can be increased to 35°. Tube conveyors, chain conveyors and bucket elevators are an alternative for steeper angles, but are impractical and involve high initial capital and ongoing maintenance costs. Flexowell® belting was originally designed to provide a belt-based material handling alternative in confined layouts, allowing the conveying of approximately 700 tph at angles up to 75° on suitably designed belts. Metso Minerals have recently acquired the rights to the Flexowell® system, and offer the Pocketlift 'S ' vertical conveying arrangement which has been used in limestone tunneling applications as well as hardrock quarries and mines for lifts between 70 and 500m and tonnages between 250 and 1200 tph. Both manufacturers and engineer's estimates indicate capital savings of 36% and power savings of 12% when compared to an inclined conveying solution in the same application (Paelke, 1997; Yester, 1997). Such developments make a Flexowell-type vertical conveying installation a viable technical alternative for the hoisting of hard rock ore in deep mines and must be considered in the material handling design. 33 A second alternative to be considered is hydraulic hoisting. In hydraulic hoisting ore is diluted into a coarse slurry or high-density paste and pumped to surface using either high-pressure centrifugal or positive displacement pumps. Much research and development has been undertaken by CSIRO in Australia into hydraulic conveying, investigating in particular optimum slurry rheology, and the design of suitable pipelines and the selection of pumps*. With regard to suitable high-pressure pumps, two manufacturers produce possible units. Warman currently produce a high-pressure all-metal slurry pump capable of pumping 1000 m 3/h slurry at 10 bar; Gardner-Denver supply a multi-cylinder high pressure positive displacement mud-pump capable of displacing 300m3/h of slurry at 30 bar. Pipelines fabricated from new engineering materials such as UHDPE and HDPE-lined steel are commonly used in overland tailings applications over 5000 - 10,000m horizontal delivery and 16 bar pressure. Anglo American Corporation had previously installed a hydraulic ore hoisting system at their Vaal Reefs North mine. The system was installed using Mars-type positive displacement pumps delivering 136 m 3/h of slurry at an SG of 1.28, at 7.85 MPa delivery pressure (Schuttler, 1977). The hoist operates in four stages over a vertical height of 2191m, and delivers a throughput of 983 dry tons per day. The system was installed to supplement the capacity of the existing 2240kW 4000 tpd rock hoist. There were further benefits in terms of improved loading and material handling efficiency due to the reduction in fines in the conventional skips. The CSIR in South Africa have also developed a system utilizing dual-medium pumping for the hydraulic transport of broken ore and waste up to a topsize of 12mm and between 45 and 67% solids by mass, using a positive displacement TORE® pump installed integrally within a feed tank (Ilgner & Kramers, 1997). Cameco's McArthur River uranium operation also employs positive displacement hydraulic hoisting subsequent to the underground crushing and milling of their uranium ore (Edwards, 2004). http://www. 34 A scoping study into hydraulic hoisting was conducted by INCO Ltd at the Copper Cliff Mine in 1990§. Both centrifugal pumping and positive displacement pumping using Siemag® pumps were considered in the design. INCO concluded that hydraulic hoisting was more attractive than hoisting or truck haulage on the basis of operating cost/t, but compares unfavourably to conventional hoisting from the point of view of capital cost. However, the poor capital cost comparison of hydraulic hoisting may be further improved when considered in conjunction with underground pre-concentration. There are several alternative pumping arrangements for hydraulic hoisting, with varying impacts on the performance of the concept with respect to pumping efficiency and wear. These alternatives include: • solids and transport medium fed directly through the pump • pumped transport medium and solids introduced into the pipestream • air-assisted lifting of solids in the transport medium • jet pumps Hydraulic hoisting systems reviewed in this section fall into one or a combination of these categories. Other hydraulic hoisting concepts which are not considered for this thesis, but must be mentioned for completeness include varieties of an alternating lock-hopper systems proposed by the U S B M (Dierks, 1964), the U K Ministry of Fuel and Power (Spencer, 1952), Falconbridge Mines (Pasieka et al, 1978), the Hitachi hydrohoist system (Singhal, 1970) comprising alternating stages of low-pressure slurry feed and high pressure medium pumping for transport, as well as a continuous capsule lifting system for deep-sea mining (Donkers, 1980). Challenges to be overcome in establishing hydraulic transport as a viable technology include overcoming wear problems in the transport of coarse slurries at high pressure, pressure limitations in conventional centrifugal pumps, wear and particle size limitations of high-pressure PD pumps, the batch nature of the dual-medium systems, and the inherent INCO General Engineering Department Estimate No. 2409-OA, 1991 35 inefficiency of pumping the transport medium as well as the ore over large vertical distances. The concept of energy recovery from the downcoming transport medium has not been covered in the literature. Energy could be recovered from the increased quantity of process water required for underground pre-concentration by means of pelton wheels or reverse-running clear water pumps. Energy recovery from all downcoming water to the underground mine is also a concept which should be explored. Presently, downcoming water headers require complex and expensive pressure reduction systems in order to limit the pressure of water delivered to each level. Unreduced, the static pressure in a water column for a shaft 1500m deep would be of the order of 15 MPa - this potential could be utilized to provide energy for hydraulic hoisting. Again, the concept of underground pre-concentration addresses a number of other issues in providing a smaller quantity of processed ore suitable for transport as a fine slurry, which would also reduce the cost of hoisting. 36 1.5 Conclusion The individual technologies for an integrated underground mining and processing system already exist at some level of implementation in the industry, and particular precedents for the practical application of these technologies also exist. It remains for these technologies to be coherently designed into an integrated mining and processing system comprising mechanized mining, underground pre-concentration, automated preparation and placement of the waste as backfill, and the hoisting of the pre-concentrated ore using unconventional methods. A case study based on the research presented in this paper has been undertaken at INCO's McCreedy East operation. Suitable process, mining, and engineering design methods will be applied to the development of the system and a preliminary economic analysis for the concept as applied will be developed. 37 Chapter 2 - Case Study for INCO's McCreedy East Mine 2.1 Introduction The process design, layout and costing for an underground pre-concentration facility at McCreedy East Mine in Sudbury, including provisions for returning the waste thus generated as backfill for the mining operation, as well as the removal of products and process water effluent from underground has been undertaken for INCO Mines Technology. The objective of this underground facility would be primarily to reject waste dilution underground and consequently improve the grade of the R O M (Run-of-Mine) ore delivered to surface. It is anticipated that savings will be made in mining, hoisting and backfill costs as well as the cost of surface transport to Clarabelle M i l l . The possibility of producing a directly smeltable concentrate from the McCreedy 153 ore has also been identified as an opportunity. A number of previous studies into sorting and the pre-concentration of ore have been undertaken by INCO Mines Technology (Buksa & Paventi, 2002; Exportech, 2002). The focus of this previous work was to identify viable process technologies and test the technologies on INCO ores. Electronic sorting methods, and in particular optical, conductivity and high-intensity magnetic separation, were identified early on as an attractive technology, and a number of tests were undertaken on ores from the Sudbury region. The results of this testwork were mixed: a number of ores showed good amenability to sorting. Results for low grade or highly disseminated ores, however, were less encouraging. The results of this previous testwork are shown in Table 1. At this stage the potential for sorting of the Sudbury ores has been clearly established; however there has been no specific study by INCO to date of the benefits to the entire mining and processing system of processing ore underground and the engineering aspects of constructing a mineral process plant in the underground environment. 38 Table 4 - INCO Ore Sorting Testwork Results Test Ore Method Feed Grade Reject Concentrate C u % N i % W t % C u % R e c o v e r y N i % R e c o v e r y 1530pt 153 Opt ica l 5.67 0.4 78.7 24.76 92.9 0.47 75.4 153 Opt ica l / Optcon 153 Conduc t i v i t y Opt i ca l / 5.67 0.4 54.6 11.83 94.7 0.73 81.2 V.384 M O B M a g n e t i c Opt ica l / 4.33 2.4 12.3 5.93 99.8 3.5 99.6 V.444 M O B M a g n e t i c Conduc t i v i t y 1.81 2.4 41.6 4.27 99.3 2.78 99 INCO S u d b u r y 90% Conduc t i v i t y 1.09 1.06 38.5 1.55 87 1.66 96.1 INCO S u d b u r y 100% 1.19 1.21 23.5 1.5 96.7 1.56 99.1 INCO B i r c h t r e e n/a n/a 0.96 80.5 n/a n/a 4.79 97.5 Exportech S t o b i e 6283 H G M a g n e t i c n/a n/a 23.8 2.74 94.87 1.874 98.9 Exportech S t o b i e 6287 L G M a g n e t i c n/a n/a 13.03 0.365 92.69 0.873 97.5 Exportech S t o b i e 6276 L G M a g n e t i c H e a v y n/a n/a 12.42 0.604 89.06 0.376 98.95 U B C 1 153 Liquid H e a v y 13.26 0.39 55 29 98 0.8 91 U B C 2 M O B Liquid 1.11 2.54 22 1.32 3.02 3.02 97.3 U B C has been involved in research into Mine-Mill Integration and underground mineral processing since 2000. Early research focused on identifying and quantifying the benefits of underground pre-concentration, as well as researching mineral processing technologies which would be suitable for underground pre-concentration. It was realised very early in the research that the particular focus should be on coarse-particle processing technologies. These technologies include: sorting (optical, conductivity, magnetic, fluorescent) gravity concentration and dense media separation coarse-particle flotation In 2001, a simulation study was undertaken for INCO by U B C to evaluate the benefits of underground pre-concentration compared to resue and drift-and-fill mining methods at McCreedy East Mine. The study indicated that underground pre-concentration would result in a 10% saving in mining costs when compared to the above mining methods (Schindler, 39 2001). Previous optical sorting testwork by INCO was used as a basis for the process design for the simulated pre-concentrator. This previous work was used in developing a proposal for a scoping study into underground pre-concentration at McCreedy East, which was presented to INCO in November 2003. It was agreed to proceed with the study, and fieldwork commenced in January 2004. At the start of the fieldwork, four basic scenarios were identified during the course of the fieldwork, and evaluated using preliminary estimates of metallurgical performance, as well as estimates of capital and operating costs in order to identify the preferred option. With reference to the overall layout of McCreedy presented in Figure 4, the four underground pre-concentration alternatives were: O p t i o n 1 - pre-concentrate the combined 153/170 and M O B ore production of McCreedy East at the present ore handling facility on 3240 level at Coleman shaft. Waste would be returned to the 3770 rockpass at Location 9 by Kiruna truck on back-haul. O p t i o n 2 - pre-concentrate only the production from the 153/170 orebodies at 3240 level with the objective of producing a smeltable concentrate. The high-grade product would be hoisted separately to surface. O p t i o n 3 - pre-concentrate 153/170 ores only at Location 8 on the Kiruna ramp, again with an alternate hoisting arrangement for the high-grade ore. Waste would be conveyed directly to the rockpass at Location 9. O p t i o n 4 - pre-concentrate both the 153/170 and M O B / L M O B ores in separate pre-concentration facilities and transfer the pre-concentration products separately to surface. 40 Figure 4 - Simplified McCreedy East Mine Section A preliminary plant design was established, and a process model for throughput, recovery, mass-pull to concentrate as well as operating cost was used to estimate cost, mass flow and grade impacts in the mining system. A preliminary cost and revenue impact study was undertaken on the above options, and on the basis of maximum waste rejection, maximum cost savings, and possible positive revenue impacts in making a smeltable concentrate, the fourth option was.deemed by INCO to be most attractive. A schematic of the preferred pre-concentration option is shown in Figure 5. The basis of evaluation as agreed at the beginning of the project was to assess the pre-concentration potential of the orebodies to be mined as per the present mining plan at McCreedy East, and to perform a Net Present Value analysis on the impact of pre-concentration of these ores against the tonnages, grades, metal values and financial parameters in the present life-of-mine plan. 2004 working cost figures from McCreedy have 41 been used, broken down by functional activity. Costs and revenues shown in the analysis are shown by difference, i.e. costs were shown as cost savings (or increases) and revenue impacts are over and above the present revenue forecast for the mine. Revenue estimates were based on INCO's present long-term metal price forecasts used in the existing business plan. The maximum savings, as well as the maximum revenue potential for the underground pre-concentration concept at McCreedy was shown for Option 4 comprising separate pre-concentrators for the two different types of ore at McCreedy. In the preliminary analysis, a budget of $15m was allowed for the design and construction of this option. Based on the rejection of an average of 30% low grade waste to backfill directly underground at a metallurgical recovery of 95%, additional pre-concentration working costs of $2/ROM ton, and a projected increase in revenue through an increase in the overall P G M recovery, an NPV of $118m and an IRR of 113% at a discount rate of 11% was indicated (Bamber, 2004). 42 McCreedy East Underground Pre-concentration Study - Option 4 Concentrate Tankeq to Smelter Colem an Hoist Ore truck to rail load out Figure 5 - Combined underground pre-concentration facility for McCreedy East Mine 43 Preliminary fieldwork included familiarization with the setup and layout at McCreedy East in January/February 2004, as well as the procurement of stope samples from the M O B and 153 for mineralogical and metallurgical testwork at U B C during the second quarter of 2004. Further sorting testwork at Mogenson in Aachen, Germany was planned, but has been postponed due to budget constraints. This report summarizes testwork, process design and preliminary cost engineering, and conclusions to date as well as recommendations for further research and testing. Sections covered include: • Background on the mining and smelting operation at McCreedy East • Mineralogical Evaluation and Process Development • Integrated Mining and Processing System Design and Costing • System Impact Evaluation Metallurgical test reports can be found in Appendices 1 -3. Preliminary isometric layouts of the plant have been developed to assist in visualizing the underground facility as well as to provide a basis for excavation sizing and costs, and are presented in Appendix 4. Detailed mechanical equipment lists have been developed from the process design, and are presented in Appendix 5. Mass balances developed from the testwork results and process design criteria are contained in Appendix 6. Capital and operating costs have been estimated for the design and construction of the facility described in this report. Capital costs are calculated using factors for civil, structural, electrical, piping, and support infrastructure based on the estimated cost of mechanical equipment for the plant. Additional indirect costs such as design and construction costs, startup and maintenance costs have also been calculated using factors based on the cost of mechanical equipment. Capital cost estimates are presented in Appendix 7. A preliminary assessment of possible system impacts, including mineral reserve impacts, cost, revenues, labour impacts and impacts to the underground and surface environment is included in Chapter 3. The details of the financial evaluation are presented in Appendix 8. Supplementary technical information can be found in Appendix 9. 44 2 Background 2.2.1 Geology and Mining The case study is for an integrated underground mining, mineral processing and waste disposal facility for INCO's McCreedy East mine in Sudbury, Ontario. It is a typical Sudbury Igneous Complex deposit. The geological setting, mining and ore handling methodology and pertinent aspects of the present milling and smelting system are reviewed. Sudbury ores fall into three principal categories: contact ores, footwall ores and offset-dyke deposits (Peredery & Morrison, 1984). Contact ores are hosted in the ultramafic zone at the contact between the igneous complex and the transitional Sudbury breccias, and are typically rich in N i , but poor in other metals. Contact ores are also associated with the occurrence of footwall ores comprising narrow-vein chalcopyrite stringers situated in the Sudbury breccias down-dip of the contact. It has been suggested that the origin of these secondary orebodies is related to copper remobilisation after the intrusive event. Footwall stringers are rich in Cu, and PGM's , but poor in N i . Other economic mineralization in the complex occurs in the form of narrow, steeply dipping dykes, offset perpendicular from the main intrusion. These offset dykes are host to Cu, N i , A u and PGMs (Coats & Snajidr,1989). McCreedy East is a medium-depth base metal mine located on the north western arm of the Sudbury Igneous complex, some 30km east of the towns of Onaping and Levack in Greater Sudbury. McCreedy East mines both contact- and footwall type deposits. Principal orebody access is via the 1000m deep Coleman shaft. Access to the orebodies is from the shaft bottom, via a 4km haul ramp inclined at a maximum of 12°, supplemented by a small-vehicle service ramp. A generalised section of the mine, showing mineral reserve estimates can be seen in Figure 6. 45 FECUNIS No.1 McCREEDY EAST F A SHAFT McCREEDY EAST R.A. SHAFT MAIN OREBODY TONS % CU % N I TPM Po/Ni 19,827.059 0.99 2 00 0.023 16.8 EAST OREBODY TONS % C U % N I TPM Po/Ni ORE 3.573.482 0.47 1.48 0.006 17.7 WEST OREBODY TONS ORE 1,977,000 COLEMAN No.1 SHAFT 153 OREBODY TONS % CU % Nl TPM Po/Ni ORE 5,338,247 10 70 1.18 0.343 1.5 7386 ZONE TONS % CU % N I TPM PtfNt ORE 179,294 8.22 0.35 0 253 2 5 TONS % C U % N I TPM Po/Ni ORE 234,963 3.63 164 0.095 2 5 COLEMAN TONS %Cu %Ni TPM PtrfNi 228 OREBODY 5,363 1 02 1 53 0.018 19.5 MAIN V\EST 367,359 0.54 1 62 0.005 2 2 8 MAIN EAST 450,010 0.72 1 56 0.008 2 2 4 HIGH COPPER 0 0 0 0 0 EAST LENS 108,514 0. 70 1.67 0.009 26 9 TOTAL 931,246 0 69 1.50 0007 23.1 22a.CREBOOY j J l k MAIN WEST . 1 Kir HIGH COPPER A MIDDLE ZONE LOWER COLEMAN TONS %Cu %Ni TPM Po/Ni MIDDLE ZONE ORE 245,079 0 48 1 51 0 008 16.1 LOWER ZONE ORE 1,244,627 0.92 1.69 0 017 16.8 TOTAL 1,489,706 0.85 1 66 0 016 18.4 COLEMAN / McCREEDY EAST MINE Figure 6 - Generalised McCreedy East Longitudinal Section (from McCreedy East 1999 5 Year Plan) 46 Future production at McCreedy comprises ore mined from four principal orebodies, the Main orebody (MOB), the Lower Main orebody (LMOB), the 153 orebody and a future orebody, the 170. The Main and Lower Main orebodies are geologically similar; it is also believed that the 170 is a down-dip extension of the 153, offset by a regional fault, and mineralogically similar to the 153. Production from the 170 orebody is scheduled to commence in 2006. For the purposes of the plant design, a representative mean tonnage and grade has been selected from the MRI (Mineral Resource Inventory) for each orebody (Table 5). Relevant locations mentioned in the text are illustrated in Figure 7. Table 5 - Mining Rates, costs and values (from McCreedy East 2004 Mine Plan) M O B L M O B 153 170 Total / average Mining rate tpd 1800 1500 1100 600 5000 Cu % 0.90 0.92 11.42 7.01 3.03 N i % 1.85 1.94 1.22 0.92 1.7 T P M g/t 0.64 0.96 10.56 14.4 2.88 Mining cash cost $/t 45 54 84 124 88 Value Equivalent Nickel Units/t 29.9 29.9 81 81 50 The Main and Lower Main contact deposits are thick and shallow dipping, with a high Ni:Cu ratio and containing complex pendlandite/pyrrhotite with chalcopyrite occurring as massive sulphides. Small amounts of chalcopyrite occur as disseminated sulphides in the host rock of the M O B / L M O B . A n exception has been observed in small areas of the WOB (West Orebody), an offset of the L M O B , where complex disseminated sulphides and P G M mineralization sometimes occur in the contact as well. As the principal sulphides in these ores are pentlandite and pyrrhotite, these orebodies are the principal contributors to nickel production at McCreedy. The pyrrhotite/nickel ratio is of great importance for the mill and the smelter, and is maintained below 15:1 for the combined ore production from McCreedy. The mining method is by overhand post-pillar cut-and fill, and estimates of dilution in each panel vary between 20 and 80% for the M O B and L M O B (Figure 8). 47 RW HR/AV m KI4 u / T 0 / t r / s k / a v COLEMAN - McCREEDY EAST MUCKING CIRCUIT 12332 CUFT 667T ROCK 806T ORE 10-DUMPS -BN N I / C U / T O / T R / A V / B L IIO-DUMPS -SUOE N I / C U / T O / T R / A V 10-SKHST -ROPES N I / C U / T O / T R / S K / A V 10-SKHST -SKP5 10-SKHST -SW3 N l / C U / T O / T R / S K / A V N I / C U / T O / T R / S K / A V 2 X 1400 TON ORE BINS 2.6 HRS • 540 TPH EACH SURFACE TRUCKING RFOtJIRFMFNTS ACTUAL PERFORMANCE AT 2 2 0 0 T / 0 WITH 40 TON TRUCKS = 55 TRIPS (2 TRUCKS O 20 MIN/TRIP = 9.2 HRS) PROJECTED PERFORMANCE AT 10,000 T /O WITH 40 TON TRUCKS = 250 TRIPS (5 TRUCKS 0 20 MIN/TRIP = 16.7 HRS) X L n o o o o II /r. TRUCKING REQUIREMENTS FOR COI FMAN ACT1IAI PERFORMANCE AT 2200 T / D THROUGHPUT WITH 30 TON TRUCKS = 74 TRIPS WITH 8 YD SCOOPS = 136 TRIPS (OR A BLEND OF BOTH) 36-X48" JAW CRUSHFR PERFORMANCE SPECIFICATIONS 4" JAW SETTING - 450TPH 6" JAW SETTING - 635TPH 8" JAW SETTING - 825TPH 10" JAW SETTING - 1012TPH ACTUAL PFRFORMANCF AT 2 2 0 0 T / D THROUGHPUT 5.5" JAW SETTING O 50OTPH RUNNING TIME 4 . 5 - 5 HRS PROJECTED PFRFORMANCF AT 10.000 T / D THROUGHPUT 5.5" JAW SETTING 8 5 5 0 TPH RUNNING TIME 1B.2-19HRS 10-TRUCKS-KIRUNA919 N L / C U / T O / T R / A V / K M / T R A E A T / H R START NAME CODE LBS NI NI LBS CU CU CYCLES CY DISTANCE KM TRIPS TR LOADS LD BUCKETS BK SKIPS SK COUNTS CO LOCATION FROM LF LOCATION TO LT TONS ROCK TR TONS ORE TO HOUR METER HR KWHOURS KH OPERATING HOURS OH NO OPERATOR NO NTRS^ IT P D "/CU/TO/TK/AV/CO OPERATING 0OWNT1ME 00 PLANNED MAINTENANCE PM BREAKDOWN MAINTENANCE BM BIN LEVELS BL FILL TONS FT 19750 CUFT 1068T ROCK 1291T ORE 10-UGCHUT-BIN3370-3 N I / C U / T O / T R / A V / B L 10-gGCHUT-CH3370-3 , NI/CU/TO/TR/AV/CO 10-CR5HNG-RCKBREAKER NI/CU/TO/TR/AV/HR 10-CRSHNC-CfiZ71Y3370 NI/CU/TO/TR/AV 1000T 1.8HRS«540TPH 10-CRSHNG-BIN3470 NI/CU/TO/TR/AV/BL 10-CR5HNG-CH3470 NI/CU/TO/TR/AV/CO , « O 3 « o t p h NI/CU/TO/TR/AV/HR 10-CRSHNG-CR3470 N I /CU / T O / T R / A V / C O / H R 900T 1.7HRS«540TPH fS/,C0/td'7rR'7A7/BL 7AV/CO 60"X16' 200-1000TPH 10-UGCNVY-FD3D90-1 NI/CU/TO/TR/AV/HR 8080' TO OR PASS N I / C U / T O / T R / A V / C O 4810-4945L CUFT" T R O C K T O R E FUTURE MUCK CIRCUIT INFRASTRUCTURE REVISION DATE REVISION DATE INCO LIMITED COLEMAN McCREEDY EAST PROCESS FLOW PRINT SUBJECT MUCKING CIRCUIT MINES WpcnfpiaUpsau Figure 7 - McCreedy Mucking Circuit Flow Diagram (s:\mtsdata\productionplan\1999 5 Year Plan\ mucking circuit.dwg) 48 Figure 8 - Main Orebody Panel Grade estimation is done on a daily basis by beat geologists in order to control the mining grade and dilution. Ore is blasted, and typically mucked to a remuck pile in the section, then mucked from the remuck to an orepass leading to a Kiruna chute located on the main Kiruna haul ramp. The MOB and Lower MOB are serviced by chutes at Locations 13 -17, approximately 1.9km from Coleman shaft. As the mining of the LMOB progresses deeper, the ramp will be developed further and an additional Kiruna chute will be established beyond Location 17. Ore is hauled to Locations 1 or 2 on 3220 level near Coleman Shaft, above the hoe-ram and crushing facility on 3370 level. The 153 is located to the East of the MOB, and the 170 extends from the lower portion of the 153 back in a westerly direction below the horizon of the LMOB. The 153 and 170 orebodies are footwall deposits, occurring down dip from the MOB. They occur as narrow-vein massive sulphides, occurring in multiple veins and stringers, varying in width from 0 - 6m, and dipping variably between 20- 60°. The veins are typically massive chalcopyrite, grading at 30% copper, with secondary veins of pentlandite, millerite and occasionally bornite. 153 and 170 ores are high in Cu and PGM's, however, Ni content is low. The veins are mined to a present cutoff width of 100mm. The overall horizontal extent of the orebody varies typically between 5 - 50m, and the average dilution in 153 ROM ore can be as high as 70%. Furthermore, the ore is extremely soft and friable; waste is substantially more competent than the sulphides, thus ore typically 49 fragments to a finer particle size distribution than the waste during the blasting of a round. This characteristic can also be observed in the MOB and LMOB. Ore is mined in multiple headings which extend laterally from a central drift developed off the main footwall drift on each level. Panel dimensions vary from a minimum of 8ft x 9ft to relatively open stopes in the thicker veins (Figure 9). Remnant pillars are often left in the open stopes for additional support. The mining equipment can be jacklegs, single or double-boom jumbos, LHD's vary from 1.5 to 6 cu.yd capacity, and either manual or automated roofbolting is used depending on the width of the seam, and thus productivity varies widely from heading to heading. Ore is mucked from the headings to a re-muck area on the central drift, and then haul to an orepass on the footwall drift. The orepass may be a transfer pass, or may lead directly to a Kiruna chute. 153 ore is hauled approximately 3.9 km from locations 50 - 69 on the Kiruna ramp, located between 4360 and 4945 levels. A future Kiruna chute is planned forthe 5130 level. Access to the 170 will be by an extension of the 153 Kiruna ramp, haul distances will be consequently even greater. Figure 9 - Typical 153 Narrow-Vein Heading Excess rock from the 153 orebody is hauled to Location 9 to be made available as rock for the MOB, from where it is mucked by scoop tram. When Location 9 is unavailable, rock is hauled to Location 1 for hoisting. Waste rock from the MOB and LMOB is mucked to rockpiles within the orebody or to Location 9, to be redistributed as fill prior to 50 final sandfilling. There is presently a net shortage of fill, and the shortfall is made up through the mining of sand on surface at a rate of 1500 - 2000 tpd and a cost of $4/t. Sandfill is batched in a surface plant and distributed to the working faces by a network of 4, 5 and 8 inch pipelines. A representation of the overall underground mining and material handling process flow is shown in Figure 9. Note Location numbers and haul distances increasing with depth. Direct mining costs and costs to surface in terms of $CDN / ROM ton for the various orebodies as per the 2004 Mine Plan are shown in Table 6 (Thomas, 2004): Table 6 - McCreedy East Mine Cash Costs to Surface, CDN$, all ores Orebody Mining Backfill/ Sandfill Mucking Haul Crush/ Convey Hoist Total M O B 17.08 4.38 4.27 3.3 1.37 1.08 31.48 L M O B 26.44 4.38 3.87 4.1 1.37 1.08 41.24 153 52.9 9.29 5.97 10 1.37 1.08 80.61 170 81.89 9.29 5.97 12 1.37 1.08 111.60 2.2.2 Milling and Smelting Milling of the ore takes place at the centralized Clarabelle milling complex near Copper Cliff, South of Sudbury. Ore is transported by truck from the Coleman Shaft to the rail loadout in Levack at the western boundary of the mine property. From there, ore is shipped 60km by rail to the ore receiving facility at Clarabelle Mil l , which utilises grinding, selective flotation and magnetic separation for recovery of the base metal sulphides and rejection of pyrrhotite. The Clarabelle Mill is designed for a maximum of 10.6 Mtpa, and processes all ore for the Ontario Division. The mill utilizes two milling circuits, a conventional 10,000 tpd crushing, rod-mill, ball-mill circuit and a 20,000 tpd SAG milling circuit. A low-grade base metal sulphide concentrate is produced using flotation, and sent to the smelter at a minimum grade of 20% Cu + Ni, and 5:1 Cu:Ni ratio (see INCO Ontario Operations Flowsheet 2002 in Appendix 4). Milling costs are reported variously at between $7/t and $9/t. Tailings costs are included in this figure, and are estimated at $2/t (Kerr, 2004). Concentrate from the Clarabelle Mil l and other toll 51 concentrates, together with future production from Voisey's Bay is smelted at Copper Cliff smelter, some 4km from the mill. Concentrate is flash-smelted to produce a matte, and Fe is then oxidised to Fe2C>3 and removed in the matte converters. A small amount of metal (<5% ) is lost to slag in the smelting process. Flash smelting is costed at $10 per ton smelted. Costs were not available for other downstream activities. For the purposes of financial modelling and evaluation, the smelting toll has been assumed to be 20% of the contained metal value in the feed concentrate. 52 2.3 Sampling and Mineralogical Testwork Observations made during the site visit as well as the results of previous testwork and ore characterization by INCO had indicated that the ore as observed in the MOB, LMOB and 153 stopes at McCreedy showed good potential for pre-concentration due to a distinct difference between the particle size distribution of the ore and waste, as well as significant disparities in sulphide content, colour, conductivity, magnetic susceptibility and density. This has been borne out by previous sorting testwork conducted by INCO on other Sudbury massive-sulphide ores (Buksa & Paventi, 2002). Testwork on McCreedy ore was carried out to confirm the mineralogy, grade, size characteristics, and grade distribution by size. This could then be used in the generation of a mass and grade balance for the development of a process flowsheet as well as to size and cost the equipment for the pre-concentration plant required for each type of ore. The sampling and testwork programme was undertaken between February and June 2004, and comprised the following objectives: 1. Characterise a typical ROM stope sample from the 153 and MOB according to particle size and grade distribution. Samples for analysis were taken from the 153, MOB and the surface stockpile. 2. Mineralogical characterization of the ore 3. Liberation and separability testwork on 153 and MOB ore. 4. Bench scale process testwork on sorting, Dense Media Separation and bulk sulphide flotation. 5. Confirmatory combined optical and conductivity sorting testwork on the AE 80 sorter at Mogenson in Germany. Within the constraints of the time and budget allowed, only mineralogical characterization and liberation and separability testwork have been completed to date. Bench-scale flotation and sorting testwork has not been completed. Additional research is recommended in order to explore these aspects of the work further. Representative stope samples were taken from McCreedy and analysed using process mineralogical techniques. The size distribution, grade distribution, liberation and separability of each sample was assessed using standard procedures (Finch & Gomez, 1989). Sink/float tests at a separation density of 3 t/m3 were performed on each size fraction, assays were taken, and the resulting size/grade distribution developed. The grade 53 and degree of liberation of the sulphides in each size fraction was noted. Polished sections of the ore- and waste fractions were also taken for mineralogical analysis. Summary results from the mineralogical analysis of the MOB and 153 ores are presented below. Detailed results can be found in the mineralogical reports in Appendix 1 & 2. These results have been used to develop basic process design criteria for the respective pre-concentration plants. Additional process design criteria presented here have been drawn from literature and from experience in designing process facilities on similar base metal ores. 2.3.1 153 Ore Evaluation A 500kg stope sample of 153 ore was taken by LHD from the 4550 level, X/C 1W on 5. February 2004. Ore was split into 4 sub-samples and subjected to a screen analysis, then sink/float testwork was undertaken on each size fraction, and the respective sink and float fractions sampled and assayed for Cu, Ni and PGM's. The size / grade distribution curve for the ore was reconstructed from the results. Micrographs of representative ore and waste fractions were also taken using an optical microscope. The fragmentation of the ore as blasted in the stope appeared to be relatively fine (Figure 10). A clear distinction between the sulphide and gangue size distribution was noted. Miners characteristically call a high-grade round 'sugar' due to the extreme fineness of the sulphides. 54 Figure 10 - 153-4550 muckpile The overall particle size distribution of the 153 ore was considered fine for a stope sample. The sulphides in the ore were clearly distinguishable from the waste in each size fraction due to discrepancies in colour, lustre, and density. Some middlings particles as shown in the upper left corner of Figure 11 were observed. The waste also has a sharp and angular appearance when compared to the more rounded appearance of the sulphides. Liberation as observed visually in each size fraction was good, improving towards the finer size fractions (Figure 12). Figure 11 - 153-4550 +53mm ore & waste 55 Figure 12 - 153-4550 +6mm fraction Liberation and grade both improved towards the finer size fractions. Hand sorting was used to separate sulphides from the ore in the +53, +75 and +125mm fractions, and heavy liquid separation was used to separate the sulphides from the waste in the finer fractions. Sulphides can be clearly distinguished visually from the waste after separation (Figure 13). Sulphides are extremely weak and friable, large sulphide particles break up in the hand. A significantly lower Work Index of the ore compared to the waste was observed in assay sample preparation: the sinks fraction milled up to 4 times quicker in preparation than waste fractions of the same size distribution. 56 Figure 13- 153 4550 +9mm sinks and floats The size distribution of the 153 stope sample is shown in Figure 14. The topsize was observed as 250mm, with dso=12mm, and 27% passing 1.6mm. The size distribution of the accept and reject fractions is also shown. As can clearly be seen, the size distribution of the waste reject fraction is substantially coarser than the sulphide fraction. Waste d5o=26mm, sulphide d5o=3.3mm. Overall, the size distribution appears to be bimodal due to the different fragmentation patterns of ore and waste fractions in the round. The size distribution of the waste fraction is normal, with a peak around the +19mm fraction, while the size distribution of the sulphides is heavily biased towards the fines. 57 153-4550-1 Size Distribution Log Size (mm) Figure 14-153 ore size distribution; whole ore fraction, sinks and floats There is a clear increase in Cu grade and weight distribution, and therefore metal distribution towards the finer fractions in the sample (Figure 15). Cu and PGM grades increase in the -3.3mm fraction. There is a good correlation between Cu and total precious metals (TPM) distribution in the sample. The +125mm fraction, comprising 3% by mass and 0.4% metal by weight can be considered barren. Ni grades are consistently low (<1%) in all fractions. The weighted average grade of the 153 sample taken was 13.26 % Cu, 0.39% Ni, and 14.81g/t TPM. 153 Size/Grade Distribution Log Size (mm) Figure 15-153 grade distribution by size 58 The metal sulphides observed in the 153 ore are almost exclusively chalcopyrite (Figure 16). The chalcopyrite is massive, and exhibits regular parallel fracture planes at approximately 3mm spacing, which possibly contributes to the characteristic fragmentation behaviour of the sulphides as noted earlier. Disseminated pentlandite can sometimes be seen in the chalcopyrite, however this is not common. Platinum group metals sometimes occur in the ore as sperrylite, however this occurrence is not analogous with the TPM grades noted for this sample. It is felt that Ni and platinum group metals may occur in solid solution in the chalcopyrite, accounting for the high grade of the TPM in particular (Naldrett, 1984; Coats & Snajidr, 1984). Figure 16 - 153 4550 Ore fraction (x 50) showing chalcopyrite with characteristic striations Some dissemination of ore into the waste can be observed at the contact to a limit of between 5 and 10mm (Figure 17). Similarly footwall is slightly disseminated into the ore at the contact. However, sulphides typically break cleanly at the contact on fragmentation, and the distribution of metal sulphides in the footwall particles from attachment or as disseminated sulphides is insignificant (Figure 18). 59 Figure 18- 153 4550 waste fraction (x 50) showing quartzite embedded in Sudbury Breccia. Note the absence of disseminated sulphides in the waste Previous size and grade distribution testwork had been conducted by INCO (Buksa & Paventi, 2001). Liberation and separability testwork was not performed. The experimental 60 results have been compared to previous results and orebody data on actual in-situ grades and estimated dilutions of the 153 and MOB for accuracy; the characteristics shown above are consistent with these results, and compare well with values from the Mine Plan and can be assumed to be reasonably representative. The 153 sample was of higher Cu and TPM grade overall, with lower levels of dilution than the average 153 ROM ore (see Table 6). The size distribution based on visual observations and previous INCO testwork is representative. A summary of the relevant ore properties determined by the testwork is presented in Table 7. Metallurgical results from the heavy liquid separation testwork that was undertaken on the ROM samples are presented in Table 8. Separation was performed at a cut density of 3 using methylene iodide in trichloroethane. Table 7-153 ore properties Property 153 Topsize 200mm dso 12mm Ore density 3.19 t/m3 Sulphide density 3.81 t/m3 Waste density 2.69 t/m3 Liberation Index 90% Concentration Criteria 2.61 Table 8 - Heavy Liquid Separation Results @ SG 3.0 Product Cu Ni T P M Wt% Grade Wt% Grade Wt% Grade Wt% 153 Ore 100 13.26% 100 0.39% 100 14.81 g/t 100 Cones 45 29% 98 0.8% 91 35g/t 93.3 Waste 55 0.5% 2 0.06% 9 1.43 g/t 6.7 The heavy liquid separation results on the 153 ore indicate that 55% of the ore could be rejected at Cu, Ni and PGM recoveries of 97.9%, 91.1% and 93.3% respectively. 61 .2 MOB Ore Evaluation A 500kg MOB stope sample was taken using a small loader-excavator from #3 Panel of the Main Orebody on 3575 level. Again, ore was screened for size analysis, and the respective size fractions separated using heavy liquid separation techniques and the sink/float fractions sent for assay. The size/grade characteristics of the ore were reconstructed from this data. Micrographs of selected ore and waste particles were taken using an optical microscope. The MOB ore appears to fragment to a much coarser particle size than the 153 ore (Figure 19). Particles >500mm which are possibly barren were observed in the stope, and 500mm static grizzlies are typically installed at the orepasses to prevent these particles from entering the muck circuit. However, these known +5 00mm particles were not observed in the sample, probably due to small size of excavator bucket on the loader-excavator. Figure 19 - MOB round 62 Liberation in the MOB ore is fair, but it is not as well liberated as the 153. Also, the sulphides are not visually distinguishable from the waste (Figure 20), and more sophisticated sorting methods such as conductivity or magnetic separation will be required to effectively sort this ore. The ore responded well to heavy liquid separation, although at the natural size distribution of the stope sample, there was a substantial amount of middlings particles in the sinks fraction. While this is not a major concern from the perspective of simple waste rejection, further comminution of the MOB ore would be required to improve liberation, and therefore increase the degree of pre-concentration through sorting or dense-media separation methods. Figure 20 - MOB 3575 +6mm fraction A single +125mm particle was observed in the sample (Figure 21). The particle exhibited disseminated chalcopyrite in the rock which is borne out in the assay of this rock, showing 3% Cu. However, this dissemination seems limited to waste rock in the contact zone between the massive sulphide lenses. It is still felt that the grade decreases significantly with increasing particle size, however the presence of metal values in this particle belies this observation. As noted previously, 'barren' +500mm particles were not captured in the sampling process. 63 Figure 21 - MOB 3575 +125mm fraction showing chalcopyrite mineralization The size distribution of the stope sample is shown in Figure 22. The topsize was observed as 250mm, d5o=19mm, with 12% passing 1.6mm. The size distributions of the accept and reject fractions are also shown. Again, the size distribution appears to be bimodal, where the size distribution of the waste reject fraction is notably coarser than the sulphide fraction. Waste d50=38mm, sulphide d5o=18mm. MOB-3575-3 Size Distribution Log Size (mm) Figure22 - MOB ore size distribution: whole ore fraction, sinks and floats 64 There is a distinct increase in grade towards the finer fractions (Figure 23). Weight and grade distribution increase in the -1.6mm fraction, leading to a high metal distribution in this size range. Grade falls significantly in the +53mm, +75 and +125mm fractions. The +125mm fraction assayed at 3.4% copper, which may not be representative as there was only one such particle in the sample. However, this fraction cannot be considered sufficiently low-grade to be rejected as waste. The weighted metal grades of the sample were 1.22% Cu, 2.43 % Ni , and 0.44 g/t TPMs which is higher than the value contained in the MRI estimate for this orebody. MOB Size/Grade Size (mm) Figure 23 - MOB grade distribution by size Examination of the MOB ore fraction under the microscope revealed a complex massive sulphide structure (Figure 24). Pentlandite (Light cream), Pyrrhotite (Beige), Silicates (Dark Blue/Grey) and Magnetite (Black) can be seen. This is typical of a North Range Main ore zone ore (Naldrett, 1984). 65 Figure 24 - M O B ore (x50) showing pentlandite, pyrrhotite, silica and magnetite Examination of a middlings particle reveals some dissemination of sulphides into the host rock (Figure 25). There is a concentration of magnetite at the interface, rapidly transforming into silicates and disseminated sulphides, mainly pentlandite and chalcopyrite away from the interface (Figure 26). This dissemination of sulphides into the host rock diminishes rapidly over distance from the mineralised zone. 66 Figure 25 - MOB contact (x 50) showing dissemination of sulphides into waste and waste into sulphides A summary of the relevant MOB ore properties as determined by the testwork is presented in Table 9. Table 9 - MOB Ore Properties Property MOB Topsize 250mm d50 19mm Ore density 3.86 t/m j Sulphide density 4.1 t/mJ Waste density 3.09 t/mJ Liberation Index 52% Concentration Criteria 3.2 Again, the ore characteristics shown above are consistent with previous testwork results and compare well with values from the Mine Plan and can thus be assumed to be reasonably representative. For the MOB size distribution the sample screened finer than a representative stope sample. The sample was collected from the stope by loader-excavator and the bucket size may not have been sufficient to sample the known +500mm fraction in the ore. The MOB size distribution is thus considered conservative, and must be corrected upwards for accuracy. A more realistic topsize of 500mm is indicated. The grade of the MOB 4550 is somewhat higher than the grade used in the MRI and life-of-mine plan, but is within a reasonable grade range of this value. Again, the dilution of a true MOB ROM sample would be higher than tested. Metallurgical results from the heavy liquid separation testwork that was undertaken on the ROM samples are presented in Table 10. Separation was performed at an SG of 3 using methylene iodide in trichloroethane. Table 10 - Heavy liquid separation results @ SG 3 I 1 1 . 2**L Product Cu Ni T P M Wt% Grade Wt% Grade Wt% Grade Wt% MOB Ore 100 1.11% 100 2.54% 100 0.44 g/t 100 Cones 78 1.32% 85 3.02% 97.3 1.1 g/t 90.2 Waste 22 - 0.9% 15 0.3% 2.7 0.15 g/t 9.8 These results are based on bench-scale heavy-liquid separation testwork. Metal recoveries overall are better than 95%. Cu recovery in the MOB sample is poor, and is attributed to some dissemination of the chalcopyrite into the interstitial gangue. Platinum group metals also 68 appear to be associated with the chalcopyrite in this ore as PGM recoveries are similarly low. Some dissemination of pentlandite in the MOB footwall ore was also observed, but this is not reflected in the Ni recoveries. Metallurgical recovery and the degree of waste rejection would be substantially increased through improving the liberation of the MOB ore through further comminution. 2.3.3 Summary of Testwork Results In analysing the results of the mineralogical evaluation, the following processes are suggested for good metallurgical results: 153 Orebody • Screen out +125mm fraction. • Crush to -75mm to improve liberation. • Screen out -9mm fraction to concentrates. • Separate -75 + 9mm fraction into ore and waste fractions, • Combine -75 + 9mm cones with -9mm fraction. • Reject -75 + 9mm waste. MOB Orebody • Screen out +125mm (+250mm) fraction. • Crush to -75mm to improve liberation. • Screen out -19mm fraction to concentrates. • Separate -75 + 19mm fraction into ore and waste fractions, • Combine cones with -19mm fraction. • Reject -75 + 19mm waste. In both cases separation could be based on the discrepancies in colour, density and metal content between the sulphides and the waste observed in the mineralogical testwork. Based on the size/grade distribution of the MOB ore, it was felt that best liberation of the ore would occur below 26mm, and that comminution of the ore to below -75mm may improve 69 the metallurgical results. Further testwork will be required to explore the optimum liberation size of the MOB ore. The substantial differential between the work index of the ore and waste in the 153 was noted from the beginning of the study. This was confirmed throughout the testwork campaign, as well as in the tendency of the sulphides to preferentially break down further during handling in the muck circuit at McCreedy. This property has been identified as a potential mechanism to separate out high-grade material from low grade waste purely on the basis of size. This method of auto-concentration of the 153 ore should be investigated further. 70 2.4 Integrated Mining and Processing System Design 2.4.1 Design Basis Four scenarios were identified during the course of the fieldwork, and evaluated using preliminary numbers in order to identify the preferred option. It had been previously identified that there was a possibility of meeting the smelter feed grade specification through pre-concentration of the 153 ore, which has been confirmed in the bench-scale testing undertaken at U B C . Preliminary scoping work at McCreedy East had been executed in early 2004, and it was determined that due to the substantial grade and tonnage differences in the mining of the 153 and Main orebodies, that the best metallurgical and economic benefits would accrue from processing the ores separately. Appropriate coarse particle separation technologies had been identified through previous research and are listed in Table 11 (Sivamohan & Forssberg 1991). Table 11 - Evaluation of Coarse Particle Separation Technologies Technology Property Comment Optical / Colorimetric Sorting Colour / Reflectivity Compact, robust, low tonnage Conductivity Sorter Conductivity Compact, robust, low tonnage X-Ray / Laser / U V Sorting Fluorescence Compact, robust, low tonnage X-ray Permeability (u) Compact, robust, low tonnage Radiometric Sorting Radioactivity Ore must be radioactive, limited to gold/ uranium Gamma Neutron Activation Expensive, low capacity Magnetic Separators Magnetism / Paramagnetism Not sensitive enough for low-grade ores Differential microwave heating Temperature Unproven technology Gravity Separation Density Limited mass pull to waste Dense Media Separation Density Compact, robust, high tonnage Scalping / screening Size Compact, robust, high tonnage, but not selective 71 Coarse-particle flotation has also been identified as a possible technology (Lloyd, 1979), but this concept will require substantial further development and testing. Optical sorting was successful on the 153 ore, but due to poor visible distinctions between the MOB ore and waste, optical sorting did not give good results on this ore. Conversely, conductivity sorting had performed with good metallurgical recovery on other Sudbury nickel ores, but was not as successful on the chalcopyrite ore. High Intensity Rare Earth Magnetic separation had been tested on nickel ores previously at Creighton and Stobie, and returned good metallurgical results. Amenability testing on MOB and 153 ores showed good responses for optical, magnetic and conductivity sensors, but fluorescent responses showed a poor correlation with grade (Figure 27). Sample Figure 27 - Nickel Sulphide Ore Conductivity and Magnetic Responses (see Appendix 3) Heavy liquid separation gave better metallurgical results than previous optical sorting testwork on the 153 ore, sufficient to warrant the consideration of dense media separation for both ores. This technology is widely used in the pre-concentration of ores on surface (Bappu et al, 1999). Coarse particle flotation was rejected as the size/liberation curve suggested that ore in the coarser size fractions (+53 & +75mm) was sufficiently well liberated to be separated at or near 72 this size range. A finer optimum liberation size of approximately 3mm and finer would necessitate consideration of coarse particle flotation. Previous testwork by INCO and U B C Mining has indicated that a Mogenson combination optical/conductivity sorter demonstrates good metallurgical results on McCreedy ores, and will achieve the process specification stated in the model (Buksa & Paventi 2002). Rare-earth magnetic separation was also shown to be effective in separating massive nickel sulphide ores at other mines (Exportech, 2002). However, this equipment has not been tested on M O B ore specifically. The combined conductivity sorter / magnetic separator would be installed on the head end of a suitably designed feed conveyor, where ore and waste would be diverted into separate chutes on discharge from the conveyor. The Mogenson AH-80 sorter operates optimally at a throughput of 80tph over a size range of 20 - 80mm and up to 200tph over a size range of 80 - 200mm, thus probably two units would be required. On consideration of the testing results, and out of a desire to explore diverse technologies in the design, dense media separation was selected as the preferred process technology for separation of the 153 ore, while a combination of conductivity sorting and magnetic separation would give good results in the separation of the M O B ore. Fundamental design criteria for underground process plants have been developed in the literature review. Further operating design criteria for the respective plants were assembled from experience and the background information obtained during the site visits, o Processes should be as compact as possible o Plant should be robust and flexible in the face of variable feed conditions, particle size, tonnage and grade o Processing should occur as early in the mining cycle as possible o Processing should be at as coarse a size as possible while achieving preferably >95% metallurgical recovery o A maximum of 60% of the ore by mass only could be rejected due to backfill space constraints incumbent through the bulking factor associated with broken ore. 73 o The process plant should integrate well into the existing underground mining and material handling layout o Dry processes are preferred over wet o Open circuit processes are preferable to closed o Process equipment should be suitable for such situations of heat, noise and dust as found in the underground environment o Process plant is a source of high levels of heat, noise and dust - the process plant should be situated on an independent ventilation circuit or directly on a Return Air Way in order to limit the impact of this on air quality in other working sections of the mine o A bypass function should be inherent in the design in order to avoid mining stoppages associated with adverse plant availability o Surge capacity in the feed and product areas of the plant should be maximised within the cost and space constraints of the situation o Plant situation should be above the upper horizon of the orebody in order to facilitate the backfilling function o Individual plant excavations should be separated by a distance of > 1 x maximum dimension to meet basic rock mechanics criteria (Hartmann, 1992) Additional operating design criteria developed for the 153 and M O B underground pre-concentration plants are summarized in Table 12. 74 Table 12 - Pre-concentration Plant Operating Design Criteria F a c t o r U n i t 153 M O B Ore type Massive-stringer chalcopyrite Massive & disseminated base metal sulphides Mining rate tpd 1800 3200 Dilution (max) % 70 30 Ore UCS GPa 15 20 Waste UCS GPa 80-100 80-100 Plant Operation Shifts / days 2 x 12, 7 days 2 x 12, 7 days Availability % 85 85 Surge Factor % 35 35 Plant Feed Rate (max) tph 120 230 Feed size mm 250 500 Plant feed storage capacity h 8 8 Crusher feed rate tph 6 24 Crusher f80 mm 125 250 Screen products High grade fines tph 61.3 115 Fine sorting tph n/a 41.5 Coarse sorting tph 58.4 57.5 Max waste stream tph 50.24 33.5 Waste storage capacity h 12 12 Max product stream tph 69.5 180 Product storage capacity h 1 5 Total water requirement m J/h 38 12 Unit process flowsheets and process descriptions for each of the underground pre-concentration systems are described below. 75 2.4.2 Mining and Processing System Description PFD 100: 153 Mining The process flow diagram in Figure 28 has reference. Ore is mined at an average of 100 tph using mechanized underhand drift-and-fill techniques suitable for narrow-vein mining. Headings are drilled and blasted, and mucked by L H D to an orepass on the mining level. It is proposed to increase the minimum panel dimensions from 8ft x 9ft to 12ft x 12ft in order to facilitate jumbo drilling and automated roof bolting. This will also enable the utilization of higher capacity 6- and 8-cubic yard LHD's in the narrow-vein areas. Dilution from these narrow vein panels will also be increased, however, present and excess levels of dilution would be eliminated at the pre-concentrator. The orepass is equipped with an inclined static grizzly with an aperture of 500mm. Barren oversize material reports to the rockfill pile situated next to the orepass, and is remucked to the stopes as rockfill; an additional rockfill L H D has been allowed for in the design, although simulation of the mining, processing and backfill system indicates that the rockfill function could be adequately catered for by the existing L H D suite on a level. Rounds that are identified immediately as waste will be mucked to rockfill as in the present method. The balance of fill is introduced from surface using the existing sandfill system. Grizzly undersize reports to the orepass and related Kiruna truck loading chute, where it is hauled to a lOOOt pre-concentration feed silo to be developed at on 3575 level at Location 8. Should the pre-concentration plant be unavailable, the system can be bypassed by delivering ore directly to Location 1 or 2 as in the present mining scenario. 76 ProductiorL. . Stope Fill stope Production tope Fill stope Production Stope Fill stope 3* 500mm vibrating grizzly 20 Spray water Scj^ pptram +500mm Waste 4360 Orepass^  - 1 3 -Spray water Scooptram Scooptram 500mm vibrating grizzly 4660 Orepass Spray water Scooptram FT 500mm vibrating grizzly 4810 Orepass - 1 5 -Kiruna Truck Spray water 500mm vibrating grizzly 4945 Orepas: _ 1 2 Scooptram Production Stope 19 Location 8] Ore Silo Scooptram +500mm Waste Fill stope. Ore to Pre-concentrator 1 8 Spray water 500mm vibrating| grizzly -11-5130 <p repass) Scooptram Scooptram Production Stope,— - Fill stope Spray water 500mm vibrating grizzly Scooptram Production Stope/---10-Future 170 Orepass +500mm Waste Scooptram ilfSTi T | J Fill stope Figure 2 8 - 153 Mining Process Flow diagram 77 PFD 200 -153 pre-concentration The process flow diagram in Figure 29 has reference. Ore is extracted from a new lOOOt feed silo at Location 8 via a vibrating grizzly feeder at lOOtph and screened at -53mm. Grizzly oversize reports to an 800 x 600mm jaw crusher for reduction to -75mm to prepare feed for the Dense Media Separation (DMS) drum and to improve liberation. Grizzly underflow is combined with the crusher discharge and transferred via belt conveyor to the screening plant. Ore is wet screened of -9mm fines in preparation for feed to the DMS drum; wet screening was selected in order to efficiently remove the adhering fines from the coarser particles and thus achieve maximum recovery of the fine, high-grade sulphides. The choice of wet ore preparation, a wet DMS separation process, and the hydraflush crushing technology will also help to allay the dust generated underground during processing. -9mm high-grade undersize material reports directly to concentrate at 50tph, the. -75 + 9mm material is fed at 100 tph (max) to the DMS drum, operating on FeSi medium maintained at an SG of 3. In selecting a drum in preference to a dense medium cyclone for this duty, it was felt that the drum was better suited to accept variations in feed size and tonnage, and would require less rigorous feed preparation. D M cyclones are superior in processing finer size fractions at constant feed rates. However, copper and precious metal losses to fines are a concern, hence fines are screened out directly to concentrate, and the drum processes only the -75 + 9mm fraction. DMS sinks and floats report to a split-deck drain-and-rinse screen for recovery of the FeSi. DMS floats report to a high-angle conveyor which delivers waste rock to the existing 3770 level rockfill pass at Location 9 via a newly constructed conveyor drift. DMS sinks are crushed to 100% -15mm in the ore crusher and report to the final concentrate conveyor, where it is combined with screen undersize and diluted with water to 60%> solids in the shaft-bottom tank prior to hydraulic hoisting. Metallurgical recovery is projected to be better than 95%. A relatively simple medium circuit comprising a make-up system, correct- and dilute medium tanks is shown in the design. FeSi recovery is by means of the drain and rinse screen, passing underflow to a double-drum magnetic separator. Mag-Sep underflow is pumped to a pipe densifier for re-densification to approximately SG 3; under-density can be corrected by supplying overdense FeSi slurry from the make-up tank. The medium circuit is compact and simple to operate. 78 Primary Crushe Flushing water 6 -9mm high grade fines Bin Q D&R Screen Feeder Concentrate Crusher 13 Overflow-Concentrate Conveyor 14-3380 Level Rockpass Shaft Bottom Tank Surface Cone Thickener -16 return water-Surface tank 15 Concentrate Tanker to Smelter Figure 2 9 - 153 Pre-concentration Plant Process Flow Diagram 79 PFD 300 - Concentrate Hydraulic Hoisting The process flow diagram in Figure 31 has reference. Hydraulic hoisting has been selected due to the sticky nature of the DMS pre-concentrate, and a desire to keep the resultant high grade material out of the conventional ore handling system. Concentrate is crushed from -75mm to -15mm in a 1356 hydraflush cone crusher. This crusher has been selected due to the improved efficiency of wet crushing, the dust-allaying action of the flushing water, and the opportunity to produce a finely-graded, high solids-content slurry directly in the crusher. Based on the expected natural concentrate size analysis, and a closed side set of 8mm on the crusher, the expected product size distribution is shown in Figure 30. It can clearly be seen that the particle size distribution of the feed is already fine, and finer comminution than projected would be quite feasible. Crusher underflow is passed to the shaft-bottom tank, and kept in suspension by a high-speed abrasion-resistant repulping agitator. 20 10 0 1 10 100 Size (mm) Figure 3 0 - 153 Concentrate Projected Size Analysis 80 -Make-up water - -6 Thickener O/flow-t t Overflow-Process water tank To DMS Plant-Flushing water -Cones from DMS ^ 2 n d stage Turbine EMF Concentrate Crusher 1sl stage Turbine 'EMF Mid-shaft Tank jr -4 Overflow -Dump line -Shaft Bottom Tank 2 n d stage pumps ^2 Dump line -Overflow to shaft bottom dam 1 s t stage pumps Shaft bottom dam Surface tank Surface Cone Thickener 3 r d stage pumps U/F pump Concentrate Tanker to Smelter Figure 31- Concentrate Hydraulic Hoisting 81 Concentrate slurry is pumped to surface at 42.3 m lh via existing ramps and shafts using a multi-stage pumping system. The pumping system comprises three stages of pumping, with standby pumps, separated by agitated break-tanks similar to the shaft-bottom tank. Gravity surge lines, sized for full flow return, overflow from the receiving tank back to the delivery tank. There is also a provision for a dump line in case of pump failure or line blockage which delivers the pumped concentrate back to the delivery tank. This is a pumping arrangement that has been successful in long distance overland tailings delivery lines. Should the shaft bottom tank overflow, the slurry is captured in a bunker where it would be dewatered and returned to the system by bobcat or 1.5 cu. yd LHD. There are a number of pumping options: • Multiple stages of high pressure, all-metal centrifugal slurry pumps. This would entail pumping the concentrate at a consistency suitable for this type of pump, probably 30-40% solids max. Increasing the solids content may be facilitated by modifying the rheology of the slurry through further comminution; this must be explored in future research. Hydraulic hoisting is by nature less energy efficient than dry hoisting due to the requirement to pump the transport fluid as well as the ore against gravity, which requires additional pumping effort - increasing the solids content will reduce this inefficiency. A further means of addressing the inefficiency of hydraulic hoisting would be to recover the energy from the inflowing process water prior to use in the pre-concentration facility. This would avoid costly pressure-reducing valves in the process water downcomer, and offset the energy required to pump the water back up the shaft as part of the slurry. This energy recovery can be achieved through a number of methods. In the centrifugal pump scenario, a second set of centrifugal pumps could be installed on the process water delivery line and mechanical effort transferred directly to the delivery pumps by means of a shaft. Hydraulic energy could also be recovered by reverse-running pumps, driving an A C motor/generator which drives the delivery pumps or alternately feeds A C back into the underground power circuit for general use. • Multistage centrifugal pumps are also attractive, and have been used in dewatering applications in deep mines for a number of years. The pumps are energy efficient and also have energy recovery capabilities when running in reverse as clear-water turbines. A good example of this type of pump is the Sulzer HPL-range of multistage pumps. These pumps typically pump up to 1300 m 3/h of dirty water with a low solids-content (<10%) 82 over heads of 200 - 500m. The pumps are limited in this application by two factors: a requirement for low feed solids content, and a limitation on the feed particle size the pump will accept. Some development work with the manufacturer on sealing arrangements, pump clearances and materials of construction will be required to develop this pump for the hydraulic hoisting application. • Several other hydraulic hoisting technologies have been reviewed from the literature. Vaal Reefs Gold Mine in South Africa operates a low solids density ore hoisting system using Mars-type positive displacement pumps to hoist ore over 2500m to surface (Schutter, 1977). INCO have investigated both centrifugal pumping and Siemag positive displacement pumping options for hoisting ore some 1200m at Copper Cliff North Mine*. The CSIR in South Africa is experimenting with high-solids content pumping using the Tore® hydraulic transport system, though while the pumps are sanding-resistant, there are severe limitations in the head that can be pumped using such technology. • A further alternative for the separate hoisting of 153 concentrates to surface is vertical conveying. High-angle and vertical conveyors have been used in hard rock applications up to 1500 tph and 500m vertical lift. Vertical conveying has also been shown to be superior in terms of capital and operating cost when compared to conventional conveying over the same vertical distance (Yester, 1997). However, the pocket belts are difficult to clean, and present a challenge when handling clays and other sticky materials. The use of such conveyors is still attractive, however, and high-angle conveying has been included as part of the system design for the handling of waste from the pre-concentration plants. • A n attractive option for hydraulic hoisting is the large-calibre positive displacement pump. Multi-piston Gardner-Denver PD pumps have been used in high-pressure mud pumping for a number of years. These pumps are also utilised in concrete aggregate pumping as discussed later in this paper. Pump capacity is high, and delivery pressure can be as high as 17 MPa. Increasing the size and solids content of the slurry to be pumped negatively affects the performance of the pump, and also has a deleterious effect on wear, particularly of the sliding surfaces. However, the performance of the pump is so superior to that which is required in the multistage pumping arrangement shown in the * INCO General Engineering Department Estimate 2409-OA, 1991 83 design that these limitations could be overcome. In addition to this, the chalcopyrite concentrate does not pose a particular wear challenge in this instance due to the extreme softness and friability of the ore. It is proposed to explore the application of these pumps to the hydraulic hoisting concept, thus these pumps have been included in the conceptual design. The pumps are not suitable for energy recovery, thus pumps similar to the Sulzer HPL series are recommended for this application. The pre-concentrate is dewatered on surface via a 16m diameter high-rate thickener and delivered directly to the smelter as converter cold charge feed by concentrate truck. Thickener overflow is clarified and gravitated underground to the process water tank at the DMS plant location via the energy recovery pumps. PFD 400 - MOB Mining The process flow diagram in Figure 32 has reference. The M O B and L M O B are mined by mechanized overhand cut and fill methods at a rate of 200 tph max. 6m x 18m panels are drilled and blasted, and mucked to a remuck pile by L H D . The +500mm oversize is scalped at the orepass by vibrating grizzly as in the 153 scenario; oversize is hauled to fill stopes and grizzly underflow reports to the orepass for collection by Kiruna truck via Kiruna chute. Ore is hauled to a new feed bin developed on the haul ramp between Location 8 and Location 7. Pre-concentration of the M O B / L M O B ore is discussed in the following section; after pre-concentration, ore is delivered to a new surge bin delivering ore to a Kiruna chute located on the haul ramp. Pre-concentrated ore is collected from here by Kiruna truck and hauled 900m to Location 1 or 2 as in the existing arrangement. 153 and M O B pre-concentration rejects are delivered by high-angle conveyor to the existing rockfill silo at Location 9. Again, an additional rockfill L H D has been allowed for to cater for the increased quantity of rockfill produced in the new scenario. Should the pre-concentrator become unavailable for any reason, the system can be bypassed completely and trucks would deliver -500mm ore to Location 1 or 2 as per normal. 84 Production^ Stope Scooptram MmW£®m 500mm / vibrating grizzly -11 Spray watei^ -Kiruna Truck Scooptram 500mm / vibrating grizzly 4090 Orepass Fill stope X x+500m Waste Production Stope Scooptram Scooptram 500mm / vibrating grizzly Fill stope X Future pre pas s| ,+500m Waste Ore to _ Pre-concentrator Waste from _ ^pre-concentrator 3770 ftockpass) Location 1 Rock Silo Kiruna Truck 9 (Optional) Waste from 153 DMS Plant --Spray water-F i g u r e 3 2 - M O B M i n i n g P r o c e s s F l o w D i a g r a m 85 PFD 500 M O B pre-concentration, ore hauling and hoisting The process flow diagram in Figure 33 has reference. Ore is extracted from the coarse ore bin via a vibrating grizzly feeder and screened at -53mm. Grizzly oversize reports to the existing 1200 x 1000mm jaw crusher for crushing to -75mm. Crushed material is wet screened into -75 + 38mm and -38 + 15mm fractions at 100 tph max respectively. -15mm high-grade undersize material reports directly to concentrate. -75 + 38mm is sorted at 58 tph in a Mogenson Model AH-80 combined conductivity / magnetic sorter comprising a 1000mm flat feed belt fitted with a 6800 Gauss magnetic head pulley moving at lm/s feeding ore to the sorter mounted in the head chute of the conveyor. Ore is sorted by means of a combination of pneumatic ejection of the waste particles based on the conductivity response and magnetic deflection of the particles by the rare-earth magnet. Similarly the -38mm + 19mm fraction is sorted at 42 tph using a Mogenson AH-80 combined conductivity / magnetic sorter. Metallurgical recovery is also expected to be of the order of 95%. Concentrated ore is transferred at 200 tph (max) via high-angle pocket belt conveyor to a new loading bin constructed between the 3370 and 3450 level on the main haul ramp. From here ore is collected via a new Kiruna chute by 40t Kiruna truck and delivered to the ore bins at Location 1 & 2, and hoisted conventionally to surface using the existing ore handling system. The hoe-ram and crusher can be eliminated due to the pre-processing of the ore by the concentration plant. Waste is transferred by high-angle conveyor from the sorters to the existing rockfill silo at Location 9, where it is combined with the 153 rejects. 86 F i g u r e 33 - M O B P r e - c o n c e n t r a t i o n P l a n t 8 7 2.4.3 Waste Disposal and Backfill: PFD's 100 & 400 Both mining methods employed at McCreedy require fill. Rounds in both the 153 and MOB are sometimes considered waste rounds and are mucked to waste piles for use directly as rockfill. Due to differences in the mining methods, there is a surplus of rock in the 153 stopes and a shortage of rock fill in the MOB stopes. Stopes that have been adequately rockfilled are barricaded and filled with sandfill which is prepared on surface at a cost of $4/ROM ton. At the present mining rate of 165 tph, the sandfill requirement is on average 104 tph. The backfill system proposed in this report integrates well with the existing rockfill system at McCreedy. The two basic rockfill systems are described below: 153 Orebody DMS rejects at -75 + 9mm and 40 tph are discharged from the drain-and-rinse screen onto an inclined conveyor for transfer to the existing 9001 waste-rock silo at Location 9. A specially developed inclined drift will be required for this conveyor between the bottom of Location 8 on 3860 level and the top of Location 9 at 3770 level. Waste is mucked by LHD from the existing rockfill draw-point on 3860, and dumped to fill stopes in the MOB prior to sandfilling from surface. Should the metal losses to waste not be acceptable due to the extremely high feed-grade of the material, 153 DMS waste could be hauled from the 3860 rockpass to Coleman shaft for processing together with the MOB ore. MOB Orebody Waste at 100 tph is transferred to the existing lOOOt waste silo at Location 9 via a specially constructed conveyor drift. Fill is collected by the existing LHD's for placement in finished panels prior to final sandfilling from surface. Should it be required to handle 153 waste additionally, feed, waste and ore design tonnages would be 240 tph, 140 tph and 240 tph respectively. A maximum of 140 tph of solid waste is generated by the pre-concentration plants at their respective locations. The waste is typically ultramafic diorite or Sudbury breccia with small concentrations of disseminated sulphides, but is not expected to have a high ARD potential. Sudbury breccia has a UCS of between 100 and 200 MPa. The expected size distribution of the 153 and MOB waste fractions is shown in Figure 34. 88 153 & MOB Ore and Waste Size Distribution Size (mm) Figure 34 - Expected Ore & Waste Fraction Size Distributions At present it is proposed to introduce the additional waste into the mining cycle as fill, utilizing the rockfill system that is currently in place. Stopes that have been rockfilled are then barricaded prior to final filling with sandfill prepared and pumped from surface. The present sandfill plant produces a finely-graded fill at an approximate UCS of l-2MPa, which is typical for this kind of fill. At the present depth of mining, more competent fill is not required. The additional waste is expected to reduce, if not eliminate, the quantity of sandfill presently required. However, when considering the high void ratio of rockfill, this may not be physically possible due to the quantity of additional waste that is generated in the process. As discussed previously, high-strength backfill is a crucial element of the support system envisaged for an ultra-deep mine. The UCS required of such a backfill may be of the order of 10 - 20 MPa. It is presently not considered feasible to produce backfill of such quality economically. A means of overcoming the high void ratio of traditional coarse rockfill, and deliver a more competent backfill in the deeper mining scenario would be to utilize the waste rejects from the underground pre-concentration process to provide the aggregate fraction and part of the sand fraction for a backfill with similar specifications to a conventional concrete aggregate mix. The backfill plant would be similar to a conventional automated concrete batching plant fed from the rockpass at Location 9. Coarse waste rock would be crushed to -38mm and screened of the -3mm fines. Coarse 89 waste material is used as aggregate and fine waste material is combined with supplementary mill tailings from surface in order to correct the mix ratio. Backfill is pumped to the fill stopes using positive displacement concrete pump. (Figure 35). Based on past experience, a starting mix ratio is suggested to be 12 parts coarse aggregate : 4 parts sand/fines : 2 parts water: 1 part cement for a backfill strength of approximately 10 MPa. Residual sulphides in the waste will be isolated from potential ARD contacts through fixing of sulphides in the cement. -68 tph rockfill 816 t rock silo Cemented backfill to stopes From surface 6.8 tph I Sand/ 3.4 tph Cement Tailings I t I 100t sand bin 6.8 m3/h Water I 100m3 Process water t a n k ^ Flushing water Batch mixer h -85 tph cemented backfill PD Concrete Pump Figure 35 - Backfill Preparation System Backfill thus batched can be pumped to fill stopes using heavy-duty concrete pumps. A schematic of a typical positive displacement concrete pump is shown in Figure 35. Such pumps are capable of delivering -38mm aggregate concrete mix at 52 m3/hr (100 tph @50% solids) over 400m horizontal, and 240m vertical distance. 90 5" delivery Hue Figure 36 - Thomas Katts BS 907A Concrete pump schematic Should this cemented backfill delivery system prove effective, it is further suggested to alter the mining layout to mine to an apparent dip of 5-8% in order to facilitate the flow of the cemented backfill to the stopes. 2.4.4 System Integration An understanding of the underground layout at McCreedy was developed from site visits, as well as plans and sections obtained during fieldwork. The simplified section through the mine presented in the introduction is shown in Figure 37. Several locations were identified for the construction of the respective process plants. It was imperative to integrate the process system with the existing material handling system in the mine i.e. 40t Kiruna haul trucks using the haul ramp. It was decided to locate the plants somewhere on the haul ramp above the 153 orebody, and away from the Coleman shaft. Areas which meet a number of the basic criteria for the underground mining and processing system as mentioned in the literature review were: 91 o existing and planned excavations around Location 8 and 9 between 3440 and 3550 level o a combination of existing and new excavations around the 3220 level hoe-ram station. o a system of new excavations situated near the haul ramp Based on its central location on the haul ramp, proximity to the 153 orebody, and the availability of the existing 3697 rockpass at Location 9 to receive the waste rejects, the existing ramp and drift complex on the haul ramp between the 3575 and 3695 levels was selected. Figure 37 - Simplified McCreedy section 92 Basic single-line layouts were developed for each of the process alternatives mentioned above. The conceptual plant design will be laid out within the existing McCreedy East mine layout, in order to confirm the constructability of the underground pre-concentration plant, and its ease of integration into the mining process. Any additional requirements for new capital excavations and any special ground control measures which may be required are identified and costed at this stage. It is especially important to confirm that the vertical height requirements of the process plant can be accommodated within the existing established levels of the mine, and that the requirement for additional capital development is not excessive. A modified longitudinal section of McCreedy East showing the layout of the integrated mining and processing system is shown in Figures 39 & 40. The location of the two pre-concentration facilities is shown, as well as the generalized ore, waste and sandfill flows. 93 S a n d f i l l F i g u r e 3 9 - 153 Integrated M i n i n g a n d P r o c e s s i n g S y s t e m s h o w i n g k e y loca t ions a n d ma te r i a l f l o w s 94 As can be seen, the proposed underground pre-concentration facilities integrate well into the existing mining layout. Stope design, mining method, and haul systems remain essentially unchanged. An additional excavation at tipping points is required to accommodate the static grizzly and oversize bunker. The 153 dense medium separation plant is laid out between Locations 8 & 9 on the Kiruna ramp, positioned at the upper extent of the MOB. This has three major benefits. o There is an existing 900t rock pass at Location 9 o There are numerous existing excavations between the planned rockpass at Location 8 and the existing rock pass at Location 9. o Rockfill is currently collected by LHD from the drawpoint under Location 9. o Kiruna trucks hauling 153 & 170 ore are required to haul as far as Location 8. o Pre-concentrated ore can be pumped via underground tanks, pumps and pipelines erected in existing haulages and shafts between Location 9 and surface. The MOB sorting plant is laid out between Locations 7 and 8. The site has a number of advantages: o The feed silo can be located on the existing haul ramp o The process plant can be laid out between existing levels of the mine o Waste can be delivered to the existing rockfill silo at Location 9 o A new transfer silo can be constructed above the haul ramp to deliver concentrated ore to the Kiruna trucks These particular aspects of the existing mining system at McCreedy East greatly reduce the expected cost of integrating the process plant into the mining system. 95 Sandfill F i g u r e 4 0 - M O B Integrated M i n i n g a n d P r o c e s s i n g S y s t e m L a y o u t 96 2.5 Conclusion The design and layout of an integrated underground mining and processing system for the pre-concentration of massive-sulphide ores is presented. The system comprises in-sertion scalping of the barren rock followed by transport of the ore to the pre-concentration facility located on the main haul ramp en route to the shaft. Pre-concentration is achieved by means of dense media separation at an SG of 3 for the massive chalcopyrite 153 ore, and a combination of high-tonnage conductivity sorting and magnetic separation for the pentlandite-rich MOB ore. The 153 pre-concentrate is hoisted to surface via a positive-displacement hydraulic hoisting system, with energy recovery from the downcoming process water stream. MOB pre-concentrate is hoisted to surface using the existing conventional hoist. However, due to the reduced tonnage of the ore handled by the hoist, operation on one shift may be a possibility. Both pre-concentration systems integrate well into the existing underground layout. Additional excavations are required for the crushing, screening and pre-concentration areas, as well as additional lOOOt surge bins for the feed and product silos for the plants. However, the largest single excavation required is of the order of 13.5 x 11 x 7 m for the dense media plant. There are several existing drifts and ore bins in the present layout which reduce the number of additional excavations required. Isometric block layouts of the proposed 153 and MOB pre-concentration plants are contained in Appendix 4. 97 Chapter 3 - System Impact Evaluation 3.1 Introduction A diagrammatic representation of the mining and smelting system at INCO's Sudbury Division is presented in Figure 41. The impact of pre-concentration on each stage of the mining and smelting process is discussed with reference to the overall system as described below. Physical, operational and financial impacts are evaluated against parameters for the existing operation. 3.2 Mining & Backfill 153 Orebody The drift-and-fill mining method is essentially unchanged. The mining rate remains constant and should require no additional equipment or support infrastructure. A maximum of 10% of the mining tonnage can be rejected at the orepass, and must be returned to the face by LHD. An additional rockfill LHD has been allowed in the cost estimate for each orebody to cater for this. However, waste remucking is currently a widespread practice at the mine which will be eliminated in the pre-concentration scenario, thereby increasing equipment availability, and possibly freeing up resources for the additional rockfill duty. It was observed that face dimensions in the 153 vary between 8ft x 9ft through 15ft x 16ft to open stopes in thicker veins, resulting in widely varying levels of productivity. It is proposed to eliminate the smaller headings entirely, and standardize on a minimum 12,5ft roof height, which will facilitate standardization on single-boom jumbos, 6 cu.yd LHDs and automated Maclean roof bolters. The additional dilution generated by this will be catered for by the pre-concentration plant. 98 Figure 41 - Simplified INCO Operations Flowsheet 99 MOB Orebody The mining method and mining rate is unchanged. Backfdl is currently through a combination of rockfill placed by LHD and sandfill delivered from surface. An additional 140 tph of waste is made available as rockfill at Location 9 which will reduce present sandfilling requirements of 1500-2000 tpd by an estimated 90%. It is envisaged that additional LHDs will be required to deliver the increased quantity of rockfill now generated at Location 9 to the MOB and LMOB. An additional rockfill LHD has been allowed in each orebody for this duty. The cost of the additional rockfill LHDs will be offset by the reduction in sandfill quantities as well as the reduced haul tonnages and distances, which will be discussed under material handling. Should the handling of such large quantities of coarse rockfill present an insurmountable challenge, an automated aggregate backfill batching and distribution plant as described in the system design could be constructed at Location 9, thus eliminating the rockfill LHDs altogether. The adoption of aggregate-specification backfill should result in improved ground conditions for a further reduction in seismicity and the frequency and severity of rockbursts at depth. 3.3 Underground and Surface Material Handling McCreedy employs a fleet of four 50t Kiruna haul trucks. Haul truck productivity and therefore cost varies with the square of the haul distance. Current linear haul distances from the 153 orebody to Location 2 vary between 1.9 and 3 km, and 500m vertically; distances for the 170 orebody will be substantially greater. Haul distances from orepasses in the MOB to Location 2 vary between 1 and 1.9km. The ore hauling system is impacted in a number of major ways: o Haul tons from the 153 and MOB orebodies to the pre-concentration facility are reduced by between 5% and 10% overall through the rejection of +500mm oversize at the tipping points o Ore from the 153 and 170 ores are delivered to Location 8 instead of Location 1, reducing the haul distance by 2km per return trip, vastly improving the productivity of the haul fleet. o Pre-concentrate from the 153 and 170 ores is pumped from Location 8 as a slurry to surface, thus eliminating any further need for haulage and hoisting of this ore. 100 o Ore from the MOB/LMOB is hauled a short distance to the pre-concentration plant at Location 7. Throughput subsequent to the pre-concentration at Location 8 is reduced by 20%, which will significantly reduce overall underground haul costs. o Due to present electrical supply constraints, a maximum of two Kiruna trucks can be on the isolation section between Location 6 and Location 2 simultaneously, which results in a limitation on the total haulage capacity of the system. The proposed underground processing system eliminates the need for more than two trucks on this section of the trolley line. o There is a reduction of 2500 tpd in material transported by road to Clarabelle Mil l . Concentrate from the 153 DMS plant is crushed and pumped as a slurry to surface. This system comprises a multi-stage system of agitated tanks and pumps, and relieves the present hoisting system of the requirement to hoist the high-grade copper ore. The quantity of waste rock hoisted to the surface rock dump is now limited to development waste only. There is an overall reduction in the quantity of ore hoisted to surface from 5000 tpd to 2574 tpd, which is a quarter of the design capacity of Coleman shaft. It may be possible to accommodate this production in a single hoisting shift, resulting in additional savings in fixed hoisting costs. In addition to these impacts, there is a general saving in surface material handling costs due to a 32% reduction in overall tons hoisted and transported from Coleman Shaft to the Clarabelle Mill some 60km away. 3.4 Underground Processing The integration of underground processing into the mining system impacts on every aspect of the underground environment. Additional support infrastructure such as power, control, reagents, water, consumables and labour are required for the two process facilities. However, this is not a significant increase over the mining infrastructure already in place at the mine. The process plants will be a source of heat, noise and dust. Supplementary ventilation will be required in order to isolate the plant on an independent ventilation circuit. It may be possible to reroute the local ventilation in order to make the new plant site the last point in the return ventilation circuit, although this will mean occasional contamination of the plant by blasting fumes and dust. Thus, it would be preferable to provide independent ventilation to the plant site. 101 Additional underground labour has been allowed for in the working cost estimation. Due to the criticality of maintaining metallurgical performance of the process, an additional day-shift process engineer is also recommended, although this is not allowed for in the costing. Labour savings in haul, hoisting and surface complements are not accounted for in the estimate, as these impacts require further evaluation. Impacts to the underground environment include air and water quality impacts. The introduction of process water, and reagents in to the underground environment can negatively impact the water quality of the local groundwater system. Process plants underground should be located outside of any known aquifer or underground intersection of the groundwater system. Care should be taken to isolate the plant from the local underground drainage system. There is a provision in the plant design to return spillage from each process back to the process in order to contain contamination and eliminate metal losses. The plant will impact negatively on air quality local to the process plant through the generation of heat and dust. However, the provision for wet processing and an additional requirement to place the process facility on independent ventilation or on a return airway will address this impact. There is a risk that low-grade waste rejects to backfdl have some ARD potential. This will be a problem if the waste rejects are left in long term exposure to oxygen and water. Sandfilling over rockfill will reduce the exposure of the waste to oxygen underground. There is water contained in the sandfill, although this is expected to drain. More permanent amelioration of the ARD potential would be achieved through the adoption of a cemented backfill, where the cementing process would consume the moisture in the fill, and the calcareous matrix would reduce the exposure of residual sulphides to oxygen, as well as neutralizing the ARD potential. 3.5 Surface Operations Surface environmental impacts are significant. In the case of a greenfields operation, all surface facilities could be designed for an effective 32% lower throughput. McCreedy is not a greenfields operation, but surface impacts are still important. There is an overall reduction in the tonnage of ore brought to surface, which will inherently require less hoisting energy with consequent positive impacts in terms of reduced electrical consumption and therefore greenhouse gas emissions. The reduced tonnage to the mill results in a lower tonnage of tailings to the tailings impoundment. The ore delivered to the mill is lower in siliceous 102 gangue, which will significantly lower the Bond Work Index of the ore, reducing the power requirements in the comminution circuit at Clarabelle. Environmental impacts at the smelter are negligible - the quantity and grade of the concentrate delivered to the smelter is reasonably consistent with the present concentrate feed. There is an additional surface environmental benefit in the reduced quantity of sandfill which is presently mined and prepared on surface. There is a possibility that the residual sandfill requirements could be supplied from Falconbridge's Strathcona Mil l ; testwork is ongoing at present to test classified mill tailings from Strathcona at McCreedy. The outcome of the testwork is not known. Several milling and smelting impacts have been identified. Feed specs to the mill are currently of the order of 30 000 tpd at 1.4% Cu and 1.4% Ni. In the proposed design 153 and 170 ore bypasses the mill completely and is introduced directly into the smelter. A smaller quantity of higher grade MOB/LMOB ore is sent to the mill for processing, which should improve the Ni feed grade overall. The pre-concentrate also has a lower Bond Work Index, due to the rejection of the more competent waste fraction, thus crushing and grinding costs should be reduced; this has not been quantified as yet. Tailings grades are expected to remain constant while feed grades are improved, and thus metallurgical recovery of Ni should be slightly improved, offsetting the additional 5% metal loss introduced by the underground pre-concentration step. The pyrrhotite: nickel ratio is expected to be affected through the elimination fo the copper rich 153/170 ore, and provisions will have to be made to cater for this in floatation and tailings disposal at Clarabelle. In addition to the impacts at the mill, the tonnage reporting to the tailings dam is reduced by up to 140 tph, which will reduce the final size of the dam at closure. In the context of INCO operations, these impacts are somewhat diluted due the investment in a large centralized milling and tailings disposal facility. The impact on the performance of a mill dedicated to McCreedy ore only would be more significant. The 153/170 ore is in the form of a high-grade concentrate. Projected concentrate specifications have been presented in Figure 35. Present concentrate specifications at various locations specified in the fNCO Operations flowsheet is presented in Table 13. 103 Table 13 - Concentrate Specification Comparison Concentrate C u % N i % Cu:Ni 80% passing McCreedy 153 Pre-concentrate 20.2 2.15 9.4:1 8 mm Clarabelle Mil l Flotation Concentrate (flowsheet) 12.5 5 2.5:1 75 um, 11% moisture Clarabelle Mil l Flotation Concentrate - present 15 6 2.5:1 75 um, 11% moisture Voisey's Bay Concentrate (future) 12.86 5.14 2.5:1 75 um, dry Slag Buttons / Converter Cold Charge 11.3 8.69 1.3:1 25mm, dry There are a number of possible options for the introduction of this concentrate into the milling and smelting flowsheet. The concentrate exceeds the specification of the nickel/copper flotation concentrate in terms of copper and precious metals content, thus inserting the concentrate prior to this point in the flowsheet is not considered attractive. Impacts at the smelter are significant. The 153 ore is high in copper and PGM's, and metallurgical testwork has indicated that the grade of the 153/170 concentrate would be higher than the present feed grade to the smelter. Firstly, the mill is bypassed eliminating metallurgical losses due to grinding and flotation. The recovery of PGM's and Cu in the mill are typically 79% and 89% respectively. These losses will be avoided through the introduction of pre-concentration, leading to a higher metal recovery for the business overall. It is expected that the 153 concentrate will meet the specification for feed to the matte converter as cold charge, further eliminating metal losses to slag in the flash furnace, and resulting in further cost savings of $10/ ton smelted. A sample of 153 concentrate produced in the lab using heavy liquid separation has been sent to INCO toll & reverts for evaluation as a smelter feed material. Should it not be attractive to INCO to introduce this feed directly to the smelter, the opportunity for McCreedy to sell the concentrate to a toll smelter which is poor in copper feed could be investigated in order to realize the economic benefits presented in the model. 104 3.6 Economic Evaluation 3.6.1 Cost Estimate Process flowsheets, mass and grade balances were developed from the mineralogical testwork described in Chapter 2. Process equipment has been sized and budget estimates for the mechanical equipment were obtained. Cost estimates will be made for capital development and infrastructure, and civil, structural, electrical and instrumentation costs will be estimated from factors based on previous similar process plant designs. Capital costs have been estimated using factors for civil, structural, electrical, instrumentation and piping based on the direct cost of mechanical equipment. The cost of mechanical equipment is usually the biggest single contributor to capital costs on a project. These factors have been used in numerous previous pre-feasibility budget estimates and have found to be representative of the total cost of a process facility. Mechanical costs have been estimated from a number of sources: • LHD's - 2003 budget price quotation from Atlas Copco • Screens, crushers, pumps - US$ quotations, base date August 2001, escalated at 3% p.a. and converted to CDN$ at 0.7:1 • Kiruna chutes - fabricators quotation to INCO, 2003 • DMS drum - 2004 budget estimate from manufacturer • Sorting equipment - manufacturer's quote to INCO, in CDN$, base date 2000 • Conveyors - rates-based estimates in US$/m, base date August 2001, escalated and converted to CDN$ • Pumps - budget price quotation from manufacturer, base date 2003 Factors used in the estimation of other direct costs are shown in Table 14. Erection costs have been allowed for at double the typical surface allowance to account for the additional cost of construction underground. 105 Table 14 - Cost Estimation Factors Item Description Factor 1 Mechanical 100% 2 Civil (incl. P&G's) 17% 3 Steelwork & platework 66% 4 E&I (incl. installation) 38% 5 Mechanical erection 7% 6 Structural erection 12% 7 Piping & Valve 36.3% 8 Spares, first fill & startup costs 60% 9 1st year maintenance allowance 60% Excavations and support have been estimated from the plant layouts based on the volume of each section of plant as laid out using good engineering practice. Equipment sizes are known. Chutes and transfer points have been developed based on an angle of repose of 40°. Access for operators and maintenance has been allowed from experience. No allowance was made for additional construction access. Conveyor drifts have been allowed for at 3m x 3m. Excavations are priced at the present McCreedy capital development cost, including support to standard, which would comprise 75mm square mesh and 1.6m long roofbolts at 1.5m centres, installed on back and walls. Additional support in the form of a 50mm shotcrete spray-on lining has also been allowed for. Excavations and support are costed for all plant excavations, thus any savings through utilizing existing excavations as described in the text have not been accounted for. Excavation and support rates used in the estimate are for a standard 5 x 5 m development end (including mucking to the rockpass), and are based on actual 2004 costs: • Excavation - $157.12 / m 3 • Support- $1.95/ m 2 106 Additional Items: The design and construction costs for the facility have been allowed for at 25% of the direct cost. This includes for preliminary expenditure of up to $588 000 on feasibility-level design work which will be required prior to detailed design and construction. A project contingency of 15% has been allowed. This is not a reflection of the accuracy of the estimate, which is intended to be -10% +30%. A summary of the capital cost estimate for both mining and processing systems is shown in Table 15. Details of the capital cost estimate are presented in Appendix 7. Table 15 - Capital Cost Estimate Summary Area Description Area Cost Mining and Processing Plant PFD100 153 Mining PFD200 MOB Mining PFD300 153 Preconcentration PFD400 Concentrate Hoisting PFD500 MOB Preconcentration 600 Excavations & Support Mechanical subtotal $4,434,747.74 Subtotal Direct Costs $879,646.00 $947,694.75 $2,359,022.62 $2,554,237.52 $3,856,927.28 $1,598,570.41 $12,196,098.58 Plant Infrastructure Piping & Valves sum Substations Substation excavations 216 m3 Office/Ctrl Office/Ctrl Excavation 216 m3 Subtotal P&l Subtotal Infrastructure $1,609,813.43 $75,390.71 $33,937.92 $75,390.71 $33,937.92 $1,828,470.69 $14,024,569.26 Indirect Costs EPCM @25% $701,228.46 Spares, First Fill & Startup @60% Mech Maintenance Y1 @10% Mech Subtotal Indirect Costs $3,506,142.32 $2,660,848.64 $443,474.77 $6,610,465.73 Subtotal Direct & Indirect Contingency 15% Total $20,635,034.99 $3,095,255.25 $23,730,290.24 Accuracy -10% -$2,373,029.02 30% $7,119,087.07 $21,357,261.22 $30,849,377.31 107 Operating costs have been estimated for the process plants. Impacts on present operations in terms of productivity improvements have not been accounted for. Operating cost factors can be broken down as follows: 153 pre-concentration system • Production rate: 1800 tpd • Power: 940 kW running load @ $0.06/kWh = $0.63 It • Labour: 2 plant operators + 1 LHD operator x 3 shifts = 9 personnel @ $60,000/mea = $0.81/t • Maintenance @ 5% capital / annum = $0.40 It • Reagents: FeSi @ 0.25 kg /1 & $2.07 / kg = $0.51 It Water consumption is assumed to be included in the overall operating cost for the facility. Total operating cost/t for the 153 pre-concentration facility is estimated at $2.36 It. MOB pre-concentration system • Production rate: 3200 tpd • Power 465 kW running load @ $0.06/kWh = $0.18 /t • Labour: 2 plant operators + 1 LHD operator x 3 shifts = 9 personnel @ $60,000/m ea = $0.46 It • Maintenance @ 5% capital / annum = $0.22 It • Reagents: no reagents required, but an allowance of $0.22/t has been allowed Water consumption and diesel for LHDs and haul trucks is accounted for elsewhere in the mine cost system. Total operating cost/t for the MOB pre-concentration facility is estimated at $1.08/t. The overall additional operating cost per ROM ton for the installation of a pre-concentration system at McCreedy East is thus estimated at $1.54 / ROM t. 108 3-6.2 Cost and Revenue Impact Evaluation Impacts on the working costs at McCreedy are shown below. Costs are cash costs of minin] and do not include smelting costs, overheads, capital depreciation, taxation or redistributed costs. Cost of smelting is accounted for in the 80% smelting fee imposed on the revenue stream in the financial evaluation. Mining costs are unchanged. Backfill savings are made due to the reduced requirement for sandfill in the MOB and LMOB orebodies. Haul costs are reduced through a reduction in haul distance of 50% for the 153/170 ores, and a reduction of 22% in the haul tons for the MOB and LMOB. Haulage, hoisting, surface transport, milling and tailings disposal costs are similarly reduced for these ores. Underground crushing and conveying, and surface milling are eliminated entirely for the 153/170 ores. A reduced hoisting cost has been substituted for the hydraulic hoisting of the reduced quantity of 153/170 pre-concentrate. Cost savings are calculated using the present working costs at McCreedy, based on the percentage reduction of tons in each unit process, calculated in terms of the quantity of ROM tons mined (Figure 42). It is assumed that unit working costs do not vary over small variations in the tonnage to each process. Working costs for the pre-concentration plants have been estimated as described in the section on operating costs. McCreedy East Mining and Milling Cost Comparison 140 120 : 100 80 60 y> 40 20 0 o 2 O MOB LMOB • Ta i l i ngs D i s p o s a l • Mi l l ing • S u r f a c e T r a n s p o r t • Ho is t • C r u s h / c o n v e y • P r e c o n c • H a u l • M u c k i n g Backf i l l • M i n i n g Figure 42 - Mining and Milling Costs Comparison with and without Pre-concentration 109 Cost savings are generated against the baseline through the reduction of tons in the functional processes downstream of the pre-concentration plant. Revenue penalties are incurred in the metal losses to waste across the plant. Cost savings in the mining and surface handling of MOB and 153 ores vary between 15 and 20%. Revenue losses for all metals in the MOB ore are of the order of 4%. In the case of the production of a smeltable concentrate from the 153 and 170 orebodies, additional revenue is generated from these ores by the reduction in metal losses through bypassing of the surface mill. Additional revenue from increased PGM recoveries is of the order of $ 10/t. Based on an expected constant tailings grade at the surface mill, there may be a benefit in increased metal recoveries from the MOB due to the increase in grade of ore delivered from McCreedy, however, this has not been allowed for in the analysis. A detailed working cost savings, revenue impacts, discounted cashflow analysis and NPV calculation based on the mass and grade balance from the testwork is contained in Appendix 8. Mining tonnages and grades used as a basis in the study are shown in Table 16. Table 16 - Mining rates, grades, costs and values MOB LMOB 153 170 Total / average Mining rate tpd 1800 1500 1100 600 5000 Cu % 0.90 0.92 11.42 7.01 3.03 Ni% 1.85 1.94 1.22 0.92 1.7 TPM g/t 0.64 0.96 10.56 14.4 2.88 Mining cash cost S/t 45 54 84 124 88 Value $/t (after milling and smelting) 100.41 107.69 298.10 264.61 133 6.3 Financial Evaluation The preliminary economic evaluation has been undertaken using the existing McCreedy East Life of Mine Plan for the selected orebodies as a basis. Tonnages, grades, and metal values as used in the plan have been used as a basis for estimates of savings and revenue impacts of the various proposals. Data has been taken from the McCreedy cost database in order to generate direct activity-based costs for the mining process in each orebody, kindly supplied by INCO Mines' Technical Services. Costs exclude amortization and 110 depreciation, as well as Distribution as it is felt that this would overstate the projected cost savings in the model. Milling costs were obtained from previous INCO technical reports and confirmed by personnel at the mill. The financial parameters that have been used are shown in Table 17. Table 17 - Financial Evaluation Parameters Discount rate 11% Escalation 2% Baseline 2004 Tonnage & Grade Profile Current McCreedy Life of Mine Plan Long Term Metal prices Nickel $3.00 / lb Copper $0.80 / lb Platinum $650 / oz Palladium $300 / oz Gold $350/oz The working cost saving has been applied to the tonnage profile in the life-of-mine plan. Revenue losses are experienced on the MOB and LMOB ores due to the decreased metal recoveries. However, the saving per ton in pre-concentrating the ore exceeds the value of the revenue lost. Cost savings as described in Figure 41 are also enjoyed in the case of the 153 and 170 ores. Revenues for these ores are increased as the pre-concentration recoveries are higher than the present mill recoveries. A possible further saving of $10/ ton smelted through bypassing the flash smelter is not accounted for in this analysis. Using the tonnage contained in the mining plan, the cost savings and revenue impacts described above, and based on an initial cost of CDN $30 849 377 for the integrated underground mining and processing facility, as set out in the cost estimate, the project demonstrates an NPV of $ 134m and Internal Rate of Return of 79% at a discount rate of 11%. Payback period based on these figures is under 2 years. A secondary impact of lowering the working costs to surface for the mining of the 153 and MOB ores is the creation of an opportunity to lower the economic cutoff grade for the mining of these orebodies. The introduction of underground pre-concentration has been shown to 111 lower the costs and increase the grade of ore delivered to surface. Based on the samples taken and the testwork presented, the increase in grade of 153 ore delivered to surface is of the order of 69%; the increase in grade of the MOB ore delivered to surface is of the order of 20%. As discussed in Chapter 2, the samples taken were of a higher grade than present ROM ore, thus projected upgrade ratios will be even higher. The evaluation shows there is a reduction in the cost of delivering this upgraded ore to surface; cost savings in the case of the 153 and MOB are 18% and 20% respectively. The facilitation of bulk mining techniques in the 153 may further increase these savings. There is an opportunity through the lowering of costs to lower the cutoff grade and increase the mining rate in each case, and yet still deliver the same grade of ore to surface as in the present mining scenario, or alternately deliver a higher grade product to surface at the present mining rate. The economic benefits of this will be diluted by the increased capital requirements to do this; however, there is an opportunity to re-evaluate the orebody model based on the lower cutoff grade and assess the increase in ore reserve under these conditions. It may be beneficial to adopt pre-concentration solely in order to increase the extraction of the present mineral resource. It has been suggested that some of the impacts of implementing underground pre-concentration are diluted in the case of INCO's Sudbury Operations. The mining and smelting operation overall is mature, and the opportunity is for a brownfields (i.e. existing) operation. It has been previously identified that adopting pre-concentration of the ore can either increase the effective capacity of downstream plant, or alternately reduce the size of downstream facility required. There is a significant investment in the central processing facility at Copper Cliff, which is already operating under capacity, with a bottleneck at the smelter and refinery, not the mill. There is no opportunity for increasing the capacity of the facility, or alternately reducing the capital investment in this now. However, it is felt worthwhile to explore these capital savings in the context of a greenfields operation in order to quantify the type of benefits to be enjoyed in this area. In project evaluation, the capital cost of a prospective operation may be extrapolated from the actual cost of an existing similar facility of a different capacity. A capital scaling formula is used to estimate the change in capital requirements based on a comparison of the prospective tonnage to the tonnage of the existing operation according to the relation: C 2 = C(t2/t,)2 / 3 - (1) 112 Where C2 = estimated capital Ci = known capital t2 = throughput of new facility ti= throughput of existing facility Based on this formula, the changing capital requirements for an operation employing underground pre-concentration based on a baseline capital amount of 100% for the facility can be calculated. The following graph indicates the change in overall capital requirements for a given operation for a range of values for % waste rejected (Figure 43). In the evaluation, capital allocations for the operation has been split into the categories of mining, haulage, hoisting, surface transport, and milling capital, each with a weighting of 100 units. The mining rate is constant, and therefore mining capital requirements remain at 100 units in all cases. A pre-concentration capital allocation of 100 units is made in all cases as the cost of the pre-concentration facility is related directly to the mining rate. A l l other capital requirements vary with the proportion of waste rejected according to Equation 1. Overall capital impacts are evaluated against a total of 5 x 100 = 500 units of original capital. Greenfields Capital Cost Impacts Figure 43 - Underground Pre-concentration Capital Cost Impacts Capital and revenue impacts can be compared directly by substituting real values for the baseline capital amounts and metal prices. As can be seen in the graph, at low degrees of waste rejection the additional cost of the pre-concentrator outweighs the savings in other 113 capital areas. At between 30 & 40% waste rejection by mass, the downstream capital savings more than offset the increased underground Capex. Taken in conjunction with the operating cost savings indicated previously, the opportunity to reduce the initial capital investment for an operation is attractive. 3.7 Conclusion The capital and operating costs for the underground mining and processing system design have been presented. The estimated cost for the system, including additional mining equipment and infrastructure, excavations, design and construction is of the order of CDN$30.8 million. Operating costs for the plants have been estimated at $2.36/ ROM t and $1.08/ ROM t for the 153 and MOB pre-concentrators respectively. However, operating cost savings of 20% overall are enjoyed due to the reduction in throughput to the system downstream of the pre-concentration plant. Due to the brownfields nature of the case study mine, no capital savings can be enjoyed as there is substantial existing investment in surface processing and waste disposal infrastructure. However, the projected operating cost savings, combined with the additional revenues from the 153 ore results in an NPV of CDN$134 million over and above the present life-of-mine NPV for McCreedy East, which must be considered sufficient justification for further consideration of the concept. 114 Chapter 4 - Conclusions and Recommendations The motivation, literature review and basic design criteria for underground pre-concentration has been presented in the thesis. The literature indicates that there is a substantial economic and environmental motivation for underground pre-concentration. The literature also indicates that enabling technologies for underground pre-concentration are well established and that the research focus should be on the integration of these existing technologies into an appropriately designed underground mining and processing system. Basic design criteria for underground mining and processing systems have been developed for utilization in the engineering of such a system. A case study for massive sulphide ores at INCO's McCreedy East Mine has been undertaken, comprising sampling, mineralogical testwork, process development as well as system design and integration. The testwork indicates that a substantial proportion of the ROM ore can be rejected with good metallurgical recoveries at a coarse particle size. Operating cost savings are derived from the early rejection of waste in the mining system. There are also positive revenue benefits arising from the opportunity to produce a smeltable pre-concentrate from the high grade 153 type ore. The size and cost of excavations for the underground plant have been evaluated, and are considered in line with existing excavation requirements for a mine of this nature. It is felt that there is a very strong business case for the consideration of pre-concentration of the ore underground at McCreedy East. It is especially attractive in the light of the possibility of producing a smeltable concentrate suitable for smelter feed. Using INCO's present financial evaluation criteria, the system shows good feasibility as well as presenting several positive environmental benefits in terms of a reduction in the quantity of solid waste delivered to surface during mining. However, there is a substantial amount of further research required to establish the feasibility of the system and evaluate the impacts of implementation to an acceptable level of accuracy. This future work includes: o Bench scale testing of the AE 80 conductivity / magnetic sorting arrangement on MOB and 153 ore at Mogenson in Aachen, Germany o Investigation of the comminution and liberation characteristics of the MOB ore o Testwork to specify the correct re-introduction point of the 153 concentrate into the beneficiation process at Copper Cliff o Pilot testing of dense-media separation on surface at McCreedy East 115 o Re-development of the orebody models at McCreedy East based on a lower cutoff grade in order to develop a revised estimate of the ore reserve o Detailed process design and costing to pre-feasibility level on the combined MOB and 153 underground pre-concentration system o Investigation of the rock mechanics aspects of additional non-mining excavations around the Location 7,8 and 9 areas in the mine o Investigation of the rheological properties of the 153 concentrate and detailed design of the hydraulic hoisting system o Investigation of the mechanical properties and rheology of an aggregate concrete mix produced from the waste products of the metallurgical process Coarse-particle grinding and flotation was identified in the literature review as a potential technology for underground processing, and must be tested. Pre-concentration through autogenous grinding and classification of the ore shows potential for ores where ore and wall rock competence are substantially different, and thus requires further investigation. As previously mentioned, there is substantial research and testwork required in the area of hydraulic hoisting as an enabling technology for underground pre-concentration. Further research is required outside of the immediate goals of the case study undertaken in this report. Long term research will include modelling of the geotechnical impacts of introducing high-strength backfill on the mine design for a deep or highly-stressed orebody with poor ore and wall rock quality. The environmental health and safety implications of introducing mineral processes into the underground environment require further investigation. Particular occupational health and safety concerns are air quality around and downstream of the process facility arising from heat noise and dust generated during processing. The use of conventional process reagents such as xanthates and cyanide in the underground environment also requires further consideration. The impact of introducing process reagents on groundwater quality and the general quality of mine water must be assessed. A final area of research will be to apply the methodology developed in this thesis to additional case studies to expand the range of applicability of underground pre-concentration. 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Coal Board Proceedings, Paper 6, pp 59-64 Swan, G., Arjang, B., Hedley, D.G., 1993, On the use of Rockfdls in Overhand Cut-and-Fill Mining, Innovative Mine Design for the 21st Century, Bawden and Archibald (ed) pp 103-110 Tamlyn, N. , 1994, A Review of Deep Gold Mining Methods at Driefontein Consolidated Limited, Proceedings XVth SAIMM CMMI Congress, Johannesburg, Vol 1, pp29-38 Thomas, C , 2004, McCreedy East Working Cost Report Anon, 1991, 'Underground Wet Crushing and Pumping of Fine Ore to the Clarabelle Mill ' , INCO General Engineering Department Estimate No. 2409-OA, 52 pages Warhurst A., Bridge G., 1996, Improving Environmental performance Through Innovation: Recent Trends in the Mining Industry, Minerals Engineering, Vol 9, No.9, pp 907-921 Way, H.W., 2004, A Vertical Grinding Solution, NETZSCH Corporation Publication www/ Yester, M.A., 1997, The Basics and Cost Savings of Utilizing Vertical Conveyor Technology, Bulk Solids Handling, Vol 17, No.3, pp 357-369 122 Appendix 1 - 153 Metallurgical Testwork Results 123 Sample McCreedy Eas t Mine , 153 Orebody , 4 5 5 0 Leve l , X /C 1 W Date 07 /02 /2004 Gross Weight 5 0 5 . 4 k g Tare 14.2 kg %Moisture 2 .05 Net 4 8 1 . 3 3 2 6 8 Split 1 Mass gross 135.8 kg Tare 2 kg Moisture 2 . 0 5 5 6 0 4 2 9 % Net 131.105 Screen pan Tare 0 .72 kg •4^  Accep t Reject S ize ( m m ) k g Net m a s s W t % C u m W t % Tota l S inks(acc) W t % W t % C u m W t % Floats(rej) W t % W t % C u m W t % + 1 2 5 5.81 5.81 4 .43 95 .57 5.90 0.00 0.00 100.00 5.90 4 .43 7.93 92 .07 + 7 5 6.98 6 .26 4 .77 90 .80 6.26 1.12 0.85 1.93 98 .07 5.20 3 .97 7.10 8 4 . 9 7 + 5 3 9.41 8 .69 6 .63 84 .17 8.80 1.80 1.36 3.07 95 .00 7.00 5.28 9.45 75 .52 +38 8.48 7.76 5.92 78.25 7.74 0.74 0.56 1.28 93.72 7.00 5 .35 9.59 65 .93 +26 12.82 10.66 8.13 70 .12 10.72 1.42 1.07 2.43 91.29 9.30 7.06 12.64 53 .30 + 1 9 19.90 18 .46 14.08 56 .04 10.56 1.46 1.94 4 .39 86 .90 9.10 12.14 21 .73 31 .57 +9 12.87 12.15 9 .27 4 6 . 7 7 1.65 0.48 2.71 6 .14 80 .76 1.17 6 .56 11.74 19.83 +6 .7 8.26 7.54 5.75 41 .02 1.64 0.53 1.85 4 .18 76.58 1.11 3 .90 6 .99 12.84 +3 .3 10.53 9.81 7.48 33 .54 1.37 0.56 3.06 6.92 69 .66 0.81 4 .42 7.92 4 .92 +1 .6 11.79 7.98 6 .09 27 .45 1.14 0.63 3.34 7.55 62.11 0.52 2 .75 4 .92 0.00 -1 .6 34.25 35 .99 27 .45 32 .09 8.81 27 .45 62 .11 0.00 0.00 0 .00 Tota l 131.11 Tota l 44.20 55 .85 Sample 153-4550 Split 1 Assay Results - Sink/Float test @ S G 3 |Size Fraction Sinks Floats mm Wt% Cu% Ni% Au g/t Ptg/t Pd g/t TPM Wt% Cu% Ni% Au g/t Pt g/t Pd g/t TPM -1.6 27.45 27.46 0.27 1.50 8.25 19.03 28.78 +1.6 3.34 35.55 0.52 6.98 6.40 13.71 27.09 2.75 0.38 0.03 0.18 0.42 0.93 1.53 +3.3 3.06 34.14 0.62 1.06 9.85 18.34 29.25 4.42 0.39 0.04 0.16 0.38 0.64 1.18 +6.7 1.85 30.19 0.79 2.86 10.34 35.15 48.35 3.90 0.37 0.05 0.14 0.33 0.52 0.99 +9 2.71 31.64 1.30 2.06 10.00 29.56 41.62 6.56 0.51 0.06 0.09 0.21 0.29 0.59 +19 1.94 29.07 0.70 1.56 11.23 23.73 36.52 12.14 0.91 0.11 0.18 0.41 0.64 1.23 +26 1.07 27.71 1.29 13.38 8.55 25.01 46.94 7.06 0.76 0.07 0.19 0.43 0.55 1.17 +38 0.56 19.77 0.96 1.76 7.39 12.40 21.55 5.35 0.23 0.08 0.75 1.74 3.67 6.16 +53 (calc) 1.36 29.44 0.81 3.90 9.00 22.12 35.01 5.28 0.51 0.06 0.24 0.56 1.03 1.84 +75 (calc) 0.77 29.44 0.81 3.90 9.00 22.12 35.01 3.56 0.51 0.06 0.24 0.56 1.03 1.84 125 (calc) 4.43 0.51 0.06 0.24 0.56 1.03 1.84 Average/Total 44.11 29.44 0.81 3.90 9.00 22.12 35.01 55.44 0.51 0.06 0.24 0.56 1.03 Recovery 97.88 91.08 92.75 92.75 94.45 Appendix 2 - MOB Metallurgical Testwork Results 126 Sample McCreedy Eas t Mine, Ma in Orebody, 3575 Level, Panel 3 Date 07 /02 /2004 Gross Weight 566 .2 k g Tare 16.6 kg %Moisture 3 . 7 0 % % Net 5 2 9 . 2 6 4 8 Split 3 Mass gross 114 .2 k g Tare 2 kg Moisture 3 .6135455 % Net 108.287 Screen pan Tare 0.72 k g Size Dry m a s s net %wt C u m W t % Accept W t % Reject W t % m m kg Total Mass Wt% Wt% C u m Wt% Mass Wt% W t % C u m W t % + 1 2 5 4 .87 4 .87 4 .50 95 .50 2.30 2.30 4 .50 5.73 94.27 0.00 0 .00 0.00 100 .00 + 7 5 7.23 6.51 6.01 89 .49 6.53 2.63 2.42 3.09 91.18 3.90 3.59 16.65 83 .35 + 53 9.87 9.15 8.45 81 .04 9.29 6.00 5.46 6.96 84.22 3.29 2 .99 13.86 6 9 . 4 9 + 38 12.01 11.29 10.43 70.61 10.82 7.12 6.86 8.75 75.47 3.70 3.57 16.54 52 .95 + 2 6 12.41 10.25 9.47 61 .14 11.90 8.80 7.00 8.93 66.54 3.10 2.47 11.44 41 .52 + 1 9 10.59 9.15 8.45 52 .69 2.33 1.82 6.62 8.44 58.10 0.50 1.83 8.48 33 .04 + 9 19.28 18.56 17.14 35 .55 1.89 1.54 14.02 17.88 40.22 0.34 3.11 14.44 18.60 + 6.7 8.39 7.67 7.09 28.47 1.96 1.57 5.68 7.24 32.98 0.39 1.41 6.53 12.08 +3 .3 11.44 10.72 9.90 18.57 1.36 1.15 8.41 10.72 22.26 0.20 1.49 6.92 5.15 +1 .6 8.08 7.36 6 .80 11.77 1.74 1.45 5.69 7.25 15.00 0.28 1.11 5.15 0.00 -1 .60 14.90 12.74 11.77 12.74 12.74 11.77 15.00 0.00 Tota l 108.287 78 .43 21 .57 Sample: MOB-3575 Split 3 Assay Results - Sink Float test @ S G 3 Size Class Sinks Floats Wt% Cu% Ni% Aug/t Ptg/t Pd g/t TPM Wt% Cu% Ni% Au g/t Ptg/t Pd g/t TPM -1.6mm 11.77 0.81 3.45 0.05 0.49 0.58 1.12 0.00 0.00 0.00 0.00 0.00 0.00 0.00 +1.6mm 5.69 1.48 3.63 0.15 0.61 0.46 1.22 1.11 1.13 0.43 0.14 0.18 0.35 0.67 +3mm 8.41 1.15 3.21 0.04 0.47 0.47 0.98 1.49 1.02 0.26 0.04 0.15 0.15 0.34 +6.7mm 5.68 1.42 3.40 0.13 0.42 0.51 1.06 1.41 1.13 0.37 0.03 0.12 0.19 0.34 +9mm 14.02 1.46 3.21 0.14 0.43 0.57 1.14 3.11 0.91 0.27 0.03 0.14 0.22 0.39 +19mm 6.62 1.51 3.21 0.10 0.44 0.56 1.10 1.83 0.82 0.26 0.04 0.16 0.26 0.46 +26mm 7.00 1.41 3.14 0.09 0.51 0.50 1.10 2.47 1.13 0.35 0.05 0.19 0.46 0.70 +38mm 6.86 0.97 3.75 0.05 0.49 0.62 1.16 3.57 1.33 0.55 0.07 0.29 0.40 0.76 +53 (calc) 5.46 0.65 2.91 0.11 0.46 0.54 1.11 2.99 0.83 0.27 0.04 0.13 0.23 0.40 +75 (calc) 2.42 0.32 2.91 0.11 0.36 0.54 1.01 3.59 0.83 0.27 0.04 0.13 0.23 0.40 +125 4.50 3.43 0.50 0.25 0.30 0.59 1.14 Average/Total 78.43 1.33 3.03 0.11 0.45 0.54 1.10 21.57 0.91 0.30 0.05 0.15 0.25 Recovery 84.10 97.32 89.29 91.70 88.79 90.24 *3.43 value rejected A p p e n d i x 3 - P rev ious I N C O Tes twork D a t a 1 2 9 To: Dr. Bern Klein cc: Simon Nickson From: Andrew Bamber RE: Previous INCO ore sortine testwork Date: 22 April 2004 Bern, I have briefly summarised data from various test reports made available to us during the fieldwork in February. Combined with the results from our testwork here at UBC, there is definitely a basis for looking at pre-concentration for the Ontario Division as a whole. 1. Test result summaries of various sorting tests undertaken by INCO and UBC Test Ore Method Feed Grade Reject Concentrate C u Ni Wt% C u % Recovery N i % Recovery 1530pt 153 Optical 5.67 0.4 78.7 24.76 92.9 0.47 75.4 1530ptcon 153 Optical/ Conductivity 5.67 0.4 54.6 11.83 94.7 0.73 81.2 V.384 McCreedy Optical/ Magnetic 4.33 2.4 12.3 5.93 99.8 3.5 99.6 V.444 McCreedy Optical/ Magnetic 1.81 2.4 41.6 4.27 99.3 2.78 99 INCO 'Sudbury' Conductivity/ 90% 1.09 1.06 38.5 1.55 87 1.66 96.1 INCO •Sudbury1 Conductivity/ 100% 1.19 1.21 23.5 1.5 96.7 1.56 99.1 INCO Bircrrtree n/a n/a 0.96 80.5 n/a n/a 4.79 97.5 Exportech 6283 Stobie HG Magnetic n/a n/a 23.8 2.74 94.87 1.874 98.9 Exportech 6287 Stobie LG Magnetic n/a n/a 13.03 0.365 92.69 0.873 97.5 Exportech «27« Stobie LG Magnetic n/a n/a 12.42 0.604 89.06 0.376 98.95 153 Heavy Liquid 13.26 0.39 55 29 98 0.8 91 MOB Heavy Liquid 1.11 2.54 22 1.32 3.02 3.02 97.3 130 2. Magnetic and Conductivity amenability testwork McCreedy Ore Magnetic and Conductivity Testing 3. Discussion Optical and conductivity sorting tests 153 Opt and 153 OptCon were performed on McCreedy footwall ore. Copper recoveries were between 92-95%. Nickel recoveries for optical sorting alone were unacceptable, but with a combination of optical and conductivity rose to >80%. Tests V384 and V444 were performed using a combination of Optical sorting and Rare Earth Magnetic Separation. Copper recoveries were >98% in both tests, nickel recoveries were also >98%. Precious metal recoveries are comparable to copper in all cases. Tests Sudbury 90 and Sudbury 100 were undertaken using conductivity on a typical low grade ore, probably from Stobie or Creighton, assaying at 1% Cu + 1% Ni and 1.2% Cu + 1.2% Ni respectively. Sudbury 90 was undertaken at 90% conductivity sensitivity, Sudbury 100 at full sensitivity, with an improvement in the sorting results: in each case nickel recoveries are >95%, however, copper recovery was low in the Sudbury 90 test, but improved to 96% for the Sudbury 100 test. Waste rejection varies between 40-80% for the McCreedy ores, and 25-40% for the low-grade ores. Electromagnetic sorting testwork was undertaken on Stobie high- and low-grade ores in 2000 by Exportech. Nickel recoveries were generally >95%, however copper recovery varied between 87 & 89% for the low grade ore to 94 & 95% for the high grade ore. Waste rejection varied between 5 and 25%. The results of magnetic and conductivity amenability testwork on McCreedy East ore is shown in FigfJ. Medium to high grade ore shows a good magnetic response of above 10%, and conductivity responses of > 40%. It is suggested that a combination of magnetic and conductivity sorting, with similar detection levels as shown would demonstrate reliable metallurgical results in sorting. 131 4. Other tests Sorting testwork has been undertaken on ore from the Birchtree Mine, and on the low-grade rockpile at Creighton. Results from the Birchtree tests are not clear. While results from the sorting of the Creighton rock pile were poor, due to the extremely low grade of the pile, a concentrated product was produced at a grade acceptable for introduction as feed to the Clarabelle Mill. Other amenability testwork was conducted on McCreedy and Stobie ore in 2001 using a Shimadzu XRF fluorescent tester. Fluorescent responses for both ore and waste in each case were inconclusive. 5. Conclusion Previous and present separation testwork on INCO ores have shown that between 12 - 80% of the ore mined (depending on the orebody) could be rejected at an acceptable metallurgical recovery. Appropriate sorting technology appears to be magnetic/conductivity based methods for contact ores and optical methods for the footwall ores. Dense media separation will work on both types of ores. The proportion of rejects from low-grade contact type ores such as Stobie low grade, Creighton and low grade MOB is low, however metallurgical recoveries are good using the appropriate separation technology. The proportion of rejects for the high-grade narrow-vein footwall ores such as the 153 and 170 could be as high as 80%, and metallurgical results are good in all cases due to the characteristics discussed in the mineralogical report. Appropriate process optimization based on a detailed mineralogical analysis of each ore will improve the results presented in the table. This would result in a substantial cost saving in material handling and milling costs for the Ontario division, as well as an increase in the metallurgical performance at the Clarabelle mill due to a substantial increase in the feed grade to the mill. Metal losses of approximately 5% will be more than offset by operating cost savings as well as improvements in recovery at the mill. The average mass rejected in the tests is 36.47%. Even taking into account the high proportion of low grade ore mined by INCO, there is a good possibility that the total tonnage of ore delivered to Clarabelle could be reduced by 30%, with a proportionate increase in feed grade. This will enable the shutdown of the rod mill/ball mill circuit as discussed with Andy Ken-in Feburary. The cost of mining sandfill on surface will also be offset at each operation due to the generation of waste which could be used as fill close to the mining face. I suggest we table these results to INCO for discussion during our visit. Regards Andrew 132 Magnetic/ Conductivity Test Results - Strathcona/McCreedy 153 Size Grade Conductivity Mag 1 4.75 Io 0 0 2 4.75 Io 0 0 3 5 Io 0 0 4 4.7 Io 0 0 5 6.7 mid 25.6 1.6 6 1.25 mid 46.4 4.5 7 1.25 mid 41.6 4.8 8 2.2 mid 12.4 6.1 9 6 hi 80.4 94 10 4.75 hi 76 11.8 11 5 hi 73.5 12.2 12 5.5 hi 71.2 12.5 13 5.5 hi 85.4 22 14 4.7 hi 77.6 28 15 5 hi 77.6 40.7 HI 85.4 40.7 LO 12.4 1.6 RANGE 73 39.1 MEDIAN 46.4 6.1 AVE 44.51333 10-24 McCreedy Ore Magnetic and Conductivity Response Testing Sample 133 Date Sample Descriptior Response. Reponse Wavelength 3.1.01 A O I coarse ore 8 435 M C E L W A 0 2 fine ore 15 435 365nm AR1 coarse was 35 440 A R 2 ffine waste 90 420 3.1.02 G 0 6 coarse ore 53 440 S T O L W BOI coarse ore 11 430 S T O LW1 G 0 3 coarse ore 32 410 365nm G R 3 coarse was 63 435 BR1 coarse was 41 440 GR1 coarse was 16 430 G 0 5 fine ore 22 420 B 0 2 fine ore 16 410 G 0 4 fine ore 17 410 G R 4 fine waste 40 430 BR4 fine waste 37 425 3.1.02 A O I coarse ore 180 370 M C E S W A 0 2 fine ore 300 380 254nm AR1 coarse wai 195 370 AR2 coarse was 850 370 3.1.02 G 0 5 fine ore 450 370 S T O S W B 0 2 fine ore 330 370 254nm G 0 4 fine ore 390 370 G R 4 fine waste 465 370 BR4 fine waste 440 370 G R 2 fine waste 360 370 134 Testwork Stobie Mine Tester Eledrimag Date Oct-00 Method Magnetic & Eddy Current Splitter Sample 1 6262 low grade Accept Reject Recovery spBt wt% Cu% Ni% wt% Cu% Ni% Cu Ni 1 71.46 0.335 0.858 28.54 0.146 0.249 85.1745 89.6133 2 80.41 0.321 0.835 19.59 0.117 0.068 91.8444 98.0546 3 86.45 0.311 0.788 13.55 0.092 0.023 95.5688 99.5446 4 95.07 0.292 0.719 4.93 0.071 0.021 98.7548 99.8488 5 100 0.281 0.684 0 0 0 100 100 Sample 3 6287 low Accept Reject Recovery spBt wt% Cu% Ni% wt% Cu% Ni% Cu Ni 1 68.93 0.403 1.043 31.07 0.208 0.192 81.1264 92.3382 2 72.62 0.399 1.015 27.38 0.194 0.152 84.5081 94.6556 3 77.32 0.389 0.962 22.68 0.184 0.153 87.8159 95.5428 4 86.97 0.365 0.873 13.03 0.192 0.151 92.6947 97.474 5 100 0.342 0.779 0 0 0 100 100 Sample 3 6279 high grade Accept Reject Recovery spRt wt% Cu% Ni% wt% Cu% Ni% Cu Ni 1 63.79 2.974 0.467 36.21 1.137 0.44 82.1681 65.154 2 68.78 2.932 0.456 31.22 0.934 0.245 87.3671 80.3938 3 72.12 2.893 0.429 27.88 0.797 0.178 90.3751 86.1773 4 83.95 2.655 0.386 16.05 0.495 0.082 96.5582 96.0971 5 100 2.309 0.328 0 0 0 100 100 Sample 4 6283 high grade Accept Reject Recovery split wt% Cu% Ni% wt% Cu% Ni% Cu Ni 2 67.14 2.775 2.051 32.86 1.024 0.204 84.7025 95.358 3 69.98 2.762 1.99 30.02 0.89 0.17 87.8557 96.4649 4 76.2 2.739 1.874 23.8 0.474 0.066 94.872 98.912 5 100 2.2 1.444 0 0 0 100 100 Sample 5 6276 low grade Accept Reject Recovery split wt% Cu% Ni% wt% Cu% Ni% Cu Ni 1 71.66 0.636 0.447 28.34 0.486 0.045 76.7928 96.1711 2 74.3 0.639 0.436 25.7 0.461 0.034 80.0293 97.3735 3 78.06 0.622 0.417 21.94 0.49 0.034 81.872 97.7597 4 87.58 0.604 0.376 12.42 0.523 0.028 89.0634 98.955 5 100 0.594 0.333 0 0 0 100 100 135 Sample 6 6290 low grade Recovery Accept Reject Ni split wt% Cu% Ni% wt% Cu% Ni% Cu 1 64.16 0.714 0.467 35.84 0.479 0.08 72.7405 91.2665 2 66.86 0.72 0.456 33.14 0.447 0.07 76.4687 92.9292 3 71.5 0.701 0.429 28.5 0.45 0.074 79.6256 93.5667 4 81.37 0.679 0.386 18.63 0.415 0.075 87.7243 95.7409 5 100 0.63 0.328 0 0 0 100 100 Stobie Ore Sorting Accept Reject Test Wt% Cu% CuWt% Ni% NiWt% Cu% CuWt% Ni% NiWt% 6279H/G 83.95 2.655 96.55 0.386 96.09 0.495 3.45 0.082 3.91 6283H/G 76.2 2.739 94.87 1.874 98.91 0.474 5.13 0.066 1.09 6262L/G 95.07 0.292 98.75 0.719 99.84 0.071 1.25 0.021 0.16 6287L/G 86.97 0.365 92.69 0.873 97.47 0.192 7.31 0.151 2.53 6276L/G 87.58 0.604 89.06 0.376 98.95 0.523 10.94 0.028 1.05 6290L/G 81.37 0.679 87.72 0.386 95.74 0.415 12.28 0.075 4.26 INCO - Stobie Electromagnetic Sorting Results 100 : r r — — — — _ J | f c — : — - — ~ r - r - T 3 96 92 88 84 80 & 1? ft Test / Ore 2.5 2 1.5 1 0.5 CuWt% NiWt% # R E F I Cu% Ni% Cu% Ni% 136 McCreedy Size/Grade Distribution After Paventi/Buksa 2001 Cu Ni TPM Size +6 -6 +4 -4 +2 -2 +1 -1 +3/4 .3/4 + 5/8 -5/8+1/2 -1/2 +3/8 -3/8 +4# -4#+8# -8# 30 25 + 11.25 1.19 0.001043 %Wt Metal Cum Grd 20 + 5.6 100 3.4 94.4 20.3 91 16.3 70.7 5.6 54.4 3.3 48.8 2.6 45.5 7.2 42.9 12.3 35.7 8.5 23.4 14.9 14.9 100 0.4 1.1 12.8 7.1 10.8 1.7 3.8 6.4 13 15.8 27.1 100 0.4 1.5 14.3 214 32.2 33.9 37.7 44.1 57.1 72.9 100 Grade Cu grade Ni 0.803571 3.639706 7.093596 4.900307 21.69643 5.795455 16.44231 10 11.89024 20.91176 20.46141 11.23953 M c C r e e d y Size / Grade Distribution^ 1 0 0 80 j#> & f & ' Size Class 137 Copper Grade by Size -8# -4#+ -3/B -1/2 -578 -3/4+ -1 -2+1 -4+2 -6+4 +6 8# +4# +3/8 +1/2 578 +3/4 Size Range 30 25 20 | at 15 j S Grade Cu E 10 O 5 0 I 138 Sortng Results - Mogenson Test 384 McCOpt Cu% Wt% Ni% Wt% Total 100 5.67525 100 0.40704 100 n/a Reject 78.7 0.51 7.1 0.39 75.4 n/a Accept 21.3 24.76 92.9 0.47 24.6 n/a McCOptCon Cu% Wt% Ni% Wr% Total 100 5.67112 100 0.40786 100 n/a Reject 54.6 0.55 5.3 0.14 18.8 n/a Accept 45.4 11.83 94.7 0.73 81.2 n/a V444 Material 1 20%ore low/50 %wt Total 15.4 100 Accept 4 26 Reject 11.4 74 high/50 Total 11.3 100 Accept 4 35.4 Reject 7.3 64.6 high/75 Total 8.2 100 Accept 3.7 45.1 Reject 4.5 54.9 Mat 2 80% ore high/80 Total 9.5 100 reject 4.2 44.2 accept 5.3 55.8 Sudbury Ore high/90 mass Wt% CuGrd CuWttb NiGrd NiWr% TPMGrd Wr% Total 18.7 100 1.09 100 1.06 100 0.035 100 reject 7.2 38.5 0.37 13 0.11 3.9 0009 9.8 accept 11.5 61.5 1.55 87 1.66 96.1 0.052 90.2 Sudbury Ore high/100 Total 28.5 100 1.19 100 1.21 100 0.042 100 reject 6.7 23.5 0.17 3.3 0.05 0.9 0.001 0.6 accept 21.8 76.5 1.5 96.7 1.56 99.1 0.055 99.4 Summary Results Test Ore Method Feed Grade Reject Wt% Concentrate Cu Ni Cu% Recovery Ni% Recovery 1530pt 153 Optical 5.67 0.4 78.7 24.76 92.9 0.47 75.4 1530ptcon 153 Optical/Cor 5.67 0.4 54.6 11.83 94.7 0.73 81.2 V.384 McCreedy OptJcal/Mai 4.33 2.4 12.3 5.93 99.8 3.5 99.6 V.444 McCreedy Optical/Ma! 1.81 2.4 41.6 4.27 99.3 2.78 99 INCO 'Sudbury ConductMt 1.09 1.06 38.5 1.55 87 1.66 96.1 INCO 'Sudbun/ Conductivit 1.19 1.21 23.5 1.5 96.7 1.56 99.1 INCO Birchtree n/a n/a 0.96 80.5 n/a n/a 4.79 97.5 Exportech I Stobie HG Magnetic 23.8 2.74 94.87 1.874 98.9 Exportech (Stobie LG Magnetic 13.03 0.365 92.69 0.873 97.5 139 Exportech6 Stobie LG Magnetic 12.42 0.604 89 06 0 376 98 95 UBC1 153 Heavy Med 13.26 0.39 55 29 98 0 8 91 UBC 2 MOB Heavy Med 1.11 2.54 22 1.32 3.02 3.02 97.3 V384 Optical + RE Magnet TesM Cu % Ni % Total 24.8 4.33 100 2.4 100 AcceptO 20.7 5.14 99 2.83 98.4 AcceptM 1.05 0.79 0.8 0.67 1.2 Accept Tot 21.75 5.93 99.8 3.5 99.6 reject 3.05 0.08 0.2 0.09 0.5 V444 Optical* RE Magnet Total 30.5 1.81 100 2.4 100 AcceptO 15.8 3.36 96 2.12 95.2 AcceptM 2 0.91 3.3 0.66 3.8 Accept Tot 17.8 4.27 99.3 2.78 99 reject 12.7 0.03 0.7 0.03 1.1 Test Accept Wf AccCu% AccCuWF* AccNi% AccNiWt% ReJectWt% RejCu% RejCuWWt RejNi% RejNiWt% 153 Opt 21.3 24.76 92.9 0.47 24.6 78.7 0.51 7.1 0.39 75.4 1530ptCor 45.4 11.83 94.7 0.73 81.2 54.6 0.55 5.3 0.14 18.8 V384.1 87.70161 5.93 99.8 3.5 99.6 12.29839 0.08 0.2 0.09 0.5 V384.2 58.36066 4.27 99.3 2.78 98.9 41.63934 0.03 0.7 0.03 1.1 Sudbury 9C 61.5 1.55 87 1.66 96.1 38.5 0.37 13 0.11 3.9 Sudbury 1C 76.5 1.5 96.7 1.56 99.1 23.5 0.17 3.3 0.05 0.9 INCO Ore Sorting Results Summary 140 Appendix 4 - Drawings MikroSort Far e Fa. Reiling, Marienfelde MOGENSEN GMBH & CO KG inco ONTARIO OPERATIONS 2002 Opera •Coppw CW North •Copper CW South • CraigMon No 0 •Ganon •MeCrMdyEttt •Stobie I ClarmbalaMni COPPER CUFF COPPER REFINERY MokttCainHCwbaniM PractotM Mstal* RnkkN CoppfAnodt Pitting Pitt* Rwlduji PORT OOLBORNE REFINERY EMrenlekd from Thompton c PRODUCTS • Sulfuric Acid, Oleum • Liquid Sutfur Dioxide • Nickel Powder . Nickel Pellets • Ferrc-Nickel Pellets for sale and Thompson • Nickel Oxide Sinter 75 . Nickel Sulfate to Clydach and Thompson •Copper Cathodes ' Silver, Sjajeniurn, imDlo) Telurlu  Dioxide • Platinum Group Metals Concentrate to Acton • Gold Sand to Mint •Electrolytic Cobalt • Electron tekel to Market 2003-14 Appendix 5 - Equipment Lists Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For: Review Revisions Marked: Italic Item No. No. 1 Description I Specification Area Type Off| I Supplier Type kW Comments 100 -153 Mining Area 100 100 6 Orepass Grizzly Heavy Duty Rail 500mm aperture 18 100 110 6 Orepass Grizzly Spray Water 2" Spray bar Sprays activate automatically on tipping 100 120 6 Oversize bunker 12' x 12' Excavation, 5% Floor sloj ?e Sliped from existing orepass excavations EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For: Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 200 - MOB Mining Area 200 100 6 Orepass Grizzly Heavy Duty Rail 500mm aperture 18 200 110 6 Orepass Grizzly Spray Water 2" Spray bar Sprays activate automatically on tipping 200 120 6 Oversize bunker 12' x 12' Excavation, 5% Floor slot >e Sliped from existing orepass excavations 200 130 1 3880 Rockpass Existing 200 140 1 Rockfill LHD Atlas Copco ST8C 242 Additional Rockfill LHD required due to increase in rockfill quantity EqListRevB Proj ect R0016798 McCreedy Underground Pre-concentation Proj ect Equipment List Revision: B Issued on: 23-Dec-04 For Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 300 -153 Pre-concentration Plant 300 100 1 Kiruna Tip Ore tip at Location 8 300 110 1 Crusher Feed Bin 14001 Blasted, sliped, concrete lined. Bolt and mesh support 300 120 1 Withdrawal chute 300 130 1 Scalping Grizzly Metso CVB1330 80mm splay 22 300 140 1 Grizzly Underpan 300 250 1 153 Primary crusher 2136 single toggle jaw 75mm CSS 75 Motor power revised 300 155 1 Primary Crusher Hoist Demag 5t electric trolley hoist 6 300 160 1 Crushed Ore Conveyor 900mm x 30m 22 Collects Grizzly and Crusher Discharge 300 170 1 Feed Prep Screen Metso 4400 x 1800 double deck inclined wet screen 22 C / W duck-foot spray nozzles, 38mm & 9mm square-slot polyurethane panels EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 300 180 1 Screen Underpan 300 190 1 2" spray bar Spray water for DMS feed preparation 300 200 1 DMS feed conveyor 900mm x 30m 22 300 210 1 DMS drum feed chute 300 220 1 DMS drum Malvern 3660 x 3660mm Heavy Medium Drum, 100 tph max 18.5 FeSi meduim Separation density SG 3 300 225 1 Drum maintenance hoist Demag lOt electric trolley hoist 11 300 230 1 Sinks Discharge Chute Ceramic lined 300 240 1 Float dsicharge chute Ceramic lined EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For: Review Revisions Marked: Italic Item No. No. Description Specification kW Comments Area Type Off Supplier Type 300 250 1 Drain and Rinse Screen 1800 x 1200 split-deck inclined vibrating screen 1mm rectangular slot poly panels Combined sinks and floats drain and rinse screen c/w duck foot spray nozzles 300 260 1 2" Spray bar 300 270 1 Screen underpan 300 280 1 Combined discharge chute 300 290 1 Dilute Medium Sump 3m3 300 300 1 Dilute medium sump agitator Ughtnin 15 300 310 1 Dilute medium pump 6x4 Warm an All Metal 55 300 315 1 Pipe densifier 100 m3/hr Discharge @ SG 3.3 300 320 1 Correct medium sump 3m3 300 325 1 Correct medium sump agitator Ughtnin 15 EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For Review Revisions Marked: Italic Item No. No. Description Specification kW Comments Area lype Uff Supplier Type 300 330 1 Correct medium pump 6x4 Warman All-metal 75 300 335 1 FeSi magnetic separator Malvern 900mm x 3000mm double drum magnetic separator 7.5 Added 17/7/04 Recovers FeSi from D&R screen u/f & returns FeSi to make-up tank 300 340 1 Make-up tank 3m3 - L 300 350 1 Make-up tank agitator Ughtnin 15 300 360 1 Make-up pump 4x3 Warman R/L 30 300 370 1 Discard conveyor 900mm x 30m 22 300 380 1 Concentrate conveyor 900mm x 60m 45 EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For Review Revisions Marked: Italic Item No. No. Description Specification kW Comments Area Type Off Supplier Type 300 390 1 Concentrate bin 5m3 300 400 1 Crusher feeder 600 x 1200mm vibrating pan feedi 2.2 300 405 1 Concentrate crusher Metso 1352 Shortltead Hydraflush Cone crustier 10mm CSS 90 Cruliser selection revised 300 410 1 Crusher maintenance hoist Demag 5t electric crane hoist 6 300 415 1 2" spray bar Flushing water nozzle for Hydracone crusher Density controlled 300 420 1 Pre-concentrator sump pump 50NB 1500mm vertical spindle pump 22 Spillage to shaft bottom cone tank EqListRevB Pro jec t R0016798 M c C r e e d y U n d e r g r o u n d P r e - c o n c e n t a t i o n Pro jec t E q u i p m e n t L i s t Revision: B Issued on: 23-Dec-04 For: Review Revisions Marked: Italic I t e m N o . N o . A r e a T y p e O f f D e s c r i p t i o n S p e c i f i c a t i o n k W C o m m e n t s S u p p l i e r T y p e 400 Concentrate Hoisting Section 400 100 1 Shaft bottom concentrate tank 9m3 conical bottom rubber l ined 400 110 1 Shaft bot tom cone tank repulper I i gh tn in 55 400 brinel l , h igh speed repulp ing agitator 400 120 125 2 Shaft bottom concentrate p u m p Shaft bottom concentrate p u m p standby Gardner Denve D2000 3-cylinder positive displacement m u d p u m p 900 45 m 3 / m i n @ 5 0 M P a max 5" cylinders, 100 r p m 400 130 1 Shaft bottom process water turbine Warman 8/6 A l l metal pump , double s ided shaft 300 50 m 3 / m i n @ 10 M P a 400 140 1 Shaft bottom d u m p valve Globe 250 N B automatic pressure relief valve Pressure relief at 1 6 M P a 400 150 1 Shaft bottom overf low line M / S p la in ended pipe Dumps to shaft bottom sump 400 160 1 M i d shaft concentrate tank 9m3 conical bot tom rubber l ined 400 170 1 M i d shaft cone tank repulper U g h t n i n 55 400 brinel l , h igh speed repulp ing agitator 400 180 185 2 M i d shaft concentrate p u m p M i d shaft concentrate p u m p standby Gardner Denve D2000 3-cylinder positive displacement m u d p u m p 900 45 m 3 / m i n @ 5 0 M P a max 5" cylinder, 100 r p m EqLis tRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For. Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 400 190 1 Mid shaft process water turbine Warman 8/6 All metal pump, double sided shaft 300 50 m3/min @ 10 MPa 400 200 1 Mid shaft dump valve Globe 250 NB Automatic pressure relief valve Pressure relief at 16MPa 400 210 1 Mid shaft overflow line M/S plain ended pipe Dumps to shaft bottom tank 400 220 1 Surface concentrate tank 9m3 conical bottom rubber lined 400 230 1 Surface cone tank repulper Lightnin 55 400 brinell, high speed repulping agitator 400 240 245 2 Surface concentrate pump Surface concentrate pump standby Warman 4/3 D frame centrifugal pump 55 45m3/min @ 700 kPa 400 250 1 3rd stage dump valve Globe Automatic pressure relief valve Pressure relief at 16MPa 400 260 1 Surface overflow line M/S plain ended pipe Dumps to surface concentrate sump EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For. Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 400 270 1 Surface concentrate thickener GL&V 16m 0 conventional thickener 11 400 280 1 Concentrate thickener underflow pump Watson Marlow 2" positive displacement hosepump 9 Pumps concentrate to road tanker 400 290 1 Concentrate tanker International 30t slurry road tanker Drop-bottom tanker 400 300 1 Concentrate dryer Howden Energi Systems 200 tpd fluidized bed dryer Optional: Dries cones from 12% to 0% H20 400 310 1 Shaft bottom dam 1500 m3 400 320 1 Concentrate area sump pump 50NB x 1500mm vertical spindle pump 22 EqListRevB Proj ect R0016798 McCreedy Underground Pre-concentation Proj ect Equipment List Revision: B Issued on: 23-Dec-04 For: Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 500 MOB Pre-concentration Plant 500 100 1 Kiruna Tip Ore tip at Location 2 500 110 1 Crusher Feed Bin 14001 • Existing 14001 fine ore bin at Location 2 500 120 1 Withdrawal chute Existing 500 130 1 Scalping Grizzly 3880 x 1200 mm inclined deck, 80mm splay 22 Rail-section screen decks 500 140 1 Grizzly Underpan 500 150 1 Primary Crusher 3648 single toggle jaw 75mm CSS 150 Existing 250 tph max 500 155 1 MOB crusher maintance hoist Demag 101 electric trolley hoist 11 500 160 1 Crushed Ore Conveyor 900mm x 30m 22 Removed 10/7/04 500 165 1 Crushed ore bin 14001 Exisitng 500 170 1 Screen feed conveyor 1000mm x 20m 22 Removed 10/7/04 EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For: Review Revisions Marked: Italic Item No. No. Description Specification kW Comments Area type Off Supplier Type 500 175 1 Feed Prep Screen 3880 x 1880 double deck inclined vibrating wet screen 22 C/W duck-foot spray nozzles, 38mm & 19mm square slot polyurethane panels 500 180 1 Screen Underpan 500 190 1 2" spray bar Spray water for sorter feed preparation 500 200 1 Coarse sorter feed conveyor 1200mm x 10m 22 0.5 m/s fitted with 6800 Gauss ElectrimagQ RE Magnet head pulley 500 210 1 Conductivity Detector Mogenson Fitted under sorter feed belt 500 220 1 Combined conductivity/RE magentic sorb Mogenson/ Optosort AS 80 coarse particle sorter 80mm / 60 tph max 15 combined magnetic/conductivity sorter tuned for 38 - 75mm 500 230 1 Sorter head chute Bifurcated chute for sorter accept and reject fractions, bypass accept 500 240 1 Fine sorter feed conveyor 1200mm x 10m 22 0.5 m/s fitted with 6800 Gauss Electrimag<! RE magnetic head pulley | EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For. Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 500 250 1 Conductivity detector Mogenson Fitted under sorter feed belt 500 260 1 Combined conductivity/RE magnetic sorb Mogenson/ Optosort AS-80 fine particle sorter 80mm / 60 tph max 15 combined conductivity/ magnetic sorter tuned for 19 - 38m 500 270 1 Fine sorter head chute Bifurcated chute for sorter accept and reject fractions, bypass to accept 500 280 1 Concentrate conveyor 900mm x 30m 22 Recieves screen underflow, coarse and fine concentrate from sorters 500 290 1 Concentrate silo 10001 Existing 500 300 1 Concentrate silo discharge feeder 450 tph Existing 500 310 1 Discard conveyor 900mm x 70m inclined flexowell pocket conveyc 45 >r Rejects to 3370 level rock pass 500 320 1 Discard bin 10001 Exisitng rockpass at 3370 level EqListRevB Project R0016798 McCreedy Underground Pre-concentation Project Equipment List Revision: B Issued on: 23-Dec-04 For Review Revisions Marked: Italic Item No. No. Area Type Off Description Specification kW Comments Supplier Type 500 330 1 Discard Kiruna chute Additional Kiruna chute for back-haul of discards to Location * 500 335 1 Dust extraction for sorters Dondaldson Ai Products 60m3/h, 3kPa 11 c/w reverse-pulse dust collector 500 340 1 Sorter compressor Atlas Copco GA20 lm3/s, 700 kPa compressed air for particle ejectior mechanism 500 350 1 Compressed air reciever 3m3 Air reciever for sorting air supply c/w 2" pipe connections to sorter 500 360 1 Sorting level sump pump 50NB x 1500 mm vertical spindle pump 15 EqListRevB Appendix 6 - Mass Balances Stream # Description Throughput tph Dilution Size Distribution (mm) H20 SG ave max % min max % 1 Level 4360 Ore 14 21 57.00 0 250 3 3.17 2 Level 4660 Ore 14 21 57.00 0 250 3 3.17 3 Level 4810 Ore 14 21 57.00 0 250 3 3.17 4 Level 4945 14 21 57.00 0 250 3 3.17 5 Level 5130 14 21 57.00 0 250 3 3.17 6 Future 170 OB Ore 14 21 57.00 0 250 3 3.17 7 4360 Scalping rejects 0.7 1.05 100.00 125 250 3 2.7 8 4660 scalping rejects 0.7 1.05 100.00 125 250 3 2.7 9 4810 scalping rejects 0.7 1.05 100.00 125 250 3 2.7 10 170 scalping rejects 0.7 1.05 100.00 125 250 3 2.7 11 5130 scalping rejects 0.7 1.05 100.00 125 250 3 2.7 12 4945 scalping rejects 0.7 1.05 100.00 125 250 3 2.7 13 4360 orepass discharge 13.3 19.95 54.74 0 125 3 3.2 14 4660 orepass discharge 13.3 19.95 54.74 0 125 3 3.2 15 4810 orepass discharge .13.3 19.95 54.74 0 125 3 3.2 16 4945 orepass discharge 13.3 19.95 54.74 0 125 3 3.2 17 5130 orepass discharge 13.3 19.95 54.74 0 125 3 3.2 18 170 orepass discharge 13.3 19.95 54.74 0 125 3 3.2 19 Combined 153/170 Ore 79.8 119.7 54.74 0 125 3 3.2 20 Grizzly spray water (typ) 0.7 1.05 100 1 MFD 100 Stream # Description Throughput (tph) Size Distribution (mm) Dilution H20 SG / r « - - ave m a x m m m a x % % I 1 Preconcentration plant f e e d 79.80 119.70 0 125 54.74 3 3.2 2 Grizzly Oversize 3.99 5.99 75 125 54.74 3 3.2 3 Grizzly Underflow 75.81 113.72 0 75 54.74 3 3.2 4 Feed prep screen feed 79.80 119.70 0 75 54.74 3 3.2 5 Feed prep screen oversize 38.94 58.41 9 75 86.00 5 3.1 6 Feed prep screen underflow 40.86 61.29 0 9 21.00 12 3.6 7 DMS plant feed 38.94 58.41 9 75 86.00 5 3.1 8 DMS plant floats 33.49 50.24 9 75 100.00 5 2.7 9 DMS plant sinks 5.45 8.18 9 75 0.00 5 3.8 10 153/170 Concentrate 46.31 69.46 0 75 18.53 11.18 3.6 11 Crusher flushing water 18.52 27.79 100 1 12 Crushed concentrate 64.83 97.25 0 15 18.53 50 2.3 13 Hydraulic hoist spillage 64.83 97.25 0 15 18.53 50 2.3 14 Concentrate thickener feed 64.83 97.25 0 15 18.53 50 2.3 15 Smelter feed 46.31 69.46 0 15 18.53 30 2.96 16 Cones thickener overflow 12.97 19.45 0 100 1 17 DMS Medium feed 19.47 29.21 0 0.06 30 3 18 Sinks screen underflow 16.75 25.12 0 0.06 50 2.6 19 Floats screen underflow 2.73 4.09 0.06 50 2.6 20 DMS water make-up 2.00 3.00 100 1 21 FeSi make-up 0.02 0.03 0.03 0.06 0 4.5 22 Prep screen spray water 4.80 7.20 100 1 23 Process water make-up 25.32 37.99 100 1 MFD 200 Stream # Description Throughput tph Size Distribution (mm) Dilution % %H20 SG ave max mm max 1 3860 Level Ore 54 81 0 750 34 3 3.7 2 4090 Level Ore 54 81 0 750 34 3 3.7 3 Future LMOB Ore 54 81 0 750 34 3 3.7 4 3860 Grizzly Oversize 2.7 4.05 500 750 100 3 3.05 5 4090 Grizzly Oversize 2.7 4.05 500 750 100 3 3.05 6 LMOB Grizzly Oversize 2.7 4.05 500 750 100 3 3.05 7 Pre-concentrator feed 153.90 230.85 0 500 30.5 3 3.8 8 Combined MOB/LMOB rejects 33.86 50.79 19 75 100 5 3.05 9 Combined 153/170 rejects 33.49 8.18 9 75 100 5 2.7 10 Combined rockfill 67.35 58.96 9 75 100 5 2.88 11 Grizzly spray water 2.70 4.05 100 1 OS MFD 400 S t r e a m # Desc r ip t i on T h r o u g h p u t t p h S ize Dis t r ibut ion*(mm) <- %; D i lu t ion % H 2 0 " S G -,' " ' - '- "' a v e m a x m m m a x 1 P re -concen t ra t i on p lant f e e d 153 .90 2 3 0 . 8 5 0 5 0 0 30 .5 3 3.8 2 S c a l p i n g gr izz ly unde r f l ow 1 3 7 . 7 4 2 0 6 . 6 1 0 7 5 27 3 3.9 3 P r imary c r u s h e r f e e d 16 .16 2 4 . 2 4 75 5 0 0 33 .5 3 3.6 4 P r imary c r u s h e r d i s c h a r g e 16 .16 2 4 . 2 4 0 75 30 .5 3 3.6 5 F ine o re b in f e e d 153 .90 2 3 0 . 8 5 0 7 5 30 .5 3 3.8 6 C o a r s e sor ter f e e d 3 8 . 3 2 57 .48 38 7 5 4 0 . 5 5 3.5 7 F ine sor te r f e e d 2 7 . 7 0 4 1 . 5 5 19 38 32 .5 5 3.7 8 P r e p s c r e e n under f l ow 7 6 . 4 9 114 .73 0 19 15.6 10 4 9 C o a r s e sor te r p roduc t 2 2 . 6 8 34 .03 38 7 5 2 .03 5 4.1 10 C o a r s e sor te r re jec ts 15 .54 2 3 . 3 2 38 75 100 .00 5 3 .05 11 F ine sor te r p r o d u c t 2 0 . 9 3 3 1 . 4 0 19 38 1.63 5 4 .1 12 F ine sor te r re jec ts 6 .77 10.16 19 38 100 .00 5 3.1 13 C o m b i n e d M O B / L M O B w a s t e 2 2 . 3 2 33 .47 19 7 5 100 .00 5 3.1 14 C o m b i n e d M O B / L M O B c o n c e n t r a t e 120 .10 180 .16 0 75 10.60 7 4 .1 15 P r e p s c r e e n s p r a y w a t e r 7 .70 11.54 100 1 M F D 500 Appendix 7 - Cost Estimate Spreadsheets 166 Area Description Area Cost Mining and Processing Plant 100 153 Mining 200 MOB Mining 300 153 Preconcentration 400 Concentrate Hoisting 500 MOB Preconcentration 600 Excavations & Support Mechanical subtotal $4,434,747.74 Subtotal Direct Costs $879,646.00 $947,694.75 $2,359,022.62 $2,554,237.52 $3,856,927.28 $1,598,570.41 $12,196,098.58 Plant Infrastructure Piping & Valves sum Substations Substation excavations 216 m3 Office/Ctrl Office/Ctrl Excavation 216 m3 Subtotal P&l Subtotal Infrastructure $1,609,813.43 $75,390.71 $33,937.92 $75,390.71 $33,937.92 $1,828,470.69 $14,024,569.26 Indirect Costs EPCM @25% $701,228.46 Spares, First Fill & Startup @60% Mech Maintenance Y1 @10% Mech Subtotal Indirect Costs $3,506,142.32 $2,660,848.64 $443,474.77 $6,610,465.73 Subtotal Direct & Indirect Contingency 15% Total $20,635,034.99 $3,095,255.25 $23,730,290.24 Accuracy -10% -$2,373,029.02 30% $7,119,087.07 $21,357,261.22 $30,849,377.31 167 Area 100 153 Orebody Item Quantity Mechanical Mechanical Erection Civil Structural & Platework Strucural Erection E&l Subtotal 100 6 O r e p a s s Gr izz ly 2 0 0 0 0 . 0 0 1320.00 3475 .32 13339.42 2440 .00 7 6 9 9 . 6 0 2 8 9 6 4 6 . 0 0 2 0 0 1 A t las C o p c o S T 1020 Rockfi l l L H D 590000 .00 0.00 0.00 0.00 0 .00 0 .00 5 9 0 0 0 0 . 0 0 Tota l $610 ,000 .00 $879 ,646 .00 Area 200 MOB Orebody Item Quantity Description Mechanical Mechanical Erection Civil Structural & Platework Strucural Erection E&l Subtotal 100 6 O r e p a s s Grizzly 24698 .75 1630.12 4291 .80 16473.35 3013.25 9508 .52 3 5 7 6 9 4 . 7 5 140 1 A t l a s C o p c o S T 1 0 2 0 Rockf i l l L H D 590000 .00 0.00 0.00 0.00 0 .00 0 .00 590000 .00 Tota l $614 ,698 .75 $947 ,694 .75 Area 300 153 Pre-concentration O Item Quantity Description Mechanical Mechanical Civil Structural & Strucural E&l Subtotal 100 1 Kiruna tip 184617.20 12184.74 32080.19 123134.31 22523.30 71073.89 445613.62 120 1 CVB 1330 scalping grizzly 24698.75 1630.12 4291.80 16473.35 3013.25 9508.52 59615.79 130 1 153 Primary Crusher - C80 190754.96 12589.83 33146.72 127228.02 23272.11 73436.80 460428.43 140 1 Primary Crusher Electric Hoist 5143.62 339.48 893.79 3430.65 627.52 1980.19 12415.25 160 1 Crushed ore conveyor 9469.29 624.97 1645.44 6315.74 1155.25 3645.48 22856.17 170 1 Feed prep screen 55587.02 3668.74 9659.13 37074.93 6781.62 21399.88 134171.32 200 1 DMS feed conveyor 9469.29 624.97 1645.44 6315.74 1155.25 3645.48 22856.17 210 1 DMS Drum 200000.00 13200.00 34753.20 133394.19 24400.00 76995.95 482743.34 225 1 Drum Electric hoist 5143.62 339.48 893.79 3430.65 627.52 1980.19 12415.25 250 1 Drain & Rinse Screen 20000.00 1320.00 3475.32 13339.42 2440.00 7699.60 48274.33 300 1 DM Sump agitator 14024.37 925.61 2436.96 9353.85 1710.97 5399.10 33850.86 310 1 DM Pump Warman 6/4 4734.63 312.49 822.72 3157.86 577.62 1822.74 11428.06 315 1 Pipe Densifier 2000.00 132.00 347.53 1333.94 244.00 769.96 4827.43 320 1 CM Sump Agitator 14024.37 925.61 2436.96 9353.85 1710.97 5399.10 33850.86 330 1 CM Pump 8/6 9241.35 609.93 1605.83 6163.71 1127.44 3557.73 22306.00 335 1 Magnetic Separator 30000.00 1980.00 5212.98 20009.13 3660.00 11549.39 72411.50 350 1 Make-up tank agitator 14024.37 925.61 2436.96 9353.85 1710.97 5399.10 33850.86 360 1 Make-up Pump 4/3 4506.72 297.44 783.11 3005.85 549.82 1735.00 10877.94 370 1 Discard conveyor 15782.14 1041.62 2742.40 10526.23 1925.42 6075.81 38093.62 380 1 Concentrate conveyor 4734.64 312.49 822.72 3157.87 577.63 1822.74 11428.09 400 1 Concentrate crusher feeder 9469.29 624.97 1645.44 6315.74 1155.25 3645.48 22856.17 405 1 Concentrate crusher 142214.07 9386.13 24711.97 94852.66 17350.12 54749.54 343264.48 410 1 Concentrate crusher hoist 5143.62 339.48 893.79 3430.65 627.52 1980.19 12415.25 420 1 Preconc sump pump 2556.98 168.76 444.32 1705.43 311.95 984.39 6171.83 Total $977,340.30 $2,359,022.62 Area 400 Concentrate Hoisting Item Quantity Description Mechanical Mechanical Erection Civil Structural & Platework Strucural Erection E&l Subtotal 110 1 Shaft bottom repulper 14024.37 925.61 2436.96 9353.85 1710.97 5399.10 33850.86 120 2 Shaft bottom concentrate pump 200000.00 13200.00 34753.20 133394.19 24400.00 76995.95 965486.68 130 1 Shaft bottom process water turbine 20000.00 1320.00 3475.32 13339.42 2440.00 7699.60 48274.33 170 1 Mid shaft repulper 14024.37 925.61 2436.96 9353.85 1710.97 5399.10 33850.86 180 2 Mid shaft concentrate pump 200000.00 13200.00 34753.20 133394.19 24400.00 76995.95 965486.68 190 1 Mid shaft process water turbine 20000.00 1320.00 3475.32 13339.42 2440.00 7699.60 48274.33 230 1 Surface concentrate repulper 14024.37 925.61 2436.96 9353.85 1710.97 5399.10 33850.86 240 2 Surface concentrate pump 13816.00 911.86 2400.75 9214.87 1685.55 5318.88 66695.82 270 1 Concentrate thickener 116999.84 7721.99 20330.59 78035.50 14273.98 45042.57 282404.47 280 1 Thickener underflow pump 8955.68 591.07 1556.19 5973.18 1092.59 3447.75 21616.47 310 1 Shaft bottom concentrate dam 20000.00 1320.00 3475.32 13339.42 2440.00 7699.60 48274.33 320 1 Concentrate area sump pump 2556.98 168.76 444.32 1705.43 311.95 984.39 6171.83 Total $644,401.61 $2,554,237.52 Area 500 MOB Preconcentration Item Quantity Description ' 10 3 1 Kiruna tip 13( 3 1 Scalping grizzly 14( 3 1 36 x 48 jaw crusher (Existing) 15! 5 1 Crusher maintenance hoist 17! 5 1 Feed prep screen 20C ) 1 Coarse sorter feed conveyor 21C ) 1 Coarse sorter conductivity detector 22C 1 AS80 Coarse particle sorter 24C 1 Fine sorter feed conveyor 25C 1 Fine sorter conductivity detector 26C 1 AS80 Fine particle sorter 28Q 1 MOB Concentrate conveyor 300 1 Concentrate silo discharge feeder (Existing) 310 1 Discard conveyor 330 1 Discard silo Kiruna chute 335 1 Dust Extraction 340 1 Sorter compressor GA20 350 1 Compressed air receiver 360 1 MOB preconcentrator sump pump Total Mechanical Mechanical Erection TcSvii Structural & Platework Strucura Erection "[i&i " Subtotal 184617.2C ) 12184.7' t 32080.1 3 123134.31 22523.31 3 71073.8 3 445613.62 24698.7! 5 1630.12 > 4291.8( 3 16473.35 3013.2! 5 9508.5-2 59615.79 333.5! > 1737.6( 5 6669.71 613.2( 3 3849.8( 3 23203.98 5143.62 ! 339.4c" 1 893.7J ) 3430.65 627.52 > 1980.1 < 3 12415.25 55587.02 3668.74 9659.1C J 37074.93 6781.62 ! 21399.8! J 134171.32 5786.7S 381.92 1005.5E > 3859.62 705.9S ) 2227.8C ) 13967.66 3207.93 211.72 557.4; 2139.60 391.37 1234.9S 1 7743.04 524036.90 34586.44 91059.78 349517.39 63932.5C 201743.6C 1264876.62 5786.79 381.93 1005.55 3859.62 705.99 2227.80 13967.66 3207.93 211.72 557.43 2139.60 391.37 1234.99 7743.04 524036.90 34586.44 91059.78 349517.39 63932.50 201743.60 1264876.62 9469.29 624.97 1645.44 6315.74 1155.25 3645.48 22856.17 0.00 0.00 0.00 0.00 0.00 0.00 19289.29 1273.09 3351.82 12865.39 2353.29 7425.98 46558.87 184617.20 12184.74 32080.19 123134.31 22523.30 71073.89 445613.62 10000.00 660.00 1737.66 6669.71 1220.00 3849.80 24137.17 20064.03 1324.23 3486.45 13382.12 2447.81 7724.25 48428.88 6200.45 409.23 1077.43 4135.52 756.45 2387.05 14966.12 2556.98 168.76 444.32 1705.43 311.95 984.39 6171.83 $1,588,307.07 $3,856,927.28 Area 600 Excavations & Support Item Quantity Description Excavation vol m3 cost Support m2 Support cost Subtotal 100 9 Orepass excavat ions 100 15712 600 1170.326 151940.9192 200 2 153 feed silo 700 109984 4200 8192.28 236352.541 300 1 153 crushing excavat ion 155 24353.6 930 1814.005 26167.60275 400 1 Screen feed conveyor drift 89.6 14077.95 537.6 1048.612 15126.56262 400 1 153 feed prep excavat ion 176 27653.12 1056 2059.773 29712.89087 500 1 D M S feed conveyor drift 108.4 17031.81 650.4 1268.633 18300.4396 500 1 D M S plant excavat ion 1040 163404.8 6240 12171.39 175576.1733 600 6 153 concentrate conveyor drift 122.8 19294.33 736.8 1437.16 124388.9659 700 1 153 discard conveyor drift 123.1 19341.47 738.6 1440.671 20782.14128 750 1 Concentrate pump stat ion 176 27653.12 1056 2059.773 29712.89087 800 1 M O B feed silo 990 155548.8 5940 11586.22 167135.0111 800 1 M O B crushing excavat ion 260 40851.2 1560 3042.847 43894.04332 900 1 Crushed ore silo 700 109984 4200 8192.28 118176.2705 995 1 Feed prep excavat ion 156 24510.72 936 1825.708 26336.42599 1000 2 Sorter feed conveyor drifts 112 17597.44 672 1310.765 37816.40656 1005 1 M O B sorting excavat ion 224 35194.88 1344 2621.53 37816.40656 1010 1 M O B fine product silo 480 75417.59 2880 5617.563 81035.15691 1020 1 M O B discard conveyor drift 540 84844.79 3240 6319.759 91164.55152 1030 1 M O B waste silo 990 155548.8 5940 11586.22 167135.0111 Total $1,598,570.41 Appendix 8 - Financial Evaluation Spreadsheets 174 Pre-concentrata all ore at Location 8 & 9 2,004 2,005 2,006 2,007 2,008 2,009 2,010 2,011 2,012 2,013 2,014 2,015 2,016 2,017 2,018 Orebody MOB 1,808 1,833 1,749 1,705 1,666 1,303 221 tons/day LMOB 1,027 1,557 1,568 1,617 1,617 1,617 1,617 1,274 1,176 1,176 1.176 956 402 153 1,078 1,078 1,078 1.078 1,078 1,078 1,078 1,078 588 170 165 485 515 662 659 662 1,073 1,846 1,720 1,507 868 774 342 Saving MOB 4,181,673 4,239,495 4,045,214 3,943,447 3,853,245 3,013,673 511,145 0 0 0 0 0 0 0 0 LMOB 2,585,985 5,723,133 0 3,920,524 5,723.133 0 3,948,222 5,723,133 935,068 4,071,604 5,723,133 2,748,533 4,071,604 5,723,133 2,918,546 4,071,604 5,723,133 3,751,606 4,071,604 5,723,133 3,734,605 3,207,930 5,723,133 3,751,606 2,961,166 3,121,709 6,080,776 2,961,166 0 10,461,428 2,961,166 0 9,747,376 2,407,207 0 8,540,288 1,012,235 0 4,919,025 0 0 4,386,319 0 0 1,938,141 153 170 Revenue MOB -3,116,368 -3,159,459 -3,014,672 -2,938,831 -2,871,609 -2.245,922 -380,928 0 0 0 0 0 0 0 0 LMOB -2,000,154 -3,032,367 -3,053,790 -3,149,221 -3,149,221 -3,149,221 -3,149,221 -2,481,204 -2,290,342 -2,290,342 -2,290,342 -1,861,877 -782,923 0 0 153 12,358,273 12,358.273 12,358,273 12,358,273 12,358,273 12,358,273 12,358,273 12,358,273 6,740,876 0 0 0 0 0 170 0 0 1.910,780 5,616,536 5,963,950 7,666,282 7,631,540 7,666,282 12,425,862 21,377,578 19,918,436 17.451,793 10,051,862 8,963,296 3,960,526 Capital 30,849,000 UDCF -29,783,694 20,049,599 22,852,228 28,373,474 28,867,922 31,189,428 30,500,152 30,226,020 29,040,047 32,509,829 30,336,636 26,537,410 15,200,199 13,349,616 5,898,667 NPV IRR 134,710,632 79% 1 Base Working Costs Mining Backfil Mucking Haul Crush/ Hoist Subtotal Surfac Mill Tailinc Subtota Cost/t Convey Haul M O B 17.08 4.38 4.27 3.30 1.37 1.08 31.48 4.45 7.34 2.00 13.79 45.27 LMOB 26.44 4.38 3.87 4.10 1.37 1.08 41.24 4.45 7.34 2.00 13.79 55.03 153 52.90 9.29 5.97 10.00 1.37 1.08 80.61 4.45 7.34 2.00 13.79 94.40 170 81.89 9.29 5.97 12.00 1.37 1.08 111.60 4.45 7.34 2.00 13.79 125.39 2 Preconcentrate MOB, LMOB, 153 & 170 at Location 8 Mining Backfill Mucking Preconc Haul Hoist Transport Mill Tailing s Total cost tons cost tons cost tons cost tons cost tons cost tons cost tons cost tons cost tons M O B 17.08 1800 4.38 885 4.27 1710 1.08 1710 3.30 1404 1.08 1404 4.45 1404 7.34 1404 2.00 1404 LMOB 26.44 1500 4.38 585 3.87 1425 1.08 1425 4.10 1170 1.08 1170 4.45 1170 7.34 1170 2.00 1170 153 52.90 1100 9.29 1045 5.97 1045 2.34 1045 5.00 1045 1.08 330 4.45 330 7.34 330 2.00 330 170 81.89 600 9.29 570 5.97 570 2.34 570 6.00 570 1.08 180 4.45 180 7.34 180 2.00 180 Revised Working Cost $/ROM t 17.08 26.44 52.90 81.89 2.15 1.71 8.83 8.83 4.06 3.68 5.67 5.67 1.03 1.03 2.22 2.22 2.57 3.20 4.75 5.70 0.84 0.84 0.32 0.32 3.47 3.47 1.34 1.34 5.73 5.73 2.20 2.20 1.56 1.56 0.60 0.60 38.49 47.65 78.83 108.77 Pre-concentration Grade Balance 1 In situ Orebody Mined tons Cu% Ni% TPM g/t dilution MOB 1800 0.9 1.85 0.64 22% LMOB 1500 0.92 1.94 0.96 22% 153 1100 11.42 1.22 10.56 70% 170 600 7.01 0.92 14.4 70% 2 Mine 3 Muck Waste t Product t Cu% Ni% TPM% MOB 90 1710 0.945 1.9425 0.672 LMOB 75 1425 0.966 2.037 1.008 153 55 1045 11.991 1.281 11.088 170 30 570 7.3605 0.966 15.12 5 Haul 6 Preconcentration Waste t Product t Cu% Ni% TPM g/t Value/t MOB 396 1404 0.988% 2.189% 0.74 159.5913 LMOB 330 1170 1.010% 2.296% 1.11 189.4813 153 715 330 18.832% 1.867% 17.05 602.4375 170 390 180 11.559% 1.408% 23.26 602.4375 7 Hoist 8 Surface Haul 9 Mill 10 Dump tailings Track Meta Values - RENU In situ Preconc Mill Precon+Mill Recovery 100% 95% 85% 80.75% MOB 130.8125 124.2719 111.1906 105.6310938 LMOB 155.3125 147.5469 132.0156 125.4148438 153 354.375 318.9375 301.2188 286.1578125 170 354.375 318.9375 301.2188 286.1578125 5 Notes Option 1 Preconc all ore at Coleman shaft Nickel price $3.50 US$/lb Exchange rate 1.25 CDN:US Mining costs include drill. Blast & U/G Services. Cost do not include distribution, A&D & taxes 170 mining costs inflated by 54% over 153. Haul costs as per Kiruna Productivity graph. Preconcentration plant costs CDN$10m & $2/t processing cost Precon waste backhauled to Location 9 @ theoretical 50% haul cost. Additional rockfill reduces sandfill required by 50% Recovery across pre-concentrator = 95% No recovery improvement at mill Smelter fees of 20% deducted from revenue Option 2 Preconc 153/170 only at Coleman shaft MOB & LMOB not preconcentrated 153 & 170 pre-concentrated at shaft. Waste hauled back to Location 9. Concentrate pumped to surface. Preconcentration recovery to produce smeltable con = 90% Preconcentration plant and hydraulic hoist plant costs CDN$4m total. Hydraulic hoisting = hoisting cost/t (Study) Mill recovery is 85%. Mill is bypassed thus metal loss credited as additional revenue. Option 3 Preconc 153/170 at Location 8 Premutation to Option2:153/170 preconcentrated at Location 8, no waste backhaul. Option 4 Preconc 153/170 and MOB/LMOB separately at Coleman shaft MOB/LMOB and 153/170 preconcentrated separately at Coleman shaft. 153/170 ore preconcentrated to smeltable con @ 90% recovery 153/170 con bypass mill, additional metal revenue credits High grade pre-con waste fed to MOB/LMOB pre-concentration plant tuned for rejection of barren waste only Final waste reject back-hauled to Location 9 


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