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By Robert B. Parsons

"It is generally acknowledged that Canadian and foreign-based mining companies are spending relatively fewer dollars in Canada and more dollars abroad than was the case not many years ago. It is common to see the Canadian tax system being fingered as one of the causes in this shift in direction. However, if one takes the time to peel away the surface of our tax system, and to probe a little deeper into its substance, it will be found that the Canadian taxation of mining is generally quite competitive. IntroductionCross-border shopping in the mining industry? - Absolutely! Economics 100 teaches that dollars are votes in the capitalist system, and mining companies are going to cast their limited number of votes where they can achieve the highest return at the least risk. The Canadian consumer knows that taxes are a major factor in cross-border shopping, and mining company executives know that the same tenet holds true in their industry. Unfortunately, mining operations are subject to such a bewildering array of tax rules that making an intelligent decision is not always an easy matter.Statutory Tax RatesConsider, for example, the ""statutory"" tax rates shown in Table 1. (Statutory tax rate in this table refers to the simple sum of federal and provincial tax rates on profits as specified in the tax legislation.) Looking only at the total statutory tax rates (such as 66070 for Manitoba) , one might quickly observe that the Canadian mining industry is subject to a truly onerous tax regime, especially as compared to the United States where statutory tax rates can be in the 34% range."

Jan 1, 1992

By M. Mahdi Aboutalebi, Roderick I. L. Guthrie, Mihaiela Isac

"A preliminary computational fluid flow model has been developed to simulate the flow field of liquid steel in the mold region of billet mold casting. This study reports the effect of different Submerged Entry Nozzles (SEN) on the fluid flow in round and square billet, curved mold, casters. In this study, ANSYS Fluent 14.5 software was used to model the 3-D turbulent flow of liquid steel. The standard k-0 model was used to simulate turbulence in the mold region. Similarly, the species transport equations were discretized to simulate the dye injection test into the mold cavity. The behavior of flow in the mold region was validated against previous experimental work. Predicted results showed that different nozzle types plus the slight curvatures in the molds can significantly influence the flow patterns generated within the mold region, and thereby affect the finished quality of the billet products.INTRODUCTIONConventional Continuous Casting (C.C) processes are presently used to solidify more than 95% of the 1.5 billion tons of steel now produced in the world annually (World Steel in Figure, 2014). This overwhelming contribution of C.C processes for steel production is the prime motivation for this study, in order to understand and optimize the casting processes and thereby reduce internal defects and minimize energy consumption. The primary objective of all C.C processes is to maintain liquid metal purity during the casting process and to create a uniform flow within the caster so as to avoid internal and external defects. In this regard, studying fluid flows in the mold region of continuous caster has attracted many research groups, in both academia and industry, in recent years. The mold region is one of the most critical parts of the C.C. process, in which the transport phenomena associated with this region can significantly affect the quality of the final products. Thus, proper control of fluid flows within this region can be expected to greatly decrease both internal and surface defects such as Oscillation Marks (OM), inclusion entrapment, and Mold Powder Entrapment (MPE). In the casting process, the fluid flow in the mold region can be affected by a number of variables, such as; the flow control system (stopper rod or slide gate), nozzle geometry, SEN (Submerged Entry Nozzle) depth, casting speed, mold dimensions, gas injection, and the application of electromagnetic fields. One of the most influential and easily-changed fluid flows are the jets of steel issuing from the SEN. Nozzle geometries are variable, consisting of a number of exit ports, different port angles, and/or port sizes, the bottom shape of the nozzle, and the SEN immersion depth (Thomas, 2001). The SEN characteristics can markedly affect meniscus surface velocities, surface level profiles, and turbulence levels that can vary in terms of the frequency and the magnitude of their fluctuations."

Jan 1, 2016

By R. W. Thompkins

"PREFACEPart 1 of this paper, published in the September issue, covered the occurrence of uranium, radiation from uranium ores, the uranium series, radioactive theory, radioactive equilibrium and application to the radon series, the working level, the theory of emanation and diffusion, transport by thermal diffusion of radon emanating from porous rocks, and definitions.The October issue carried Part 2, which dealt with radiation theory applied to uranium ores, other radon migration factors, factors affecting the radon gas release rate, the measurement of radon gas emanation rates, the tunnel concept, variation of emanation rate with grade and estimating radioactivity from grade analysis.Part 3, in this issue, concludes the series by discussing the techniques for controlling radiation and providing some example calculations.17.0. Techniques for Controlling Radiation17.1. Altering the Factors In the Diffusion EquationIt has been noted that radon gas emanates from the surfaces of exposed ore to the mine atmosphere. The diffusion laws establish the rate of this flow. The fundamental relationship is that the rate of flow (J) is a function of the pore concentration and the distance from the surface. It is expressed as: []If any of these terms is varied, either naturally or artificially, the rate of gas emanation will be altered.Altering the Diffusion CoefficientThe literature describes many ways to lower the radon gas emanation rates which are dependent on changing the diffusion coefficient:1. A layer of ice on top of a tailings dam or on a soil.2. Humid clayey materials and swamped soils over a radon source."

Jan 1, 1982

By J. S. Stachulak

"IntroductionInco Limited operates nine underground mines on the rim of the Sudbury Basin. These operation s are situated near the city of Sudbury, 589 km west of Montreal, and 363 km north of Toronto. The ores are mined primarily for nickel and copper, and the principal minerals are pentlandite, chalcopyrite , and pyrrhotite. Total ore production from underground is in excess of 51 000 tons per day.The total volume of fresh air supplied to the nine mines is six million cu.ft/min. There are 50 main fans in operation at these mines, powered by electric motors totaling 11 200 kW (15 000 hp).The trend in metal mining during recent years towards mechanization and high-tonnage production methods has greatly increased ventilation requirements. The now of air must quickly dilute and carry away diesel exhaust fumes, dust created by mining operations, and the fumes of the explosives used in blasting. Mechanical ventilation is required to obtain the necessary volume and control, as natural ventilation pressures alone are, for the most part, inadequate and unreliable. In the event of a mine fire or other needs, close control of ventilation is essential.There are two unique heating systems presently in operation at Inco: Creighton Natural Heat Exchanger and Stobie Ice Stopes."

Jan 1, 1991

"Technological innovation is an extremely complex activity. The author presents his interpretation of innovation and the innovative process. He reviews the strategy used by participants in coal mining innovation in the past and discusses changes required to meet the challenges of the future.Introduction if an industry is to be successful and viable in today's world, the companies active within the industry must remain close to the frontiers of knowledge. The coal mining industry is no exception. Technological problems facing the coal mining industry are basically economic and social. Economic problems are those concerned with the long-term growth and profitability and social problems, with the health and safety of the mine worker and the impact of the mining operation on the environment.Technological innovation can be a key factor in meeting these problems successfully. The general belief is that the information base for technological innovation exists within mining companies and governmental agencies but current industrial and governmental strategies and structures impede the transformation of this knowledge into the mining products and processes required for growth.The purpose of this paper is to determine where Canada is in terms of coal mining on the innovative ladder. The paper as presented: defines innovation and innovative process; gives an historical ""broad brush"" review o f innovation in the industry; and suggests the challenges to be faced in the future."

Jan 1, 1985

By Roland Froidevaux

"The main object of ore reserve classification is to assess the reliability of an ore reserve estimate. Because traditional classification schemes fail to provide a means of calculating this reliability, the terms 'proven "" 'probable' and 'possible ' ore were introduced as an attempt to qualify it.Geostatistics, on its part, allows the quantification of the precision of any ore estimate and thus appears as an appropriate tool for ore reserve classification.However, the problem of fixing quantitative boundaries for ore classes remains. In order to be meaningful, ore class limits based on the degree of reliability should be customized to the commodity, and the type and size of the deposit. In addition, the precision of an ore estimate will be affected by the cut-off grade and the size of selective mining units.IntroductionPrimarily, ore reserve classification should be seen as a communication tool between the geologist and the mining engineer on one hand and the management and policy maker on the other. By classifying ore reserves, the geologist tries to answer the very fundamental question: ""What quantity of metal is there in the ground and how sure am I that it really exists?""The difficult y of quantifying the degree of reliability of reserves has led to the creation of a large suite of terms and definitions, with almost every mining company having its own set of definition standards. As lower-grade deposits are brought into production, the margin of tolerable error in estimates of the grade and tonnage is becoming much narrower. Consequently, the need for unequivocal definition of the reliability of ore reserves is increasingly acute. Most of the generally accepted systems of classification, however, describe the reliability of reserves in qualitative, rather than quantitative, terms."

Jan 1, 1982

By J. E. Udd, L. B. Geller

"Responding to requests by the Canadian mining industry and the mine regulatory authorities, the Canada Centre for Mineral and Energy Technology (CANMET) undertook a major research project on evaluation of wire rope testing in 1986. The objectives were to document, analyze, and clarify - in basic operational terms the reasons for ongoing mine shaft wire rope failures that occur despite mandatory non-destructive testing (NDT) with electromagnetic (EM) instruments on a routine basis.The project's investigative scope and depth was unprecedented. It could not have been successfully completed without the wholehearted cooperation of industry, research organizations, and mine-inspectorates at both national and international levels.The project's goals were attained. II was shown that, at present, the problem areas in Canada are oj a human , rather than of a hardware-related nature. Consequently, the principal recommendations are concerned with a more widespread, more uniform and thorough operator training and certification process at industrial and regulatory levels. Secondary suggestions concern areas of hardware-related improvements.In this paper an up-to-date general overview is given of the major project results. A n extensive reference list is also provided of papers that contain all relevant documentation in detail."

Jan 1, 1992

By Yongfeng Jia, Shaofeng Wang, Danni Zhang

Ferrihydrite is ubiquitous in soils, groundwater and aquatic sediments and is usually classified as "2-line" or "6-line" according to the number of diffuse XRD peaks. It has high surface area and reactivity; therefore, it may serve as a sink for sequestering trace metals, metalloids and organic matters via adsorption and co-precipitation. Although considerable attention was paid to the transformation process of this thermodynamically metastable solid, there are still some controversies on the mechanisms of the formation of goethite or hematite from the ferrihydrite precursor. This work systematically investigated the transformation of 2-line ferrihydrite produced by using different neutralization regent (CaO vs. NaOH) into crystalline products at different pHs (4, 7 and 8), three temperatures (25, 40, 75 °C) and in different media (sulfate vs. nitrate). X-ray diffraction (XRD), Raman, Fourier transform infrared spectroscopy (FTIR) and chemical extraction were employed to characterize the transformed solids as well as the redistribution of Fe between the amorphous and more crystalline phases. The results showed that the transformation product in nitrate media was hematite (major) and goethite (minor), which was independent of pH, temperature or neutralization reagent. In sulfate media, goethite was the dominant solid under slightly acidic conditions (pH 4) with 6-line ferrihydrite as the intermediate product, whereas neutral to slightly alkaline conditions still favor the formation of hematite. Chemical extraction results indicated that there was only ~3 % of amorphous phase detectable after aging at 25 °C for 280 days at slight acidic pH. In contrast,~70 % of the iron remained as an amorphous phase under neutral to slightly alkaline conditions. The results suggested that the pH as well as the reaction media (sulfate ion) plays an important role in the 2-line ferrihydrite transformation process.

Jan 1, 2016

By B. R. McEwen, A. N. Samis, J. H. Gray

"The Quebrada Blanca Copper Mine is an open pit copper project located at 4200 m elevation in the Chilean Andes. The deposit is a world-class supergene enriched porphyry with a mining reserve of 89 Mt of 1.3% copper. The supergene zone will produce 6.3 Mt of ore per year for more than 15 years to produce 75 000 t of cathode copper per year.This paper will review the project geology the computer model ling of the deposit, the mine planning process and the property production scheduling. With copper production .fixed by the onsite electrowinning capacity, the planning and scheduling aspects to optimize the project NPV will be discussed. Some of the logistics, equipment. personnel and process difficulties of operating a mine at 42()() m elevation will also be described. IntroductionProject DescriptionThe Quebrada Blanca Copper Mine is an open pit copper project located in the northern part of the Chilean Andes. The project is based on the supergene enriched portion of a porphyry deposit which has a mining reserve of 89 million tonnes at 1.3070 copper. Below the enrichment blanket, the protore represents a further resource.All ore will be leached, the higher grade material will be crushed and stacked for thin layer heap leaching, and the lower grade and oxide material will be dump leached as run-of-mine material. Solvent extraction and electrowinning will be used to produce 75 000 tonnes of cathode copper per year. Power will be supplied by a 40 MW on-site diesel generating plant with the waste heat being used for process heating. The leaching operation will operate in excess of 15 years and may be followed by open pit mining and conventional ore concentration of the underlying primary sulphide mineralization.Because of the dry location water will be piped 37 km from a ground water source at the Salar de Michincha. Water requirements will be approximately 1.5 million m3/year. The operation will eventually employ approximately 550 people on a 7-day-on, 7-day-off rotation with two twelve-hour shifts per day with an on-site camp. Busing will be provided to the port city of Iquique."

Jan 1, 1994

By W. Cholewa, T. S. Golosinski, J. Wedzicha

"Power requirements of belt conveyor drives are usually estimated based on the calculated belt tension which equals the sum of unit resistances to motion which the belt has to overcome. In most situations, the value of various unit resistances can be accurately predicted using published information (5,6,8,12). A simpler approach calls for use of standard formula and tabulated values of related factors which reflect various features of both the conveyor and the transported material (2,3,7) . None of these approaches is reliable when materials with unique properties are conveyed, nor if applied to conveyors working in extreme climatic conditions. Such is the case of oil sand mines in northern Alberta.There the bitumen content of oil sand causes it to behave in a unique manner. Furthermore the conveyors operate year around, with winter temperatures often lower than - 40°C. It was the objective of presented investigations to quantify the power requirements of conveyors, and related resistance to motion for the conveyors used in oil sand mines during winter operations.The other objective was to quantify the drive power requirements of oil sand conveyors during start-up. These requirements are usually much higher than those during norm al operation, and their quantification is of paramount importance for operation of large conveyors."

Jan 1, 1992

By C. Bazin

"Discrete transfer functions which use the grinding circuit feed tonnage to predict mass flowrates at reagent addition points in the flotation circuit were calibrated using a dynamic grinding/flotation simulator. Predicted values are multiplied by g/t ratios to calibrate reagent flow. The transfer functions are currently integrated into the distributed control system at Les Mines Selbaie.INTRODUCTIONThe development of adequate tools for the control of flotation processes is a continuous challenge. Since the early 1970's there have been numerous publications dealing with practical applications of flotation process control. A review of that literature shows that basic feedback and feedforward controls are the two preferred strategies for flotation processes. Some advanced control schemes started to emerge (e.g. Zaragoza and Herbst, 1988). Feedforward and feedback aim at reducing the effect of the following flotation process disturbances (e.g. McKee, 1991):0 Changes in throughput;0 Changes in feed grade;0 Changes in particle floatability (size, liberation, surface oxidation ... );0 Deliberate changes of set-point;Feedforward strategy acts on manipulated variables, mainly reagent addition, to attenuate the effect of disturbances such as changes of feed rate and grade. The feedforward strategy is driven by measurements of disturbances. Feedback control acts to compensate unmeasured disturbances, such as ore floatability, and to bring process variables to targeted set-points of concentrate or tails grades. The feedback control action is driven by set-point error, caused by either a deliberate change of set-point or unaccounted disturbances.This paper proposes a way to improve feedforward control schemes by addition of a simple model of solids transportation. The feedforward strategy considered in this paper is of ratio type, i.e. the manipulated variable is changed according to a predefined ratio between this variable and the measured disturbances. A typical example is the ratioed reagent addition to mass flow of ore."

Jan 1, 1993

By Claudia Gasparrini

"Silver occurs in about 200 minerals in variable amount s. Silver ores usually have 5 or 6 and in some instances as many as 10 species of silver minerals. All these minerals may have different flotation and leaching properties and should be correctly identified for successful recoveries. Grain size and host minerals which are critical in gold recovery, are not always as important for silver minerals.IntroductionMineralogical factors affecting the metallurgy of silver are , as for other metals: (i) the silver-bearing mineral s ; (ii) their grain size, and when dealing with very fine sizes; (iii) their host minerals; and (iv) the association they form with the silver-bearing minerals. Of these, the silver-bearing minerals are more often the cause of problems during recovery as opposed to gold whose metallurgy is generally more affected by grain size and host minerals. The problems derive mainly from the complexity of the silver mineralogy: over 200 minerals exist bearing the element in major, minor and variable amounts.As a member of group I-B of the periodic table, silver forms extensive so lid solutions with the other two elements in the same group , copper and gold. It combines with sulphur, tellurium and selenium in covalent bonds and with chlorine, bromine and iodine in ionic bonds. Where the fluid composition, temperature and pressure during crystallization allow it, silver also combines with arsenic, antimony, lead, bismuth, tin, mercury, indium, thallium, and germanium in many instances as a minor element in sulphide , sulphosalt and native element minerals. Cations which silver replaces to a lesser extent than copper and gold are iron, manganese, zinc and probably nickel and cobalt."

Jan 1, 1984

By H. Soto, J. Jara, F. Nava

A method has been developed to recover cyanide and copper from cyanidation effluents contaminated with thiocyanate and copper-cyanide complexes. Copper is precipitated as copper thiocyanate or as copper cyanide depending on pH and the concentrations of thiocyanate and cyanide present in the effluent. The copper precipitate presents good dewatering characteristics and can be easily separated from the effluent by thickening and filtration. The solution containing the bulk of the cyanide is then oxidized with ozone to transform the remaining thiocyanate into cyanide. This results in a regenerated solution rich in free cyanide that can be recycled to the cyanidation process. Careful control of the pH of the solution is required. First pH of the effluent is lowered to optimize the precipitation and oxidation stages, then the pH is restored to an alkaline value to regenerate the cyanide solution. Recoveries of over 96% copper are obtained. Free cyanide in the regenerated solution corresponds to up to 160% of the total cyanide contained in the original impure effluent. (The balance being accounted for by the cyanide obtained by the oxidation of thiocyanate). The efficiency of the thiocyanate oxidation stage is 85% to 90%. The method has been demonstrated at bench scale with effluents of various cyanidation plants of northern Quebec. A preliminary economic evaluation indicates that the procedure would be profitable for effluents rich in copper and thiocyanate. This is the case for most barren bleeds of Merrill Crowe and heap leaching operations and also certain CIP tailings of gold ores containing copper impurities. Environmentally the procedure is a better alternative than the current detoxification methods based on the oxidation of cyanide that do not allow for copper recovery and that have difficulties in producing a final effluent with acceptable levels of contaminants such as ammonia, thiocyanate and copper.

Jan 1, 1996

By E. B. Spice

"IntroductionExperience at Eldorado has been gained with a variety of mill feeds. That is, the actual mode of occurrence of the pitchblende varies to such an extent that gravity concentration has at times been decidedly unattractive and at others highly successful. Because of this, the type of ore necessary for economic recovery of pitchblende have become well known. Also, those types which do not lend themselves to gravity concentration have become familiar. In general, it may be said that when the pitchblende occurs free, to a large extent, from sulphides or arsenides of high specific gravity, and when the bulk of the pitchblende is localized rather than disseminated, excellent recoveries, high ratios of concentration, and good grades of concentrates , will result from proper gravity concentration. On the other hand, if the ore contains pitchblende intimately or otherwise mixed with large quantities of arsenides or sulphides, or if the pitchblende is finely disseminated, gravity concentration alone will yield high recoveries only at the expense of very low-grade concentrates and low ratios of concentration.According to Taggart, the ease with which gravity separation may be effected is determined by the value of the following ratio: []It is stated that when the ratio is above 1.75, gravity separation can be expected to be easy. Thus with pure U308 (S.G. over 9) and gangue consisting chiefly of siliceous material, the process can he expected to be entirely successful even when carried on at very fine sizes. Unfortunately, the term pitchblende covers a great diversity of material. Although some of the mineral occurs pure enough to have a specific gravity of about 7, the average is much lower than this and is probably well under 5. Some of the pitchblende particles have a specific gravity so low that gravity concentration fails completely. It may also be said that no degree of grinding will produce free pitchblende of sufficient density from such particles. This, then, represents the most important limitation to the application of gravity separation to pitchblende ores."

Jan 1, 1950

By R. F. Favreau, R. A. Morin, G. W. Kuzyk, P. J. Babulic, N. J. Tienkamp

"Atomic Energy of Canada Limited is constructing an Underground Research Laboratory (URL) as part of a comprehensive program to evaluate the concept of nuclear fuel waste disposal deep in crystalline rock formations. Careful blasting methods have been used to minimize damage to the excavation surfaces. Good wall quality is desirable in any excavation. In excavations required for nuclear waste disposal, the objective will be to minimize blast-induced fractures which may complicate the sealing requirements necessary to control subsequent movement of groundwater around a sealed disposal vault. The construction of the URL has provided an opportunity for the development of controlled blasting methods, especially for drilling accuracy and optimization of explosive loads in the perimeter and cushion holes. The work has been assisted by the use of blast simulations with the mathematical model Blaspa.This paper reviews the results of a recent project to develop a controlled method of full-face blasting, and compares the observed field results with the results of a blast simulator called Blaspa. Good agreement is found between the two, and the Blaspa results indicate quantitatively how the blasting may induce damage in the final excavation surface. In particular, the rock in the final wall may be stressed more severely by the cushion holes than by the perimeter holes. Bootleg of the rock between the perimeter and cushion rows occurs when the burst out velocity imparted to it by the explosive loads in the perimeter holes is inadequate. In practice, these findings indicate that quantitative rock stress and rock burst-out velocity criteria can be established to minimize wall damage and bootleg. Thus, blast simulations become an efficient way to design controlled blasting and to optimize quality of the excavation surf ace."

Jan 1, 1989

By Jorgen Elbrond, Yve Lizotte

"A mathematical procedure for determining the underground mining level layout which minimizes the total costs of excavation and haulage is described. This procedure has the advantage over previous graphical procedures of being easily amenable to microcomputers. The paper demonstrates through examples how the procedure can be applied to minimizing discounted costs of level layout as well as providing the optimal location of a production shaft.IntroductionIn recent years there have been numerous publications in mining journals concerning the minimization of mine haulage costs. With the rapid escalation of energy costs during the last decade it is not surprising that several studies applying queuing theory, simulation and mathematical programming to mining related transportation problems have been successfully undertaken. Because of the increasing availability of various sizes of computers these techniques have gained considerable popularity as evidenced by the numerous commercial packages presently being marketed.The publications concerned with haulage cost minimization are generally concluded along the lines of ' ... haulage costs at production rate x will be minimized by assigning y units to location z...'. Details about the route followed and whether it guarantees a minimization of costs are seldom considered. Route data has to date been manually resolved and used for input to developed programs. It would appear that no program has been developed to date which simultaneously minimizes costs by both fleet assignment and routing network resolution.Another important factor in transportation problems is the investment cost associated with the routes; road construction costs or, in the case of underground levels, drift excavation costs. Except for special cases the transport network which minimizes haulage costs only will not provide a minimum excavation cost solution.Zarnbo(1) illustrates by graphical and mathematical procedures how central 'facilities' can be located to simultaneously minimize investment in the transport network and haulage costs. Unfortunately these procedures and others recently applied to mine haulage problems do not provide solutions to the generalized problem. Also because these procedures are geared to graphical solutions and hand computations whereby certain possibilities can be 'visually' eliminated, they can be tedious to program and present storage capacity constraints with current microcomputers."

Jan 1, 1985

By W. G. Brown

"Review of prototype testing of two types of three-phase, argon stabilized reactor/furnace indicates specialized advantages for both, as compared with existing metallurgical furnace technology. They are: the direct-arc, plasma-torch furnace (Knipp); and the indirect, plasma extended-arc reactor (PEAR), developed at the University of Toronto.IntroductionA unique type of three-phase AC, plasma extended-arc reactor, (PEAR), shown in Figure la with gravity feed, has been undergoing development at the University of Toronto, during the past fifteen years: argon gas, introduced axially through drilled-holes in ordinary graphite electrodes, increases the length of indirect arcing and improves electrical stability while decreasing electrode consumption as compared with the performance of AC electric arc furnaces (EAF 's). Meanwhile and independently, a three-phase, transferred arc plasma furnace (Fig. 1b) has been developed by the Krupp Research Institute in Germany, using argon-stabilized, water- cooled plasma-torches with thoriated tungsten electrodes.Accordingly, the two types of furnace are similar as to both three phase and stabilized operation, but otherwise they differ in several respects, ranging from indirect versus direct-arcing to principal intended application and current state of practical development. Between them they raise questions: (I) of their present or fore seeable economics in comparison with existing processes they are intended to replace , (see Heinke and Neuschutz, 1986, for a review of the merit s of AC and DC plasma-torch furnaces); and (2) of the state of development and accessibility of their plasma theory-in-common."

Jan 1, 1991

By Frédéric Flament, Mary Goldman, Gille Barbery

"Pilot plant testwork on autogenous and semi-autogenous grinding was conducted at the Centre des Recherches Minérales du Québec (CRM) to investigate the effect of ball addition, solids concentration in the feed, specific energy consumption and feed size distribution.A mathematical model that incorporates breakage and surface mechanisms of loss of mass was developed to predict load size distribution of autogenous and semi-autogeneous mills. An empirical model was used to predict product size distribution of autogenous and semi-autogeneous mills.IntroductionAutogenous (AG) and semi-autogenous (SAG) grinding circuits offer potential capital and operating cost savings when compared with conventional grinding. Empirical relationships have been developed to calculate specific power consumption for SAG mills using pilot plant test results (Mular et al., 1989; Barrett and Allan, 1986). Empirical models can be obtained from plant practice data but are applicable only to the particular ore and set of operating conditions used in the model. Grinding mechanisms have to be taken into account in order to develop a model that can be applied to different ores and mills.In AG and SAG, the ore itself acts as grinding media, with the finer sizes being crushed by the coarser ones. Surface mechanisms of loss of mass, in which the inner core of the particle remains intact, also take place. In SAG milling, the effect of balls also has to be taken into account. The role of each one of these mechanisms changes with particle size and the model has to account for such changes."

Jan 1, 1991

By J. G. Whellock, R. J. Phillips, J. W. Matousek, P. J. Marcantonio

"Stillwater Mining Company (SMC), a partnership of Chevron Resources and Manville Corporation, produces a copper-nickel, platinum-palladium flotation concentrate at its Nye, Montana mine. The concentrate is then toll processed in Europe. Recognizing the economic advantage of treating the concentrates at the mine site and of retaining this strategic resource in the United States, the Company commissioned Tolltreck International Limited and International Process Research Corporation to conduct pilot scale smelting tests with a flowsheet that would maximize metal recoveries while satisfying the environmental requirements of operations at Stillwater.Primary smelting was conducted in a single phase, 50 k W, submerged-arc, electric furnace. Refining the product mattes to copper-nickel ""white-metal""; containing approximately 2% iron, was accomplished in a Tolltreck, IV-litre, top blown rotary converter. Over-all platinum and palladium recoveries were approximately 99%; copper and nickel recoveries averaged 94%. INTRODUCTION(l-4)Stillwater Mining Company (SMC) a partnership of Chevron Resources and Manville, operates a copper-nickel, platinum palladium mine and concentrator at Nye, Montana, located in the valley of the Stillwater River, approximately, 100 km southwest of Billings. It is North America's only primary producer of these precious metals. The Stillwater area was mined for chrome ore during World War U and the Korean War, but it was not until 1961 that attention was drawn to the similarities between the Stillwater Complex and the Bushveld Complex of South Africa with its deposits of platinum group metals.A concentrated search for platinum was begun by Manville Corporation in 1967. Its presence, together with palladium, in mineable quantities was confirmed in the period 1976-1978, and the first partnership to develop the deposits was formed by Manville and Chevron in 1979. Anaconda joined the group in 1983, and in 1985 its share was purchased by Lac Minerals. In 1988 the interests of Lac Minerals were purchased by Chevron and Manville.The decision to develop the mine and build a concentrator was made in 1986. Construction began in the fall of that year, and the first concentrate was produced six months later, in March 1987. Underground mining is done in a two-shift, five day- week operation. Since startup, the concentrator has operated on a 24 hour/day, 7 day/week schedule, treating an average of 635 tonnes of ore per day and producing approximately 10 tonnes of concentrate."

Jan 1, 1989

By Rodgers Peale

The Waite mine, as it is locally termed, is situated in the Rouyn district, in the townships of Duprat and Dufresnoy, Quebec, and lies about 340 miles north of Toronto. Due to the impetus given prospecting in the Rouyn district by the discovery, some years ago, of copper-gold ore on the Horne claims, many exploration parties were put into the field. Early in 1925, J. H. C. Waite, C. H. Ackerman, and T. Montgomery staked the claims which now comprise the mine beating their names. Trenching and diamond drilling soon disclosed an ore-body of commercial size and grade, and by the end of 1928 a shaft had been sunk 300 feet and the mine was producing 150 tons of direct-smelting copper ore a day. In the summer of 1925, Dr. H. C. Cooke, of the Geological Survey, made a reconnaissance survey of the property (1), and since that time work of the same nature has been done by several others. MacGregor has published some interesting data on the Amulet mine (2). Using as a background a summer's work on the Horne claims at Noranda, the writer spent the summer of 1929 at the Waite.

Jan 1, 1931