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By H. Sahay

"DISCLAIMER NOTICEThe views and conclusions contained in this presentation are those of the author and should not be interpreted as necessarily representing the official policies or recommendations of the Department of Mines and Energy, Government of Newfoundland and Labrador. ABSTRACTThis paper describes the main testing procedures that should be carried out to evaluate the capacity and capability of the service brakes on the large-size diesel-electric trucks now in common use at open-pit mining operations. Brake testing programs in general are described, and then consideration is given to different techniques of service brake testing, brake performance evaluation, braking energy and typical performance characteristics.IntroductionThe first diesel-electric truck with a capacity of about 85 tons was introduced in 1963 by Unit Rig and Equipment Company of Tulsa in cooperation with General Electric Company, who successfully developed and supplied the most important component- motorized wheels. Since then, its popularity and numbers have continued to grow, and now more than 3700 such units are in operation in mines around the world. In Canada alone, their number is more than 700, including about 150 in the Province of Newfoundland and Labrador. Similarly, their size has also grown from that initial capacity of 85 tons to 350 tons.Although the introduction of the electric trucks proved to be invariably economical, their indiscriminate use was not without severe consequences. Many runaways and even fatalities occurred when the dynamic retarder failed while the truck was travelling loaded down severe grades. The service brakes in such cases proved incapable of stopping runaway trucks, mainly due to their limited capability and the fading caused by excessive heat produced during the act of braking itself. This eventually led the various governments, the open-pit operators and the manufacturers to seriously question the braking capability as well as function of both service brakes and dynamic retarders. Extensive testing"

Jan 1, 1980

By G. K. Livingstone

"This paper describes mining conditions in the Kaiser surface and underground coal mines in the Sparwood area of southeastern British Columbia. Mining methods are explained, major mining equipment is listed and production rates are given. Reclamation methods and results are described. Data are given on safety organization, safety practices, accident statistics, safety inspections and absenteeism. IntroductionThe coal mines of Kaiser Resources Ltd. are located in the southeast corner of British Columbia. All present mining is in the Sparwood area, where operations began in 1900 with the opening of the Michel underground mines. Production increased rapidly at first, then went through successive periods of depression and prosperity. Following are the highlights. Quantities are in clean, short tons. 1900 1910-Rapid rise in production to a peak of 512,000 tons in 1910.1911-1924-Reduced rate of production, averaging 238,000 tons per year, due to strikes, manpower shortage during World War 1 and post-war depression.1925-1929-Second prosperous era, with production climbing to 383,000 in 1929.1930-1933-Second depression-production falling to 231,000 in 1933.1934-1942-Third era of prosperity-production recovered and new production peak of 908,000 tons reached in 1942 during World War II.1943-1956-Corttinued prosperity, with production averaging 720,000 tons per year.1957-1959-Loss of markets to oil and gas; production dropping to 517,000 tons in 1959.1959-1969-Recovery due to new coal market in Japan; production climbing back to the 900,000-ton level.197D-Present-Jump in production reaching 6.6 million tons in 1975. This was the result of the purchase of the property by Kaiser Steel, who negotiated a large-volume, long-term contract with Japan, built the coal port of Westshore, reduced shipping costs by instituting unit trains, and reduced production costs by investing in large-scale mining machinery and a new preparation plant."

Jan 1, 1980

By Earl Gebhardt

"The planning and design of mining panels in the Potash Corporation of Saskatchewan Inc., Lanigan Division mine Block B involved a number of investigative techniques. The techniques utilized to gain insight into the viability of mining included2-D and 3-D seismic surveys, longhole drilling, exploration wells, photo lineament, magnetic and gravity surveys, and actual room excavation.IntroductionPotash Corporation of Saskatchewan Inc. (PCS) owns and operates four potash mines in the province of Saskatchewan. The Lanigan Division mine is located 110 Ian east of Saskatoon, Saskatchewan, near the town of Lanigan.The Lanigan Mine commenced production in August 1968under the ownership or Alwinsal Potash of Canada Limited. The mine was purchased by PCS in 1977.Lanigan Division's productive capacity is 3.4 million tonnes of product per year. Ore is extracted from the "" B"" zone or Layer III or the Patience Lake member of the Prairie Evaporite, more than 1000m below surface . The Patience Lake member (approximately 20 m thick) consists of potash interbedded with salt. The average run-of-mine ore grade is 21070 K20 (Fig. I).The mine was initially developed using the room-and-pillar technique. Improvements in mining resulted in the use or extensible conveyors in long panel rooms."

Jan 1, 1993

By H. N. Tran, D. Barham

"A systematic survey of the appropriate literature in terms of phase equilibria and possible reactions involving deposits in Municipal Solid Waste (MSW) incinerators are analyzed particularly with respect to the theoretical first melting points of typical ashes and the possible presence of a liquid phase at the temperature of ash-metal interface. Conclusions are drawn on the basis of the possible presence or absence of various metal chlorides and sulphates in such situations, and their effects on the problems of fouling and metal corrosion. From preliminary results, it would appear that ZnCl2 is unlikely to be present in any molten phase which causes corrosion because it evaporates rapidly from such liquids. It would also appear that PbCl2 and, by inference, ZnCl2 will sulphate at relatively low temperatures if the flue gas contains SO2 and O2.IntroductionMunicipal Solid Waste (MSW) incinerators produce fly ash which deposits on heat transfer surfaces throughout the furnace, leading to corrosion and loss of energy recovery efficiency. Despite a massive amount of research carried out on the subject of corrosion in MSW incinerators, the problem of corrosion of heat transfer metal is still with us, and no definite over-all cause and effect relationship has yet been identified. In order to make a contribution to the eventual solution of this problem, the authors have examined the problem from the beginning, using their knowledge of fouling and corrosion problems in black liquor recovery boilers as a starting point.The incineration of MSW has been used for many years to reduce the volume of material sent to landfill (ash volume compared to raw waste volume), and, in most cases, to recover energy at the same time by raising steam The steam may be ""low- or high pressure/ temperature"" depending on its end use. The higher the steam pressure and temperature, the greater seem to be the problems with corrosion of heat exchanger metal, presumably because the metal must withstand higher surface temperatures.The chemical and physical composition of MSW is extremely variable, not only on a regional basis, seasonal basis, and day-today basis, but also on a minute-by-minute basis, due to the inhomogeneity of MSW. This makes it extremely difficult to define cause and effect relationships, because both cause and effect are transient in nature."

Jan 1, 1991

By Arnold R. Pasieka, George V. Sotirow

"The object of this paper is to outline a computerized net cash flow approach for practical cutoff grade determinations. A cutoff grade carries not only geological but also economic value, in which costs are balanced against revenues. For different situations, cutoff grades are a 'standard of value' that are determined for the purpose of classifying economic or uneconomic mineral reserves in a specified time frame. In a constantly changing technical and economic environment, the decision to develop or maintain a mining operation at its best economic potential, requires an appropriate cutoff grade strategy, a strategy consistent with short- and long-term operational and corporate objectives and one based on realistic marketing and metal prices forecasts. The use of a project evaluation computer program, in combination with specific corporate net cash flow requirements, shows the most promising and flexible team approach for cutoff grade optimization.IntroductionThe recent recession has had a devastating effect on the Canadian mining industry. Exploration and mining R&D expenditures have been critically reduced and now are believed to be below healthy levels. Ore reserves, mine life and cutoff grades are careening in several directions because of low base metal prices, relatively high costs and sensitive precious metal prices.These conditions demand an improved understanding of the terms ""ore"" and ""cutoff grades"" as they are applied to the planning process and day-to-day operations. ""Ore"" or 'ore reserves"" is a dynamic concept, reflecting a cumulative effect of changes in geological, technical and economic conditions in terms of material which can be mined at a potential profit. By contrast, if economics are not the primary consideration, the term ""mineral reserve"" or ""mineral inventory"" should be used to describe the mineral deposit. For different situations, cutoff grade is ""a standard of value"" determined for the purpose of classifying mineral reserves as economic or uneconomic in a specified time frame. The question of mining at average reserve grade or low-grade material during high metal prices is of paramount importance for cutoff grade strategy."

Jan 1, 1985

By Douglas C. Fraser

"Geophysical data are often presented on maps, sections or profiles with minimal data processing. A common example is magnetic data, in which the diurnal-free raw readings are merely contoured on a map. Other examples include profile presentation of VLF-EM data, IP pseudo-sections and airborne EM anomalies which may be shown on a map with only the measured amplitudes indicated.In recent years, efforts have been directed to the portraying of data in more useful forms. A number of algorithms (i.e., manipulative procedures and equations) have proven to contribute significantly to the usefulness of geophysical data. Some can be applied manually; others are sufficiently complicated to require a computer. Some merely alter the data to a more useful geometric form without altering the units. Others yield quantitative estimates of source body parameters such as conductivity and shape. Several examples follow:1. A simple algorithm is in common use which transforms noisy non-contourable VLF-EM data into less noisy contourable data. The input is the dip angle in degrees or per cent and the output is also in these same units.2. A simple algorithm has recently been developed which allows multi-level dipole-dipole IP data to be contoured on a map. Again, the units are not altered.3. A relatively complicated algorithm is in common use which transforms non-contourable height-sensitive helicopter EM data into contourable resistivity data,4. An algorithm has been recently developed which transforms non-contourable height-sensitive EM data into contourable weight per cent magnetite.The purpose of all the above techniques is to present the data in a simplified form. This, in turn, allows an easier assimilation by the user of much of the information content of the data."

Jan 1, 1981

By D. E. Macdonald

"Approximately 2.0 million tonnes of phosphate rock is imported from the U.S. to Alberta annually and is used in the production of agricultural fertilizers. No phosphate rock is presently mined in Canada, although sources of igneous phosphate have been mined in the past. Sedimentary phosphate rock is found in nine geological units within the Cordilleran Foothills and Front Ranges region of Alberta. The Devonian-Mississippian Exshaw Formation, Permo-Pennsylvanian Rocky Mountain Supergroup, and the Jurassic Fernie Formation have the best potential f or future development. Most of the phosphate rock in Alberta is of low grade and lies within restricted land use areas. The best over-all resource potential lies in the Fernie basin area of southeastern British Columbia, in the Fernie Formation. The possibility of deep phosphate rock deposits in the plains region should be examined.Phosphate rock is mined in Idaho, Utah, Montana and Florida, and is shipped to Canadian fertilizer plants for processing. Raw phosphate rock is acidified, usually with sulphuric acid, by a wet process. Varying amounts of ammonia are added, and the fertilizer is pelletized. Future research, in western Canada, is needed in the beneficiating of low-grade ores and in the area of in situ mining. At the present time, phosphate rock sells for around US$50Itonne, and Alberta fertilizer companies probably pay in the range of C$7D-$85Itonne for U.S. phosphate rock delivered to the plant. Environmental concerns about phosphate rock processing are centred around gypsum tailing ponds, and radioactive wastes.U.S. supplies of phosphate rock are secure until the year 2000; however, by this time, the richer deposits in the U.S. may be exhausted. Alternate deposits, such as the ones in Alberta and British Columbia, might then be considered. Phosphate byproducts, such as uranium, vanadium, fluorine and rare earth elements, may help in making the processing of low-grade Canadian phosphates economically feasible."

Jan 1, 1988

By E. Kouame, M. D. Everell, D. Hodouin, T. Kasongo

"To properly operate a mineral beneficiation plant it is often necessary to have systematic evaluations of the performances of a piece of machinery, of one of the processes or of the plant as a whole. The state of a process can be described from a collection of data: flow-rate values, particle-size distributions and chemical compositions of the various streams of the circuit studied.However, the obtained raw data are generally redundant and inconsistent from a material balance viewpoint, due to natural disturbances, sampling errors, unreliable instrument readouts and laboratory analysis inaccuracies.BILMAT is a computer program which upgrades mineral processing data, adjusting them in a least-squares sense in order that they may verify material conservation constraints. As a consequence, the program is able to calculate the best values of unmeasured flow rates (such as. circulating load ratios). BILMAT is a very general program in the sense that it can accept various data from any type of mineral processing circuit, and its usefulness is demonstrated here for three processing circuits of a pyrochlore concentration plant: grinding, desliming and flotation.IntroductionComputation of the material balance around any metallurgical process, that is the determination of stream flow rates and composition. is a basic step in its evaluation. In spite of the simplicity of the physical laws of mass conservation. and of the very fundamental need to know the flow of components around a process. it appears, particularly in mineral processing. that material balancing is quite often a difficult task. The reasons for this involve the problems associated with collecting measurements in mineral dressing plants. They are generally poorly instrumented and invariably data must be collected through hand sampling. Due to the heterogeneous nature of the streams, segregation occurs and samples are often not adequately representative, a problem made worse by the lack of circuit sampling facilities, which are often completely ignored by plant designers. Furthermore, the nature of the measurements to be performed (for instance particle size distributions in the fine range) frequently gives inaccurate results."

Jan 1, 1981

By Jack A. Caldwell

"This paper describes the nature of stage curves and the ways in which they have been used to study the construction sequencing of tailings impoundments. They provide a simple graphical method of representing the main aspects of an impoundment, studying the effects of different impoundment geometries, and predicting and controlling impoundment operations.IntroductionThe designer o f a tailings impoundment must, at all stages of design, examine the relations hip between the volume o r tones of tailings that will be deposited at the impoundment, the height to which the tailings will rise in a given time and that rate at which the various parts of an impoundment will be constructed. The operator, if he is to keep proper control of this impoundment, must be able to examine the effect of different deposit ion rates at the impoundment, and the consequences of placing more or less tailings at different parts of the impoundment. The operator must be able to see at a glance just how much higher an impoundment will be at the end o f the following yea r of deposition, and this he must be able to do in order to formulate his annual construction and operating plans.Stage curves are the best way o f modeling an impoundment. The simple impoundment model represented by the stage curves enables the designer and operator visually to examine the relationships between the various factors, such as impoundment volume, deposition rat e and construction rate, that dictate the essential features o f impoundment design and operation.A design model for a tailings impoundment may be formulated as in Figure 1. This involves the site selection process, the definition o f a geotechnical model, a water balance model and a height-capacity-production model."

Jan 1, 1983

By Carl Edmunds

Talk Outline Deposit?s Vital Statistics and Economic Environment Regional Geologic Setting Deposit Geology UBC Research Search Model and Current Exploration Plans Kemess Operations Kemess South Mine Mining rate (2006) ?135,000 tonnes / day Mill throughput (2006) ?52,000 tonnes / day Average Annual Production ?300,000 oz gold and 75 million lbs copper ?Production until mid-2009 Kemess North > Open pit reserve 423.8 million tonnes @ 0.3 g/tgold and 0.18% copper > 4.1 million ounce gold reserve (127 tonnes) > 1.45 billion pound copper reserve (636,000 tonnes) > 0.75:1 Strip Ratio

May 1, 2006

By P. A. Geusebroek, N. A. Duke

Abstract - The Lupin mine, located in the central Slave province just east of the western boundary of Nunavut Territory, is a world-class example of a Neoarchean-aged banded iron formation (BIF)-hosted lode-gold deposit. At the minesite the gold-mineralized Lupin BIF, separating stratigraphically underlying psammitic wacke and overlying argillaceous turbidite sequences, delineates the Lupin dome, a hammerhead-shaped F2/F3 interference fold structure occurring at the greenschist to amphibolite facies metamorphic transition within the thermal aureole of the Contwoyto batholith. Detailed paragenetic relationships indicate that peak thermal metamorphism coincided with the switch from regional D2 compression to rapid D3 unroofing of the Neoarchean orogenic infrastructure. Gold initially precipitated with pyrrhotite, replacing amphibolitic BIF at the apex of the Lupin deformation zone, separating the east and west lobes of the Contwoyto batholith. Over the course of associated prograde/retrograde metasomatic overprints, gold was further remobilized during garnet and loellingite/arsenopyrite growth in chlorite-altered selvages of late-forming ladder quartz veins. A metamorphic model of ore genesis, with gold being scavenged and transported by metamorphic fluid that was shed and structurally trapped at the amphibolite recrystallization front, is favored over the previously proposed syngenetic and exogenic models of gold concentration that have tended to polarize genetic interpretations to date.

Jan 1, 2004

By L. J. Bryson, C. A. Biley

"Anglo American Platinum’s Rustenburg Base Metals Refiners (RBMR) produces Ni and Cu cathode from slow cooled converter matte after Platinum Group Metals (PGMs) are magnetically separated for further processing. In 2010, execution of the RBMR expansion project was commenced to increase the production capacity to 33 ktpa Ni. The expansion, which is currently still in a ramp-up stage, brought about several fundamental changes to the circuit chemistry, including the installation of a dedicated Pressure Iron Removal (PIR) autoclave. This provided an independent outlet for iron from the refinery, which previously deported to the residue from the leach circuit and was toll refined. At the same time as changes in the circuit associated with the RBMR expansion were being made, the plant experienced a significant increase in iron and copper units per nickel unit in the feed. The decoupling of the leaching and iron removal process steps has had significant advantages for the robustness and control of the overall circuit, in particular pertaining to the preparation of a nickel electrolyte of sufficient purity to produce LME grade nickel. This paper details the primary changes to the RBMR circuit chemistry with a particular focus on iron deportment and the benefits derived from the installation of the PIR autoclave.INTRODUCTIONAnglo American Platinum mines and refines Platinum Group Metals (PGMs) through a complex processing pipeline in its South African operations. Mined ore is first treated through several mineral processing stages, geared to the recovery of base metals and associated PGMs, followed by smelting and refining operations that aim to separate base metals from PGMs and recover the value components to purified, saleable products. The focus of this paper is the Rustenburg Base Metals Refiners (RBMR) complex which receives a PGM-containing converter matte, then physically separates PGMs and base metals, and recovers nickel and copper to cathode by utilising leaching, precipitation and electrowinning processes.With new mining operations coming online and the continued increase in PGM ore arisings, it was necessary to undertake an expansion project at the Base Metal Refinery (BMR) to increase production capacity from 21 ktpa to 33 ktpa Ni. Complicating the expansion project was the requirement of continued production through all stages of design, construction and implementation and, due to stringent capital constraints, the utilisation of all existing equipment installed during a previous expansion in 1981. Against this backdrop, the expanded BMR flow sheet involved significant changes to the chemistry, especially iron management. These changes in iron management allowed for the separation of iron and copper residues and significantly increased the robustness of the plant to changes in feed mineralogy as well as the potential to increase production duties of the existing equipment."

Jan 1, 2016

By T. W. Sullivan

"This paper explores the history of AC static variable frequency control drive systems on open cast mining machinery, in particular electric mining shovels and electric walking draglines, over the past ten years. It critiques the peformance of these drive systems from a productivity and operating cost standpoint and evaluates the over-all success of the original design concepts as they were envisaged in the 1970s. The paper details the state-of-the-art as it exists today as well as design improvements contemplated for the immediate future.IntroductionConsidering the fact that the very first successful application of an alternating current electrical distribution system occurred at the Gold King gold mine near Ames, Colorado, U.S.A., in the year 1891, it seems somewhat ironic that almost 100 years later, AC excavator drive systems are just now becoming commonplace in the mining industry.If one were to read the biography of Lucien L. Nunn, the first person to successfully install a commercially viable alternating current power distribution system, one would quickly visualize a parallelism between his fight to achieve acceptance of his new lower cost technology and that of those commissioned with the belief that AC static variable frequency control drive systems are not only a better performing, lower cost drive system for mining excavators today, but moreover, the drive systems of the future in our industry.This paper will deal with the evolution and incorporation of AC static drive systems on open cast mining excavators. It will trace the history of the development of these drive systems in the mining industry, the current state-of-the-art concerning these systems, and what the future holds for these excavators in our industry."

Jan 1, 1990

By D. A. Nelson, Wedding. R. E.

"A planned production increase of 1.25 million tonnes per year at Domtar's salt mine in Goderich, Ontario required an increase in the underground mine ventilation system from 320,000 to 700,000 scfm, The design of the expanded facility included the addition of a 22-foot-diameter shaft to augment the existing production and service shafts.A preliminary review of the ventilation requirement s indicated that in order to pr ovide good economic design for the expanded system, a comprehensive ventilation survey of the existing system was required. The survey was completed in the latter part of 1980, and subsequent engineering design and equipment selection proceeded throughout 1981.This paper presents the methods used to carry out the survey. Particular attention is given to the numerous challenges presented in determining air flows in the large 45-ftwide by 43-ft-high airways. Emphasis is placed on practical field survey techniques.A mathematical model of the ventilation system using a computer program is described, and the application of this program to provide design data and assist in equipment selection is discussed. Economic sizing of airways and fan selection are analyzed considering horsepower and operating costs.IntroductionThe Goderich Mine of Sifto Salt Division, Domtar Inc., is located on the eastern shore of Lake Huron in the town of Goderich, Ontario, and has been operating since November 1959. The salt is extracted by a modified room-and-pillar method, and the workings advance at a rate of 1.5 miles per year by undercutting, drilling and blasting. The surface mine area consists of a 12-acre site in the Goderich harbour, at the mouth of the Maitland River, and the underground property consists of approximately 2,700 acres lying mainly under Lake Huron."

Jan 1, 1983

By Y. Wang, J. D. Miller, C. L. Lin, Virginia Tech

Conventional flotation machines are typically limited to a particle topsize of 150-200 microns due to inherent constraints created by the pulp and froth phases. To overcome these limitations, a novel fluidized-bed flotation system called the HydroFloatTM Separator has been developed specifically for the purpose of floating coarse particles containing only small amounts of exposed hydrophobic minerals. Over the last decade, this technology has been successfully applied to industrial minerals with several full-scale units installed to recover particles up to and exceeding 3,000 microns diameter. More recently, sulfidebased test work has shown that this novel device is also capable of recovering metalliferous values at a grind size that is much coarser than currently used in industrial concentrators. In the current study, high resolution X-ray microtomography (HRXMT) was used to experimentally compare the degree of exposed grain surface area necessary to recover coarse particles using the HydroFloatTM technology to that attainable using a traditional mechanical flotation cell. The data indicate that exposed grain surface area is a critical factor for coarse particle flotation. For the gold-bearing sulfide ore examined in this study, the HRXMT data suggest that near complete recoveries of coarse (850x500 µm) multiphase particles containing as little as 1% exposed grain surface area were realistically attainable using the HydroFloatTM technology. As such, this new technology may offer a unique opportunity for increasing concentrator capacity by increasing the primary grind size needed for rougher/scavenger separations.

Jan 1, 2016

By E. Rosseland

Glencore Nikkelverk is operating its Chlorine Leach Process in Kristiansand, Norway, producing nickel, copper and cobalt metals by electrowinning. The copper tankhouse, with an annual capacity of 40 000 tons, is based on traditional technology with copper starting sheets and cast lead anodes. Due to the integration of this copper sulphate process in a nickel chloride refinery, operating conditions are unique. In recent years, operating costs have escalated, and health, safety and environment are top priorities. Test work on a new copper process involving permanent cathodes was therefore initiated in 2007, and a pilot plant was started up in 2012. The objective was to introduce innovative technologies and tailor these to be robust in the Nikkelverk process, to form a basis for a new sustainable copper tankhouse with a high degree of automation. Key learnings from more than 10 years of R&D work is presented, focusing on cathode materials and copper metal deposition including product quality, robotic stripping, coated titanium anodes, electrode contacts and alignment, electrode current monitoring with Hatch HELM tracker technology and acid mist capture using cell hoods with off-gas scrubbing. The new copper tankhouse project was approved in December 2018, and commissioning is planned in 2022. INTRODUCTION The Kristiansand nickel refinery was started up in 1910, processing nickel-copper matte originating from Norwegian mines. It was the first industrial plant in the world to produce nickel by electrolysis using the Hybinette electrorefining process. After roasting to burn off sulphur in the matte, copper was leached and recovered as cathodes by electrowinning with lead anodes. In the first years, the copper cathodes were fire- refined and cast into ingots for sale (Thonstad Sandvik, 2004). The Nikkelverk refinery was acquired by Falconbridge Ltd. in 1929, and since then matte from Sudbury, Canada has been the main feed source. The Hybinette process was in operation until 1975 when conversion to the newly developed Chlorine Leach Process was started (Stensholt, Zachariasen, & Lund, 1986). In this process still used today, nickel and part of the copper in the matte is leached using chlorine gas generated on dimensionally stable anodes in the nickel and cobalt electrowinning tankhouses, as illustrated in Figure 1. The presence of dissolved copper ions in the strong chloride leach solution is essential for effective chlorine absorption and nickel dissolution. To improve copper-nickel separation, slurry from the atmospheric leach is pumped to autoclaves and then to cementation tanks for precipitation of copper as copper sulphide, which is collected in filter presses. The leach residue, also containing the precious metals, goes through a 3-stage wash to remove chlorides before being introduced as a slurry to fluid-bed roasters for sulphur removal. The resulting calcine is then leached in spent electrolyte from the copper electrowinning tankhouse, to dissolve most of the copper and also some nickel, cobalt and other elements. After separation of the leach residue, elements like Bi, Te, As and Ag are removed from the strong copper sulphate solution by reaction with metallic copper in three columns in series filled with cut copper cathodes. The purified solution then passes polishing filters before being mixed with spent electrolyte and returned to the electrowinning cells.

Jan 1, 2019

By K. M. Deen, E. Asselin

Hydrometallurgical processes, i.e., Cu extraction from a mineral concentrate and Zn electrowinning, are coupled in a battery-like system. A naturally-sourced CuFeS2 concentrate mixed with activated carbon is used as a positive slurry electrode separated by a membrane from the negative compartment. The Zn is deposited on the negative Al current collector from the circulating solution containing 100 g L–1 Zn2+ in 0.2 M H2SO4. In this battery-like system, both CuFeS2 oxidation and Zn deposition are possible during the charging cycle. The additional benefit of this setup is the energy storage if discharged by re-dissolving the Zn in the negative compartment. Cyclic charge/discharge testing revealed oxidation of the mineral concentrate supported by the Zn2+/Zn redox reaction. The coulombic and energy efficiencies increased monotonically to 95.6% and 43%, respectively, with approximately 7% Cu extraction during the first 10 charge/discharge cycles (completed in 2.4 hours). The discharge energy density also increased from 2.7 to 36.2 mWh L–1. In the following 90 cycles, the low energy (~14%) and coulombic (~33%) efficiencies that were obtained confirmed the progress of irreversible reactions e.g. Cu extraction from the mineral concentrate. Overall, approximately 16% Cu was extracted from the mineral concentrate over 100 charge/discharge cycles (or about 12.2 hours). INTRODUCTION Chalcopyrite (CuFeS2) is the main economic copper-bearing mineral, which is generally processed through pyro-metallurgical methods to produce Cu. However, environmental challenges with smelting, depletion of high-grade ores and a large number of impurities (some toxic) have resulted in continued research aimed at the development of hydrometallurgical technologies to process CuFeS2. There are a number of different proposed hydrometallurgical routes in the literature to treat CuFeS2 concentrates, but only a few have attained commercial acceptance (Dreisinger, 2006). Zn is another commercially important metal that is mainly produced via acid leaching of roasted sphalerite (ZnS). The pregnant leach solution is rigorously purified to remove all the soluble metal impurities, particularly those that are electrochemically more noble than Zn and present large kinetic activity towards H2 evolution (Hu & Piron, 1992; Mackinnon, Brannen, & Fenn, 1987; Mureşan, Maurin, Oniciu, & Gaga, 1996). Metal impurities such as Cu, Co, Ni, Cd, Pb, Sb, As, Ge and Fe etc., if they exist in the leach liquor beyond the tolerance limits, can decrease the current efficiency during electrowinning and deteriorate the deposit quality (Alfantazi & Dreisinger, 2001). A complete flow sheet, describing the process steps of Zn metal production from mine site to pure metal is presented elsewhere (Brown, Meisenhelder, & Yao, 1983).

Jan 1, 2019

By D. Milne, P. Germain, G. Sauriol

"The practical determination of optimum stope geometry has, until recently, been an art relying primarily on experience at a given operation. Due to the complex and variable nature of rock masses in situ, stopes are often designed either very conservatively, resulting in excess development, or non-conservative resulting in excessive dilution and, in some cases, in loss of ore reserves.The Geomechanics group of the Mining Technology Division, Centre de Recherche Noranda, uses a combination numerical modelling and empirical design techniques to optimize allowable stope geometries. The above methodology allows the optim ization of stope geometry for varying rockmass classifications within an operation or between different operations. Such techniques require the development of a thorough geotechnical database at each operation and are best used when calibrated to the specific operation through back analysis.This paper discusses the above optimization techniques in light of specific examples from the Noranda Group. The design methods, limitations and results are discussed.INTRODUCTIONMine design has traditionally referred to production planning, general facility layout, (i.e. ore passes, crusher stations, main haulage drifts, etc.) and ventilation design. The dimensioning and sequencing of the stopes and pillars, however, has been almost solely based on trial and error and precedent practice.More recently, numerical stress analysis has received much more attention as part of the design process. It is common now at many operations to look at the effect of proposed stope layouts and sequences on stress redistribution through the surrounding mine workings. Through calibration of actual stope and pillar behaviour in situ, one can then relate the model predictions to support requirements, safety problems, etc."

Jan 1, 1989

By R. W. Zeindler, D. A. Fawcett

"A summary is presented of the report completed by the Coal Mining Research Centre in 1980. This report was of a cross-pit conveyor system applied as an auxiliary mining complement to the major stripping unit, a dragline. The purpose of the CMRC report was to evaluate selective mining and replacement of the upper horizons of the strata as an aid in reclamation. These strata were the topsoil, subsoil and glacial till.Past utilization of cross-pit conveyor systems and related engineering studies were assessed. The parameters of the study were based on the mining and geological conditions of the Alberta prairie coal mines. The principal excavator for stripping was a dragline. Excavation of the upper horizons was done by a bucket-wheel excavator discharging onto the crosspit conveyor. Alternative equipment applications were economically compared. Four cases or geological sections were evaluated in detail. The economics of the alternative mining systems for each of the cases were determined. In all instances, the most economical solution was a tandem system utilizing a dragline with a bucket-wheel excavator/ cross-pit conveyor system."

Jan 1, 1981

By V. J. Hucka, J. H. Tu

"A computer model has been developed for evaluating the performance of truck and shovel operations in an open-pit mine. The data required for describing the haulage network are simplified and shift startup and ending procedures specific to any given mine can be incorporated into the model readily . The model uses a FORTRAN-based simulation language SLAM and can be tailored to the conditions and practices of a particular mine.The model is used to analyze truck and shovel operations, evaluate different haulage layouts in the planning stage, or compare the expected performance of alternative truck dispatching systems.Stochastic simulation is used to model the random variation in time required f or the individual activities. This gives a more realistic representation of the processes under study. Simulation runs show good agreement with actual performance. The model is especially useful for examining alternatives in equipment acquisition , haulroad design, truck-shovel allocation and dispatching methods. The results of such simulations provide a sound basis f or making major management decisions.IntroductionTruck haul age is widely used in open-pit mining operations. It is also the largest cost item in the operating costs of an openpit mine, in some cases accounting for an estimated 50% of operating costs(1) , Truck replacements, which are necessary from time to time, involve large amounts of cap ital. Increasing depth of mining and consequent increase in stripping ratios require additional capital investment in the truck-shovel fleet. The sharp rise in energy costs over the last decade has also contibuted significantly to an increase in truck fleet operating costs . All these factors have led mine operators to search for ways to improve the effectiveness of the truck-shovel fleet, in order to lower costs and maintain a profitable operation in the face of declining markets and increasing world wide competition.Control of truck haulage in an open-pit mine is usually done in one of two ways."

Jan 1, 1985