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By J. Olsson

The paper describes and discusses the use of stainless steels for leaching plants emphasising the duplex family, i.e., austenitic-ferritic stainless steels. This group has generally twice the proof strength of the more common 300-series austenitic grades, which implies possibilities to reduce the gauge at construction of pressure vessels and atmospheric tanks and vessels and thereby possibilities to reduce the weight. The duplex grades have also higher surface hardness resulting in better resistance to mechanical wear, i.e., abrasion and erosion corrosion. The lower contents of nickel and molybdenum imply lower cost and a more stable price considering the present market situation for these alloying elements. Corrosion data from laboratory and pilot plant tests, as well as examples of applications of duplex and highly alloyed austenitic grades, are presented. Cu2007 - Volume IV (Book 1) The John E. Dutrizac International Symposium on Copper Hydrometallurgy Edited by P.A. Riveros, D.G. Dixon, D.B. Dreisinger, M.J. Collins Toronto, Canada,

Jan 1, 2007

By Roger Reed

It is well known that the performance and carbon emissions associated with the modern gas turbine engine depend critically upon the properties of the nickel-based superalloys, particularly those used in single crystal form. However, the design of all alloys in current use has been essentially empirical, or at best heavily reliant upon regression/neural network analysis of experimental datasets. In this paper, alloy design approaches [1] are described which are able to identify new alloy composition spaces which are worthy of experimental study. It is shown that a large amount of design data can be generated, which can be interpreted in different ways. Of course, the success of the approach depends heavily upon the accuracy of the underlying sub-models. Estimates of the creep resistance, cost, castability, density and oxidation resistance are made. It is demonstrated that these methods allow quantitative trade-offs to be made between the properties of interest.

Jan 1, 2010

By J. M. Menacho

As usually occurs in technological development, heap leaching knowledge has been built upon two parallel lanes, theory developed on fundamentals but relatively far from industrial applications, and operational know-how generated from empirical trial and error. As a contribution to close the gap between the above-mentioned lanes, the De Re Metallica process consulting group has been developing and applying fundamentalsbased modeling tools for industrial process design and process optimization of existing plants over the last ten years. As result, a library of models is now available with application to the whole production chain from blasting to electro winning. This presentation is centered in heap leaching and several real case studies are developed, showing how paradigms are broken when fundamentals are applied. Cu2007 - Volume IV (Book 1) The John E. Dutrizac International Symposium on Copper Hydrometallurgy Edited by P.A. Riveros, D.G. Dixon, D.B. Dreisinger, M.J. Collins Toronto, Canada,

Jan 1, 2007

By Julie Q. Shang

The electrokinetic (EK) technology for dewatering of two tailings from the uranium mines in Northern Saskatchewan is studied in bench scale experiments. The study includes four phases: 1. Characterization of two uranium mine tailings; 2. Assessment of voltage loss on the coated Ti anode and water interface. 3. EK cell test series to measure electrokinetic permeability (ke) of two tailings; 4. EK dewatering test series to study electrokinetic enhanced gravity dewatering on two tailings samples. The results have demonstrated that both tailings responded positively to the EK process. For the tailings containing gypsum, the EK dewatering should be a short term process (within one hour) with anode conditioning (pH adjustment). The tailings containing clay minerals responded well to EK dewatering, in which pH adjustment is optional and will further enhance the EK flow. Both tailings as tested have low hydraulic conductivity and are particularly suitable for EK dewatering and consolidation.

Aug 1, 2010

By E. Ferguson

In evaluating refractory designs for cast house applications, thermal calculations are frequently performed to determine cold face shell temperatures, temperature profiles and metal freeze plane locations within the refractory layer, and heat loss in the steady state condition. A problem arises when looking at molten metal transport crucible designs in that these structures are usually operating under non-steady state conditions throughout their production cycle. Because the molten metal temperature varies with time, and material properties such as thermal conductivity are temperature dependant, mathematical solutions are at best... complicated. This paper presents a relatively simple, iterative solution, that can be set up on a standard computer spreadsheet program, to calculate heat transfer in non-steady state conditions. These calculations can provide curves for (1) molten metal temperature vs time and (2) cumulative heat loss vs time, given variable starting conditions, to aid in the evaluation of molten metal transfer crucible refractory designs.

Jan 1, 2004

By F. Schulz

NA's primary smelter has continually increased its production output in the past decade, making it necessary to improve the lifetime of the casting moulds for copper anodes. The current throughput of more than 1,100,000 tonnes per year of copper concentrates amounts to an output of about 450,000 tonnes per year of anode copper at NA's Hamburg plant. The anodes are cast on a single casting wheel with 24 anode moulds. Besides the stable chemical analyses, the reproducible geometry of the anodes is an important precondition for the success of the subsequent electro-refining process. The anode moulds are subjected to thermally-induced alternating stress and side reactions with the melt. Typical signs of wear will be shown and discussed. Different processes to produce anode moulds will be introduced and the results in terms of their lifetime and the quality of the anodes themselves will be discussed. Finally an outlook is given on possible efficiency improvements.

Jan 1, 2007

By K. Koyama

As a study to establish a new energy-saving copper recycling process, the copper deposition at a cathode and oxidation of copper(I) at an anode from ammoniacal alkaline solutions containing Cu(I) or Cu(II) were investigated in order to examine the feasibility of electrolysis in this solution.. The current efficiencies were greater than 95% in the current density range of 200 to 1000 A m-2. The current efficiency decreased with the increasing Cu(II) concentration and temperature. In the batch-type electrolysis at the anode, during the initial stage, the current efficiencies were 100%, and the anode potentials ranged from 0 to 0.15V vs. SHE at a current density of 100 A m-2. During the subsequent stage, the current efficiency decreased with time, and the anode potential was around 1.35V vs. SHE where oxygen evolution at the anode surface was observed. As a result, the present electrolytic method can be the core technology in the said copper recycling process.

Jan 1, 2007

By S. Parada, P. T. Jones, M. Guo, S. Smets, E. Boydens

The corrosion behaviour of a pitch-bonded MgO-C refractory in contact with a vacuum oxygen decarburisation slag with high Al2O3 content (12-17 wt%) is investigated under VOD process conditions by rotating finger tests in a vacuum induction furnace. The refractory degradation mechanisms are studied by the characterisation of the corroded microstructures using scanning electron microscopy, equipped with energy dispersive spectroscopy. The influence on the refractory wear of the process temperature, corrosion time and the Al2O3 content in the slag is discussed. The degradation mechanisms include slag infiltration, direct MgO dissolution from the refractories into the penetrated slag, erosion of magnesia grains at elevated temperatures, removal of carbon through the MgO-C reaction, and finally, oxidation of refractory carbon by reducible slag components. Both a porous MgO layer and a dense (Mg)(A1,Cr)2O4 spinel at the slag/brick interface were observed. Their formation mechanisms and industrial relevance are discussed.

Jan 1, 2004

By James J. Roettger, Robert J. Mitchell

"The results of a series of model stope pours designed to investigate working pressures on bulkheads during hydraulic pouring of classified tailings backfill are reported. Total bulkhead pressures and fill water pressures were monitored under the condition of a fully saturated backfill. Similitude of the model testing is discussed and equations are developed to estimate bulkhead pressures under prototype conditions. The model results show that good base drainage, combined with a bulkhead setback, can reduce the expected bulkhead pressures to values such that timber or concrete block structures should be capable of supporting a saturated backfill. Introduction Bulkheads are used in underground running operations to block drift openings in order that hydraulically placed sand backfill can be contained in stope pours. Whenever possible, the sand fill is designed to have sufficient percolation rate so that excess hydraulic transport water will gravity drain out through the bulk headed drifts. This avoids costly decanting systems. Early bulkhead design assumed that the bulkhead pressure could be as high as 0h = Hy where H is the fill height and ""y is the saturated unit weight of the fill (Thomas et al., 1979)(1). A typical design pressure for a 50 m fill height would be 1250 kPa giving a design force of 12,500 kN on a bulkhead located in a drift of 10m2 in area . Typical bulkheads were 1 m thick heavily reinforced concrete structures formed and poured into a notched drift. Not only were these expensive structures but the associated problems of inside drainage systems and curing time delays increased production costs. As experience with bulkheads for large bulk pours increased and fill pressure measurements became available [see, for examples, Libby et al., 1975(2), Askey et al., 1978(3)] , 'free drainage' timber and concrete block bulkheads were designed. Substantial savings have been realized using these types of bulkheads but greater engineering design and monitoring of backfill pours is necessary to ensure that backfill pressures do not exceed the bulkhead design pressures.The effective and total stress components of fill pressure, the engineering concepts associated with the 'free drainage' design and some detailed design considerations have been discussed by Smith and Mitchell (1982)(4) . It is important to the safe use of these designs, however, that the basic engineering concepts be demonstrated and communicated to those responsible for the day-to-day backfill operations. This paper describes and explains observations from a series of model studies designed to measure bulkhead pressures in a saturated fill."

Jan 1, 1984

By F. W. Wellmer

"This paper discusses the changing nature of reserve/consumption ratios, using data from Canadian operations and focussing particularly on copper.IntroductionStudies in world metal consumption and world metal reserves are done regularly by the U.S. Bureau of Mines, the German Federal Institute of Geosciences and Raw Materials, and other institutions. Formerly, on the basis of these data, the following figures were calculated:• a static lifetime of reserves-reserves at year end divided by the consumption of the previous year;• a semidynamic lifetime of reserves-reserves at year assumed constant growth rate.These figures were severely criticized because they were purely theoretical and did not consider the availability of the raw materials.Recent practice is to calculate reserve/consumption ratios (R/C ratios) for periods(1) or for different years(2,3) and to interpret the trend. For base metals, these R/C ratios are increasing (Table 1). This had led to the conclusion that today the relative availability of base metal reserves has increased as compared to 10 or 15 years ago(4,5) As there are normally numerous deposits for each commodity, the anticipated production rates of deposits are irrelevant to this conclusion as long as the statistical distribution of possible production rates does not change significantly.The above conclusion is only justified if the quality and the nature of the data base remain the same. This is certainly the case for the consumption data. It is also assumed for the quality of the reserve figures, which usually take into account only proven and probable reserves. For the nature of the reserve figures, however, this does not seem to be true.Undoubtedly, the reserves increased in absolute terms between 1966 and 1978. To put the absolute reserve figures into perspective, it is common, as said above, to compare them with the mine production, which increased in the free world between 1966 and 1978; e.g. for copper by 40%, for Pb by 17% and for zinc by 36%. However, the reserves are also influenced by other factors, such as average size of the mines, changes in the grade of the mined ore, price-to-cost relationships and the like.The purpose of this paper is to examine the nature of the reserve figures in more detail. Because the completeness of mining data published in Canada is rather exceptional, Canadian .data serve as an example and it is assumed that the conclusions can be extrapolated to other mining countries or at least to those with free market systems."

Jan 1, 1981

By Euler M. De Souza

"This paper presents a study developed to assess the potential use of salt tailings as backfill to control surface subsidence and as a mine bulkhead used to isolate areas of potential brine inflows. A series of index tests, centrifuge model tests and quasi-static creep consolidation tests were performed to evaluate the creep and consolidation properties of salt backfl1l and to evaluate the effect of brine inflow and brine infiltration on backfill behaviour. The centrifuge models were specifically designed to study the mechanical behaviour of salt backfill under the influence of brine inflow and long-term creep. Test results indicated that salt tailings could be applied underground as a bulkhead to seal off areas of potential inflow with the additional advantage of controlling room closure and local back instability. It is hoped that this study will assist mine planners to develop contingency plans to control brine inflow occurrences. IntroductionThe use of salt tailings as backfill is a technique that would assist in the control of potential brine inflows and surface subsidence, and would have a very positive impact on the long term adverse environmental effects associated with surface salt tailings disposal. A centrifuge testing program was designed to evaluate the behaviour of salt tailings as backfill; the effectiveness of salt backfill as a bulkhead used to isolate areas of potential brine inflow and the potential use of salt backfill to control surface subsidence. The testing program was also used to evaluate the creep-consolidation properties of salt backfill and to evaluate the effect of brine inflow and brine infiltration on backfill behaviour. This paper presents the results of the experimental centrifuge testing program and a comparison of results with models tested under I g using an electrohydraulic servo-controlled loading frame (MTS) and a creep consolidation cell. Applicability of the results to practical underground soft rock mining is also discussed.The centrifuge model tests indicated that salt tailings could be applied underground as bulkheads to seal off areas of potential inflow with the additional advantage of controlling room closure and local back instability. Because the applied stresses (1.2 MPa) and fill consolidation were relatively low, the centrifuge simulated bulkhead did not retain all of the inflow brine. At typical field stresses it is expected that the fill would be able to seal off panels with brine inflows. Further testing at higher stresses are suggested to determine the consolidation and stress levels at which the fill would completely retain any potential brine inflow head and positively act as mine bulkhead."

Jan 1, 1994

By A. Abadías Llamas, N. Bartie, M. Stelter

Evaluating the economic viability as well as the sustainability of the Circular Economy (CE) system requires a deep understanding of the distribution of all elements, compounds, alloys, materials etc. in flows. In this paper, the circularity of the copper value chain, including primary and secondary processing, is rigorously evaluated. The studied system comprises the metallic copper production from primary sources (from mineral to metal), copper-containing commodity production (copper is mixed with other metals) and copper recycling through secondary smelting to close the loop. This is linked to photovoltaic (PV) panels and battery storage. A simulation model of this system is created using HSC Sim, considering more than 30 elements (and its various compounds), 180 unit operations and 800 flows. From the mass and energy balances obtained through the simulation, an exergy analysis is conducted to evaluate the resource consumption from a second law (entropy) perspective. Additionally, these results are complemented through a Life Cycle Assessment (LCA), the recovery of technology elements and by-products is discussed, while quantifying the losses through the value chain. Through the digitalization of the complete system, a better CAPEX and OPEX understanding of the metal recovery and losses can be obtained, as well as the associated resource consumption and environmental impacts. New flowsheets and technologies can be evaluated. Several scenarios show how the resource consumption and the environmental impacts are affected by the recovery of different materials to produce different products. INTRODUCTION A key challenge of the circular economy (CE) is the linkage and optimization of both resource and energy efficiency of the (renewable) energy infrastructure and the resource cycle depicted by Figure 1 simultaneously. In addition, the CE requires the linkage of all stakeholders as shown by Figure 1. This has been discussed in some detail by Reuter and Bartie (2019) and Reuter et al. (2019). Copper and all its associated technology elements is one of the key enabling metals that is an economical copper colored thread through CE society (Reuter & Kojo, 2014). The Metal Wheel (Figure 2) shows the linkage of and recovery of various elements associated with copper and other carrier metals. This reveals that copper acts as an important carrier metal because of its properties i.e. the wealth of technology elements it carries from geological ores, but also its capacity to act as solvent of valuable elements and thence a key closer of various complexly connected element and material cycles when these technology elements are released during copper refining. The same applies to all other carrier metals of modern CE society as shown for example recently in a lead policy brief (Blanpain, Reuter & Malfliet, 2019).

Jan 1, 2019

By K R. Brentley

"Tshepong Mine, Harmony’s largest gold producing mine, is situated in the Free State portion of the Witwatersrand Basin, east of Odendaalsrus and north of Welkom, South Africa. The Basal reef is a narrow tabular conglomerate and is the main gold bearing ore body. The strata strikes mainly in a northsouth direction with an average dip of 22 degrees. Currently mining takes place at depths of between 1 690m and 2 150m below surface. The Basal reef unit consists of a conglomerate ranging between a few and 30cm in thickness. Some 80-90% of the gold occurs in the lower portion of the conglomerate. Directly above the conglomerate is a quartzite placer unit, the Basal quartzite, which is a clean, hard competent rock and varies in thickness between 50cm and 200cm. The Basal unit is overlain by the Khaki shale which can vary between 30cm and 500cm in width. It is weak and incompetent khaki coloured chloritic shale.A method of stoping, termed “undercut”, is practiced in order to remove the lower 15-20cm of the gold rich strata. The method entails removing this portion along with some 70-80cm of footwall host rock which is a strong competent quartzite. The hard competent quartzite overlying the conglomerate is left in-situ to form a competent hangingwall which can stabilize and prevent exposure of the weak Khakishale. This method of extraction was developed through trial and error and has been used reasonably successfully since its inception in the early 1960’s where the method was developed to extract the Basal reef at the Western Holdings Mines. Recently prospect drilling and a number of localised hangingwall failures indicated that the residual quartzite beam could be less than 60cm and in some instances as thin as 30cm in certain areas of the mine. The result of this is a less stable hangingwall which not only poses a safety risk to workers in these areas but also had the potential of locking up some 840 000 Tonnes of gold bearing ore. At the present gold price this amounts to revenue of approximately ZAR1.5 billion (U.S. $143 million).This paper discusses the method of “undercut stoping” and existing practices in place at Tshepong Mine. It also explores new initiatives such as the use of a simple freely-supported elastic beam solution to perform back analysis, on-site data capturing of beam thickness and support geometries in order to develop empirical design charts and numerical modelling to assist in the verification of the empirical design method."

Jan 1, 2015

By Takao Oishi, Go Ohara, Toru Kitazaki, Masahiro Tanaka

"The High Pressure Acid Leach (HPAL) plant for nickel-cobalt mixed sulphide production at Coral Bay Nickel Corporation (CBNC) in the Philippines has been in operation since 2005 with a capacity of 10,000 t/year of nickel. Subsequent to the commissioning of the second plant in 2008, the production capacity increased to 22,000 t/year of nickel (Llerin et al., 2011). CBNC’s annual nickel production reached more than 24,000 tons of nickel in 2014. The concepts to achieve this capacity are the high availability and the reciprocally flexible operation of both plants. The availability depends on some factors, such as the control of impurities in the process and the equipment condition, to mention a few. In the HPAL process, it is not only the autoclave operating conditions that affect nickel extraction but also the impurities in the ore greatly influence it. The major element in the ore is iron and most of this iron is precipitated as Fe2O3, while the rest remains as Fe2+ or Fe3+ sulphate in solution. These ions in solution reduce the availability, so it is important to control the formation of iron ions by controlling the impurities in feed ore as well as the HPAL conditions. This paper presents the general HPAL process and the behaviour of impurities, focusing on iron behaviour in the HPAL process.INTRODUCTIONCoral Bay Nickel Corporation ProfileCoral Bay Nickel Corporation (CBNC) was founded on April 2002, as a joint venture of four companies; namely, Mitsui Corporation (18%), Sojitz Corporation (18%), Nickel Asia Corporation (10%) and Sumitomo Metal Mining Corporation (SMM) with a controlling stake of 54%. Aside from its capital share, Nickel Asia Corporation also supplies the ore and limestone for the High Pressure Acid Leach (HPAL) plant located in Rio Tuba, Philippines, as shown in Figure 1 (Tsuchida, 2008). The objective of the CBNC project is to produce 22,000 t/year (original design) of nickel and 1,500 t/year of cobalt, as sulphide, by the treatment of nickel laterite ore which is a low grade nickel oxide ore in the Philippines (Mines and Geosciences Bureau, 2014). The sulphide product has been exported as a raw material to the SMM Nickel Refinery in Japan for 30 years. The product of this plant, a mixed (Ni/Co) sulphide, is shipped to the SMM’s Nickel Refinery in Niihama, Japan. SMM Corporation provides the technology and management personnel to the plant. SMM Corporation also supplies the sulphuric acid, which is the main reagent for the HPAL process."

Jan 1, 2016

By S. Nakamura, K. Fujiwara, S. Hirosawa, S. Tsukamoto, Y. Koshino, Y. Aruga

"In this study, hardness test, differential scanning calorimetry(DSC)and transmission electron micros copy were carried out to investigate the age-hardening behavior of Al-Mg-Si alloy with a Mg+Si concentration of 1.5mass%, and a Mg/Si ratio of 1.0. Pre-aging treatment was conducted for the solution-treated and then water-quenched specimen at 100, 80, 60, 40, 20 (natural aging), 0 and -20°C for 20min or 5h, followed by artificial aging at 170°C for 20min. It was found that when pre-aging time is prolonged from 20min to 5h, the bake-hardening response is deteriorated irrespective of the pre-aging temperature. In particular, the specimens subjected to pre-aging at 20-40°C for 5 h hardly increased the hardness with the subsequent artificial aging. From the results of DSC measurement, it was suggested that as pre-aging temperature is lowered, the amount of solute clusters formed during pre-aging, i.e. Mg-Si clusters and Si-rich clusters, is decreased, but the ratio of Si-rich clusters is increased, resulting in the deteriorated bake-hardening response.INTRODUCTIONIn recent years, Al-Mg-Si alloys have been widely used forautomotive body panels due to their excellent characteristics such as high strength-to-weight ratio and good formability. Their strength is increased not only by artificial aging during paint-bake treatment; i.e. “bake-hardening”, but also when kept at room temperature(RT);i.e. “natural aging”, in association with the suppressed bake-hardening response. Such a “negative effect of two-step aging” has been a serious problem because Al-Mg-Si alloy coils are unavoidably maintained at RT in the manufacturing process of body panels; e.g. shipping and storage. However, it was reported that pre-aging above 70°C before the two-step aging can restore the bake-hardening response, and such an improvement is related to the formation of solute clusters(Sato, 2006; Abidet al., 2010; Takaki et al., 2014; Zandbergen et al., 2015). From the comprehensive analysis by three-dimensional atom probe (Aruga et al., 2014 and 2015), it is believed that Mg-Si clusters formed during pre-aging transform into the strengtheningß"" phase, whereas Si-rich clusters formed during pre-aging act as a detrimental precursor of ß""phase. In this study, it was aimed to investigate the effect of pre-aging conditions, especially pre-aging temperature and time, on the formation of solute clusters during pre-aging, in order to elucidate the mechanism of the age-hardening behavior of an Al-Mg-Si alloy."

Jan 1, 2018

By R. Bruce Ferguson

"The Kettle River Operation of Echo Bay Minerals Company is located about thirty kilometers north of the town of Republic in northeastern Washington State. In tum, Republic is about a hundred and seventy kilometers north and a little west of the city of Spokane, Washington. It is also about fifty kilometers south of the town of Grand Forks, BC, Canada. At this time it is the only gold mining and milling operation in the State.Tailings are stored in a double lined, zero discharge facility. It became evident in mid 1996 that the rate of free water accumulation in the tailings pond was going to result in the level in the pond reaching its permitted freeboard elevation long before the solids plus pore water capacity was reached.We then embarked on a mission to manage tailings pond water on several fronts: reduce fresh water usage in the process, divert spring seeps and surface runoff and install a system to enhance evaporation.IntroductionThe Republic area is an historic gold mining district. Hecla operated its Republic unit from 1892 until closure in 1994. The Kettle River Operation processing plant is located in the San Poi! river valley at an elevation of just under a thousand meters. The facility consists of a ninety thousand ton ore storage pad as the mines are remote and truck haulage is subject to interruption due to weather and road restrictions, jaw and cone crushing with screening, rod and ball milling to an eighty percent passing size of seventy five microns. This is followed by preaeration with high purity oxygen supplied by a vacuum pressure swing adsorption plant plus lead nitrate addition to achieve passivation of the sulfide mineralization such as chalcopyrite, pyrite and pyrrhotite. Conventional cyanide leaching and carbon in pulp adsorption completes the loop. Tailings are treated with air/S02 to achieve cyanide destruction. Plant capacity is 1,815 tons per day. Our license requires that there be no detectable free cyanide in the tailings pond water and a maximum of forty parts per million weak acid dissociable cyanide. The operation started in 1989 with production from the Overlook Mine located about five kilometers further up the same valley at an elevation of twelve hundred meters, hence the location of the mill. Current mine production comes from Lamefoot in the Curlew Lake valley, about ten kilometers distant, and K2 in the Kettle River valley about fifty kilometers away. Both are at an elevation of about 800 meters. Lamefoot produces 1,360 tons per day and K2, 455. Both are underground operations."

Jan 1, 1998

A Joint Venture between Hope Bay Gold Corporation and Miramar Mining Corporation recently acquired the Hope Bay project in Nunavut, which represents the first discovery of a major new Archean greenstone gold belt in Canada since the discovery of Hemlo in 1982 and the Yellowknife belt in 1938. Located some 685 km North of Yellowknife, the Hope Bay Belt is geologically comparable to the other major gold producing regions of Canada. The initial discovery of gold mineralization in 1992 led BHP to acquire virtually all of the Hope Bay belt. Over the subsequent 6 years, BHP expended C$85 million on exploration, including 180,000 meters of drilling, 2.3km of underground workings and a 27,000 tonne bulk sample, as well as extensive environmental and economic studies. Hope Bay Gold and Miramar acquired the Hope Bay property in December 1999 for a cost of C$27 million, considerably less than the C$85 million spent by BHP in exploring the property. They both hold 50% of the project, and share expenses equally. Hope Bay Gold is the manager of all exploration phases and Miramar will manage feasibility, construction and operations. The Hope Bay Joint Venture controls 1,117 sq. km., covering virtually the entire 80 km long Hope Bay and adjacent Elu greenstone belts. We believe that Hope Bay is one of the best-undeveloped gold projects in Canada. Work to date has defined a global resource in excess of three million ounces of gold in three principal deposits, each of which is open to extension along strike and to depth. A C$17 million work program for year 2000 focused on advancing a significant portion of the gold resource to a measured and indicated category. This will provide the basis for initiation of a feasibility study in 2001. The 2000 drill program was one of the largest undertaken in Canada last year, and is a clear indication of our confidence that Hope Bay will develop into a major gold producing district. The majority of the potential mineable areas known to date lie within 350m of surface and would be accessed through a combination of open pit and declines. Currently, development of a 1,800 to 2,000 tpd gold mine producing in excess of 300,000 ounces of gold per year at cash costs below $US200 per ounce is envisioned. In addition, the exploration potential on the 80 km gold belt at Hope Bay is clearly tremendous.

May 1, 2001

By D. Bronkhorst, J. Rheault

"Backfill practices at the Williams mine have evolved significantly since underground mining began in 1986. The use of conveyors as a means of transporting fill material to stopes has been discontinued in favour of load-haul-dump equipment. The method of adding binder to the fill product has been refined along with the placement technique. Slurry preparation has been fully automated and flyash is now included as a binding agent. Surface preparation of the fill product has evolved substantially with the quarry operation coming into full operation and the construction of pennanent crushing, screening, and storage facilities. Company personnel have replaced all contracted labour involved in backfill operations. Some changes have been made to achieve a higher standard of quality control while others were implemented as a matter of cost reduction. BackgroundThe Williams mine is one of three mines operating at the Hernlo gold deposit, located 40 km east of Marathon, Ontario, along the north shore of Lake Superior. Daily production at the mine is 6000 Ud yielding an annual gold production of approximately 500 000 oz. The orebody extends from surface to a depth of 1300 m dipping north at 70 degrees. The average orebody width is 25 m, but varies from 45 m on the eastern boundary to about 10 m at the western probable limit and at depth. Proven and probable reserves currently stand at 32 million tonnes grading 6 g/t.The Williams orebody has been divided into four separate mining zones: the A-zone, the West-of-Dyke zone, the B-zone, and the C-zone (Fig 1). Ore production originally was from the A-zone open pit operation while the A-zone and B-zone underground workings were being developed. The mining method employed underground is blasthole open stoping with delayed rockfill. Mining block dimensions are nominally 20 m along strike by 25 m high. Stope extraction is sequenced to achieve a relatively short pillar life and 100010 ore recovery. Pillars are extracted as soon as the adjacent primary stopes are two cycles ahead. Primary stopes are filled with consolidated rock fill after mining. This allows pillars to be blasted right against the consolidated fill, but also requires that the fill be self supporting to a height of 30 m.The Williams mine places approximately 1.4 million tonnes of rockfill and 50 000 tonnes of binding agents each year, with about three quarters of the fill placed being consolidated. The consolidated fill placed at the Williams mine contains Portland cement and flyash as binding agents."

Jan 1, 1994

By T. C. Chudy, C. C. Hamilton, G. R. Lane

The selection and provision of mineralogical services to assist in developing a GeoMetallurgical program for a Cu Porphyry Mine in South America is covered in this paper and presentation. The objectives of the study were to provide a rapid mineralogical scoping protocol to support and confirm results of sequential Cu leach data and to provide detailed mineralogical analysis of four composites and products on which locked cycle test-work were performed. Results showed good reconciliation was achieved between measured chalcopyrite and secondary Cu mineral contents with leach data using a linear intercept mineralogical procedure by TIMA (Tescan Integrated Mineral Analyser), as well as a generally positive correlation between pyrite and Cu-sulphide grain size. Size-by-size analysis of Composite feeds and test products was used to outline comparisons between feeds on a mineralogical basis and provide liberation and association data with which to diagnose and better explain performance. Identifying key mineralogical drivers to Cu recovery and concentrate grades using these results, a greater understanding of the role of chalcocite and pyrite on metallurgical performance was established. INTRODUCTION As a global supply of copper, porphyry copper deposits are typically very large, often in excess of a billion tons of ore (Keeney, Walters, & Kojovic, 2011; McMillan & Panteleyev, 1980) with grades typically between 0.2 and 1.2% Cu, and up to 2% Cu in places. They are also a source of gold. The complexity of copper porphyries arises from their classically zoned nature, the type of host rock, as well as whether or not they have been overprinted by oxidation, weathering and supergene enrichment (Sillitoe, 2011). Zones of alteration and other geological attributes give rise to varied mineralogy, ore mineral textures and material properties, all of which result in different processing characteristics. Due to the interdependencies of geological and processing characteristics, it is vital to study links between the geology at the macro-scale and mineralogical features down to millimeter and micrometer scale (mineral content and texture) and their influences on mineral processing behaviour (Cropp, Goodall, & Bradshaw, 2013). The growing field of GeoMetallurgy aims to provide an Orebody Knowledge framework by which to gather such information as it pertains to: (1) Understanding and predicting likely flotation response; (2) Guiding relevant test work to investigate and validate links, interdependencies, theories and data interpretation/evaluation; (3) Populate a block model with relevant information and; (4) Implement and refine the model for ongoing optimization and control. Specific questions addressed by using mineralogy often include flotation diagnostics as aids to interpretation, causes of dilution and losses to tailings, as well as guidance to sampling for test work (Cropp et al., 2013).

Jan 1, 2019

By S. E. Malouf, R. Hinse

"The Mineral possibilities of the Chibougamau area were first recognized by Richardson in 1870 and Low in 1885, both of the Geological Survey of Canada. The first record of interest by the Quebec Bureau of Mines is in 1904, when J. Obalski reported on the discovery of copper and asbestos by Peter McKenzie in the previous year. Reports of 1906 and 1907 covered the McKenzie Gold Mine find and other deposits. In 1908 Prof. Dulieux described the quartz vein on Merrill Island and the mineralized zone on the near-by mainland. In 1910 the Chibougamau Railway Commission, though commenting favourably on the possibilities of the area, concluded that a railway was not then warranted.In spite of this, prospecting in the district continued. The main Campbell Chibougamau deposit, the Blake showing on Merrill Island in Dore Lake, was found in 1920, and other discoveries were made soon after. Between 1927 the work of Mawdsley, Retty, and Norman tor the Geological Survey and the Quebec Bureau of Mines stimulated further interest. In 1934 Consolidated Chibougamau Goldfields Ltd., a subsidiary of Consolidated Mining and Smelting Co. of Canada Ltd., acquired the Merrill Island and Cedar Bay holdings of Chibougamau Prospectors Ltd. They embarked on an aggressive campaign, directed essentially at the gold quartz veins at Cedar Bay. Exploration there included a 522-foot shaft, 4,732 feet of drifts and crosscuts, and about 34,000 feet of surface and underground diamond drilling. Thirteen holes were also drilled on the replacement sulphide lens at Merrill Island. Operations were suspended in 1936, as the occurrence was too low grade for development at the existing metal prices and under difficult operating conditions.The completion of the Quebec Bureau of Mines highway from St. Felicien to Chibougamau in 1950 was followed by a project of detailed geological mapping by the Bureau. It also encouraged Campbell Chibou-gamau Mines to acquire the assets of Consolidated Chibougamau Gold-fields and commence a program of surface exploration. Between August 1950 and April 1951, 41,183 feet of diamond drilling was done on the Cedar Bay and Merrill Island holdings. The Blake showing off the west shore of Merrill Island developed at depth into a major orebody. A lease was obtained on adjacent ground from Merrill Island Corporation, and underground work commenced in November 1951. Development at Merrill Island to June 1st, 1955, included a four-compartment shaft to 1,225 feet, with 8 levels at 150-foot intervals, 5,317 feet of raises, 16,709 feet of drifts and crosscuts, 52,514 feet of surface and 96,593 feet of underground diamond drilling."

Jan 1, 1949