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By Hagen B. Schultes

"Refratechnik, a leading German refractory company, acquired a majority shareholding in Baymag in 1979. This acquisition included the mining rights for a magnesite deposit of exceptionally high purity in the Canadian Rocky Mountains.The deposit, situated in the Kootenay region of southeastern British Columbia, is of Cambrian age and has ore reserves of 50 million m. t. of magnesite. The Baymag magnesite deposit is one of the purest in the world. The deposit was opened up and mining commenced in early 1982. Production of various caustic calcined magnesia products in a rotary kiln operation in Exshaw, Alberta, was started in mid-1982. The product application ranges from pulp and paper production, to the chemical industry, and to animal feed supplements. Since the end of 1983, Baymag has produced high-quality refractory grade fused magnesium oxide f or specialty applications in the steel industry.IntroductionAlthough magnesium is the eighth most abundant element in the earth and forms about 2.06% of the earth's crust, common knowledge about magnesium and especially magnesium oxide (MgO) as an industrial mineral is not widespread. The reason might be that, with the exception of magnesium metal as the element, magnesia products are not generally sold as final products to end-users, but rather are needed for the production of final products such as steel, cement, paper, etc. Therefore, some general comments are given about the different sources of MgO, the different products, world production and consumption , before Baymag specifics are addressed. Magnesium metal will not be discussed in this paper."

Jan 1, 1986

By D. S. Harries, R. I. L. Guthrie

"The use and development of scientific disciplines involved in heat transfer technology, related in general to steel masses and in particular to batch annealing, is demonstrated.The measurable benefits of cooperation are tabled; on industry's side, this includes a 10 to 20 per cent decrease in annealing time, and a corresponding higher productivity potential. The returned benefits to the academic engineer include the development and refinement of new computer algorithms related to heat transfer mechanisms.IntroductionThe object of this discussion is not really to describe an improvement in the batch annealing operation, but to use the project as an example of joint work between industry and the academic field. Both parties have derived a number of substantial benefits. These benefits will be listed, of course, but the batch anneal project will be used just as a typical example of cooperation. Why it happened will be explained, what actually happened will be described and finally some thoughts about areas in the field of heat transfer, where cooperation will continue in the future, will be highlighted. The paper will close with a message.Industry OverviewAt the outset, perhaps it should be understood that, in general, the Research Department at Dofasco believes that the universities and their staff are not destined for the problem-solving role in industry. Its opinion is that the university staff is best suited to the gathering and dissemination of knowledge. There are other more professional problem solvers for industry, including the steel industries' own individual investigation groups. Dofasco, therefore, looks to its academics for consultation. It benefits by bouncing ideas off them, and through rubbing shoulders with them in such a way that their knowledge can be tapped, and thus establish more effective scientific routes. In addition, the younger and brighter engineer s can broaden their scope through the intellectual contact with experts. The university professors can, it is believed, benefit through closer contacts with result-orientated research. They can also develop an appreciation of the fiscal control necessary for industrial investigations, without government subsidies, grants, etc., and with massive overheads."

Jan 1, 1982

By C. -G. Berg, J. T. Tuominen

Since the beginning of the 20th century, energy efficiency in copper smelting has been improved by a factor of 30 as a result of adopting autogenous smelting processes with more and more efficient waste heat recovery systems, increasing use of oxygen enrichment and countless smaller improvements. Nevertheless, energy in the form of electricity and fuels continues to be the biggest contributor to smelter operating cost. Metallurgists around the world are under constant pressure to find ways to save those precious megajoules wherever they can. One area that is often neglected is the removal of water contained in concentrates. Copper flotation concentrates, as delivered to a smelter, typically contain some 8–12% moisture. This moisture must be removed and the only way to do this is by evaporation, which consumes a lot of energy. Several studies have shown that the most efficient way of removing this moisture is to use a steam- heated indirect dryer—the least efficient way being the feeding of wet concentrate directly into a smelting furnace. Even in the most efficient smelters that employ modern steam drying, the energy needed to remove the moisture from the concentrates constitutes over 15% of the total energy consumption of a smelter – acid plant complex. The major steps that have been taken in the past to lower the energy consumption of moisture removal are the use of drying in general and the adoption of steam drying. The energy efficiency of concentrate drying has been improved by 25% over the last 20 years. The latest giant leap in this area, the Kumera Heat Recovery Dryer, is the result of Kumera’s long-term development in improving the efficiency of concentrate drying. The Heat Recovery Dryer is a novel application of steam drying that uses the waste heat from dryer exhaust gas for drying concentrates, thus, cutting the energy consumption by a further 30%. This paper discusses energy consumption in copper smelting in general and analyzes the energy consumption of concentrate drying in detail. Different dryer types are compared especially from the perspectives of energy consumption and environmental emissions. Applicability of different energy sources for drying is evaluated and discussed. The principle of the Kumera Heat Recovery Dryer is presented along with results from extensive continuous pilot plant tests. Potential applications and the feasibility of Kumera Heat Recovery Drying technology are discussed.

Jan 1, 2019

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 Owsiacki. Leo

Both the 8 and A-C ore zones of the Heath Steele Mine contain extensive areas of brecciated ore. This texturally distinct ore type has been tentatively defined as a fragmental or conglomerate massive sulphide. It consists of subangular and round fragments of pyrite up to 30 em in diameter, smaller quartz clasts and chloritized, mineralized, lithic fragments in a faintly banded matrix of pyrrhotite and chalcopyrite. The chalcopyrite is concentrated in the pressure shadows of the fragments and appears to reflect remobilization and recrystallization during a deformation which postdates the formation of the clasts. A generally conformable relationship with the underlying banded pyrite-sphalerite ore, discordancy with the host banded argillites, spatial distribution and similarity to Kuroko ores suggest that the fragmental ore is of primary origin and that breccia formation is probably due to soft-sediment slumping or sliding initiated by volcanic explosions, as hypothesized for similar massive sulphides in various Japanese mines. It is nevertheless recognized that certain structural features in some clasts may result from deformation preceding the formation of the breccia, and at this stage a tectonic origin cannot be completely eliminated.

Jan 1, 1979

By Egon Berger

"In recent years the trend in underground metal mining has been toward ever increasing blasts (often in excess of 100 000 tonnes per blast). The Canadian mining industry has been in the forefront of these developments (VRM, LBHand cable bolting to stabilize the surrounding rock /avoid dilution).In order to move these large tonnages, efficient loading systems have to be developed. This paper shows the technical and economic implications of various loading scenarios based on a predefined (however typical) stoping layout by:1 . discontinuous loading using scoop trams;2. discontinuous loading using scoop trams in combination with low profile dump trucks;3. continuous loading (based on loading performance/operating costs of the AL-60) in combination with low profile dumptrucks.The cost advantages of production concentration (loading large tonnages from one drawpoint}, made possible by continuous loading as opposed to sea tiered mining (small tonnages), using discontinuous loading in tunnelling will be presented, based on actual results obtained in a decline development project using the A L-60 continuous loader. At as pacing oj150m between turning bays, it was found that continuous loading was out performing discontinuous loading in approximately80% of the overall decline development.Fuel consumption and total cycle times are compared, and a cost assessment is presented.INTRODUCTIONThe subject of this paper deals with the introduction of continuous loading in underground metal mining and single heading tunnelingContinous mining has been widely introduced both in underground and surface coal mining. Mechanized longwall mining, utilizing moveable shield supports in combination with double drum- shearers, has resulted in a tremendous increase of coal production per working face.This has led to very concentrated mining operations. In a typical German coal mine two mechanized longwalls nowadays would produce the same amount of coal as 12 to 15 longwalls in the 1950s. One longwall in a South African coal mine has recently set a world record with 120 000 tonnes of coal in a single month.In underground metal mines-often due to the irregular nature of the ore body and the mining methods used-mining is far less concent rated. Ore is often produced from a large number of stopes and production sections. Sometimes, a substantial amount of ore may be obtained from development in the orebody (40070 at a German base metal mine)."

Jan 1, 1989

By E. O. Reinbold, J. H. Johnson, D. E. Keski-Hynnila

A field laboratory for use in monitoring underground mine air quality has been developed and proven in underground service. The laboratory includes two separate monitoring systems and data analysis computer programs. One is the Mine Air Monitoring Laboratory (MAML). This is an enclosed trailer-mounted laboratory capable of monitoring CO2, CO, NO, NO2 particulate matter, S02, NH3, temperature and air velocity in the mine drift. The MAML can be up to 1000 feet away from the area being monitored. The second system is the vehicle data system, mounted on the load-haul-dump (LHD) vehicle, which monitors engine speed, engine fuel rack position, exhaust gas temperature, ambient temperature, vehicle speed, CO2 concentration near the operators' breathing zone and the mode of operation of the vehicle. The two computer data analysis programs and the two field instrument systems comprise a complete Mine Air Quality Laboratory (MAQL) for underground pollutant studies. Data from the systems have been used for studying vehicle duty cycles, ventilation systems, exhaust after-treatment devices, vehicle exhaust system design and portable instrumentation. Descriptions of the laboratory are presented along with a summary of some of the experiments performed. The capabilities of the lab are discussed, along with future potential uses.

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 Corby G. Anderson

"Laboratory testing using industrial electrowinning solutions was performed to determine the best Faradic cell efficiency for copper cathode recovery. The variables tested were solution flow rate, recycle ratio, reagent addition combinations of guar and thiourea, and addition amount of acid mist suppressant CAL FAX DBA-70. The ideal flow rate was determined to be 37.85 litre/min. Additionally, the ideal recycle ratio was determined to be 10:1. The reagent addition scheme from testing which showed the highest efficiency was with 680 g/tonne guar of cathode copper won and with 135 g/tonne thiourea of cathode copper won. These values for reagent addition only reflect the best efficiency and should not be assumed to be the best reagent concentration for copper purity. The amount of foaming agent that showed the best efficiency was the test with 2 ppm of foaming agent. Again this amount of reagent is based on efficiency and should not be taken as the amount of reagent that does the best acid mist suppression. Additional proprietary surface characterization work indicates these additives may have enhanced the cathodic reduction surface morphology of some noble minor impurities. Subsequently, these operating parameters were introduced into the plant operation and proved successful after their implementation.BACKGROUNDCopper electrowinning has become a dominant technology for the production of marketable high grade copper cathode products. The literature is robust with applications, optimizations, additives and developments in this field (e.g., Classens & Harris, 1990; Pfalzgraff, 1999; Tozawa, Su, & Umetsu, 2003; San Martin, Otero, & Cruz, 2005; Houlachi, Edwards, & Robinson, 2007; Romero, Avila, Fester, & Yacher, 2007; Marsden, 2008; Beukes & Badenhorst, 2009; Vainio & Weatherseed 2009; Bender, 2010; Joy, Staley, Moats, Perkins, Uhrie, & Robinson, 2010; Khouraibchia & Moats, 2010; Moats, 2010; Rutledge & Anderson, 2015).Tests of forty hours duration and at room temperature were performed for this project using samples of actual industrial electrowinning solutions. These industrial solutions were comprised of 60 g/L Cu plus 65 g/L Ni, 3 g/L Fe, and 30 g/L free sulfuric acid. Other confidential minor impurities were also present including some solubilized precious metals. The testing apparatus utilizing stainless steel anodes and cathodes is illustrated in Figures 1 and 2."

Jan 1, 2016

By Peter N. Calder, Garston Blackwell

"The history and analysis of a structurally complex open-pit rock slope displacement are described. The failure involved part of the only haulage access to a major ore zone. The ongoing movement was monitored using an electronic distance measuring device and a continuous electronic system specifically designed for the problem. A major fault, infilled with a thick clay gouge, and intersected by other structural features, resulted in the transfer of weight to a rock mass acting as a buttress in the lower regions. Movements were affected by blasting vibrations and ground-water conditions. Remedial measures included limiting blasting levels and lowering the water table, following which mining was successfully resumed in the area.IntroductionBrenda Mines is situated in a mountainous region of the southern interior of British Columbia; the location is indicated on Figure 1 . The elevation is approximately 5000 ft , giving a climate typical of more northern areas. Temperatures vary from -30°C to + 30°C. A snowfall level of 400-800 cm is experienced from November to April. During the spring, a heavy runoff period occurs, increasing the pumping rate in the pit from a nominal 50* gal/min to 2000 gal/min.The orebody occurs in a fractured Jurassic rock mass referred to as the Brenda Stock; it is approximately 3000 ft in diameter and 1000 ft deep. The Brenda Stock was intruded into the Upper Triassic Nicola Group tuffs, breccias and argillites. Mineralization occurred in several stages in a relatively homogeneous quartz diorite host rock. Shearing, faulting and jointing occured later in many stages, producing mineralized clay gouge zones up to several feet in thickness'> , These gouge zones have a major effect on the stability of the pit walls.The structural geology of the pit walls is otherwise relatively simple. Figure 2 is a conventional pole diagram indicating the orientation of the rock structural systems which are present. The major system has an average dip of approximately 70 degrees to the south and strikes approximately east /west , it is not undercu t by the pit slope. Figure 3 is a photograph of part of the North wall on which this system is indicated.The basic pit design consists of a working bench height of 50 ft, with a final wall height of 100 ft between berm s and a 100-ft-wide berm every second lift. Combined with a local face slope of 50-70 degrees, the resulting wall slope, not including haulage roads, is 45 degrees. Figures 4 and 5 are generalized plans and sectional views of the pit."

Jan 1, 1980

By P. D. Duever, P. J. Readyhough, P. L. Silveston

"The Plastojrost apparatus, developed by Ritter and Juranek in the 1960s has been modified primarily, to automatic temperature control. A limited investigation of carboniferous and cretaceous coal samples showed that both coals undergo the same changes when coked in the Plastofrost unit. Dilatometer plastic range and total dilatation correlated well with the Piastofrost plastic range for the carboniferous Appalachian coal, but not for the cretaceous western Canadian coal. Work to elucidate the characteristics of some western Canadian coals using the Plastofrost is in progress. IntroductionCoke quality is critically affected by the properties of the plastic phase in the coal-to-coke transformation. Since these properties exist in a coke oven and are not measurable, they are approximated by standard laboratory tests, such as free swelling index, plastometry, and dilatometry. Of particular interest to us is the anomalous behaviour of western Canadian coals of cretaceous origin. By many predictive methods, these coals should not produce coherent cokes. When coked in pilot scale test ovens, in fact, they produce cokes of good stability. They typically make up 20% of Japanese coking coal blends and are also used in South Korea, Mexico, Brazil, and several European countries. This anomalous behaviour has been studied in our laboratory<8.9.1O). Plastic properties or rheological tests examine just the external ramifications of the coal-to-coke transformation. To understand the process, however, microscopic examination of the coal and coke seems to be needed. The Plastofrost unit has been shown to be a convenient means to this end."

Jan 1, 1987

By R. D. Brown

"Concerns over possible additional involvement of governments in Canada&apos;s mining industry and its adverse effects on the industry &apos;s growth were a frequent topic of conversation at the recent Annual General Meeting of the CIM.Still painfully fresh in the minds of industry leaders are the disastrous results of the federal/provincial scramble over resource revenues which followed the relatively high metal prices witnessed in 1973/1974. At current metal prices, it is not surprising, then, that a nagging fear persists that these governments may once again restructure tax and other rules to the detriment of the industry and the country.This month, Tax Notes recalls some of the progress which the industry and governments have made in 1979 in the taxation of the resource sector, and urges that higher metal prices be viewed in light of the costly exploration, development and production efforts which will be required for the industry to continue to play its important role in our economy.IntroductionAt the Annual General Meeting of The Canadian Institute of Mining and Metallurgy, held in April, 1980, it was abundantly clear that there is widespread concern in Canada&apos;s mining industry that this country&apos;s tax regime may once again be blasted by ill winds of change.A stable and more favourable tax climate, as well as higher metal prices and profits, made 1979 almost a golden year for Canadian mining. However, from the tax viewpoint at least, there are some spots on the silver lining as apprehensions grow that the well-heralded battle between federal and provincial governments over oil and gas may spread to the mining sector, with serious damage to the industry and its newfound federal-provincial cooperation."

Jan 1, 1980

By D. Assmus, S. Daly

"Poor fragmentation, high toe areas, excessive shovel wear and tear and frequent re-drilling and blasting of certain tough waste rock zones in the Lomex pit in the early 1980snecessitateda better understanding of these zones.A combination of geotechnical mapping, penetration rate data, toe elevation contouring and observations on shovel and drill performance were used to determine the character and extent of these zones. This complementary information was brought together by the geotechnical mapper and the explosives supervisor.Distinct zones with particular geological characteristics were determined. A classification system with three types of blast and dig-ability was developed for the Lomex pit and later applied to the Valley pit, with considerable success. The three blast zone types are called ""poor"", ""f air"", and ""good"". This system will be described in detail.The explosives supervisor adjusted the blasting parameters to each zone, in particular, pattern spacing and burden. The resulting powder factors showed considerable variation between the poor, fair and good zones within each pit as well as between the two pits. Pattern geometry and length of collar were also adjusted to improve fragmentation. Cost savings resulted from less re-drilling and less wear and tear on drills and shovels in the poor zones and more efficient use of explosives in all the zones. An annual estimated cost savings of well in excess of $1,000,000 is realized at Highland Valley Copper. A by-product of the mapping was the determination of unstable structures and water flowing zones. Blasting was then adapted to this information for enhanced slope stability."

Jan 1, 1992

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 T. G. Carter

"This paper traces the introduction and use of fire-resistant (FR) fluids in hydraulic systems at underground mines in British Columbia. Reasons are given for the decision to make compulsory the use of such fluids in underground mines after January 1, 1975 by provincial government legislation. Examples are quoted of fires caused, or supported by, mineral oils which have resulted in loss of life and property.The basic principles of hydraulic systems are reviewed in order to present the effects of changing a system operating on mineral oils to one which must operate on FR fluids. Approximations of the cost of system conversions are given for certain machines and a brief survey is made of the various pieces of underground equipment which contain hydraulic systems operating on FR fluids.IntroductionIn 1948, Admiral Ben Moreel , who was then president of Jones and Laughlin Steel Corporation in the United States , wrote:""Although safe and healthful working conditions can be justified on a cold dollars-and- cents basis, I prefer to justify them on the basic principle that it is the right thing to do. ""(1)This paper will outline the introduction and use of fire resistant fluids at mine operations in the Province of British Columbia. That it was the right thing to do in enhancing the safety and well-being of every person employed at the mines was the prime consideration. That it may be justified on a cold dollars-and-cents basis is left for the reader to judge. The prime purpose of this paper is not so much to trace the efforts made by the individual mines to successfully introduce and use fire-resistant (FR) fluids in hydraulic systems, but to present an over-all picture of what occurred, in a collective sense, at the majority of mines which have ended up with a working FR fluid system. The problems attendant to the introduction of an FR fluid will be recounted in a manner which, it is hoped, will prove useful to those contemplating system conversions. Because the predominant fluid now in use is an invert, or water-in-oil. emulsion, this will be the FR fluid considered.Fluid Loss from SystemsThe quantities of various fluids required for use in hydraulic systems are extremely high. This statement in itself may appear to be redundant in that the capacities of the various systems should demand that a fixed amount of fluid is required for the initial filling of a system, with a further amount of fluid scheduled for use during subsequent fluid changes, i.e. at the recommended change intervals . If only this were the case!"

Jan 1, 1981

By Gary Hawthorn

"Laboratory testing, by Lakefield Research determined that, in spite of the relatively high pyrite content of the Golden Rose (GR) orebody, it would respond well to gravity concentration followed by either direct cyanidation or cyanidation of a flotation concentrate.Economics favoured the former choice which became the adopted process flowsheet.During the latter stage of the feasibility study, the Dekalb 500 tpd copper flotation mill became available in the Highland Valley of British Columbia. This plant was purchased by Emerald Lake Resources, and was moved to the Golden Rose minesite.The crushing and grinding circuit were retained and the original flotation circuit was replaced with a 400 mtd gravity plus cyanidation circuit. The plant commissioning commenced in August 1987, and by early September was processing 300 tpd.HistoryThe Golden Rose Mine was operated by Cominco during the period 1935 to 1941 processing 146000 tons grading DAD oz/t Au at a rate of 110 tpd .The process flowsheet consisted of gravity plus amalgamation followed by cyanidation.Metallurgical TestingThe underground mining areas were below the level of Emerald Lake, and had flooded on abandonment in 1942.Samples for metallurgical testing were collected from exposed faces underground once the mine had been dewateredin 1984.In spite of the anticipated high pyrite content of the ore samples, it was apparent from the success of the previous operation that the pyrite did not interfere with the gold recovery.Laboratory testing, at Lakefield, determined that, from a technical perspective, two metallurgical options existed, as shown in Tables I and 2.Note that the very high pyrite content of the ore precluded the possible direct marketing of concentrate. The process flowsheet, therefore, had to contain a cyanidation circuit for the operation to succeed financially.Although the anticipated metallurgy was excellent with either option, the 2.1 % increase in gold recovery with the gravity plus cyanidation option made this the clear choice."

Jan 1, 1989

By D. G. Osborne

"The paper describes current &apos;tools&apos; available to the coal preparation engineer for carrying out the evaluation of data, process design, selection of equipment and layout and design of coal preparation plants.Laboratory and pilot-scale equipment is widely available and test data can be used to provide valuable design information regarding capacities and sizing for specific applications. Pilot plants, unless they fit closely to the basic design envisaged for the eventual commercial plant, C(In often provide only limited useful information.Computing aids such as equipment simulation, plant modelling, etc. are becoming widely applied for establishing design criteria and investigating flowsheet options . Other computer tools such as computer-aided design systems are taking over some of the time-consuming tasks of detailed layout and design and estimation of quantities, etc. and providing the means for detailed examination of more than a single option.Besides computer modelling there is also physical modelling of plant buildings, plant sites and entire proposed mining areas; the former can be especially useful in finalizing design and threading&apos; pipework and electrical reticulation systems through the plant without the resultant risk of impairment of flow, obstruction or wastefulness. Not the least of the benefits of plant models is the assistance provided for training new operating and maintenance staff."

Jan 1, 1985

By J. M. Harrison

"The Nor-Acme gold mine is near the shore of Snow lake, northern Manitoba, about 75 miles east of Flin Flon. The claims that form the nucleus of the present property were staked in 1927 by C. R. Parres, but it was not until 1941 that diamond drilling disclosed the potentialities of the deposit. In that year a large-scale drilling programme was begun under the direction of F. Ebbutt, of the Howe Sound Exploration and Development Company. Two orebodies were outlined containing about 5,000,000 tons of ore assaying $5.25 in gold per ton. In August, 1945, construction of the mining plant was begun.All information and interpretations contained in the following brief description of the deposit were obtained from surface mapping.GEOLOGYThe ore lies at the contact between a coarse, basic, pyroclastic rock, which is the hanging-wall, and grey, acidic lavas and sediments, which form the footwall. These rocks are cut by irregular basic intrusions that vary in composition from diorite to hornblendite, and by small acidic dykes . Tourmaline is locally common in the hanging-wall rock near the ore.The ore occurs in brecciated, silicified, and ·slightly carbonatized rock. Arsenopyrite is the main sulphide and the gold is associated with it. Pyrrhotite is locally abundant and minor pyrite, chalcopyrite, sphalerite, galena, and cubanite have been noted. Two orebodies are known, the Dick and the Toots. The Toots, which is much the smaller of the two, is about 1,000 feet west of the Dick."

Jan 1, 1954

By Andrew Neale, Len Anderson, Rob Perry

"P.T Freeport Indonesia's new SAG mill commenced operation in February 1995. The reclaim system consists of four apron feeders and a series of three conveyors. The feeders are used in various combinations and relative speeds to blend SAG mill feed. It was recognized during design of the control system that effective automatic control would require dead-time and gain compensation. The successful implementation of the Dahlin algorithm to provide variable dead-time and gain compensation is described and compared to Pl control.RésumeL'opération du nouveau SAG de P.T. Freeport Indonésie a commencé en Février 1995. Le système de réclamation du minerai comprend 4 alimentateurs à tablier et une série de 3 convoyeurs. Les alimentateurs sont utilisés dans une combinaison variée et à des vitesses relatives différentes pour mélanger l'alimentation du SAG. II a été reconnu pendant la conception du système du contrôle que le contrôle automatique demandera une compensation pour le temps de délai et le gain. L'implantation de l'algorithme de Dahlin a été un succès et ii est comparé avec le contrôle Pl.ResumenP.T. Freeport Indonesia inició la operación de un molino semi autogeno (SAG) en febrero de 1995. El mineral es alimentado desde cuatro alimentadores tipo apron feeders y una serie de correas transportadoras. Los alimentadores se utilizan en variadas combinaciones y velocidades para mezclar apropiadamente el mineral que alimenta al molino SAG. Durante el diseño del sistema de control se detectó que el control automático requería de un tiempo muerto y compensaci6n de ganancia. En este documento se describe la implementación del algoritmo de Dahlin, que ha sido exitosa y ha permitido proveer tiempos muertos variables y compensaci6n de ganancia. Ademas se hace una comparaci6n con el Pl de control."

Jan 1, 1996

By Robert B. Parsons

"On February 15, 1984 Finance Minister Lalonde presented to Parliament a budget heavily flavoured with the current consultative mood of the Government. Notably absent from the budget was the typically long list of tax changes of immediate effect. Instead, the bulk of the budget papers consisted of a wide variety of white papers on significant topics that would have a far reaching impact on the future of all Canadians.This month. Tax Notes examines the highlights of Mr. Lalonde&apos;s budget.IntroductionThe mining industry need not feel slighted in Finance Minister Lalonde&apos;s budget of February 15, 1984: the budget offered very little that would have an immediate impact of significance on any sector of the economy. Apart from some welcome announcements for the oil and gas industry to postpone the implementation of certain negative tax changes, the budget was surprisingly scant in terms of definitive tax proposals.While massive revisions in the rules relating to pensions, small businesses, commodity taxation and profit-sharing were brought forward, many of these proposals are set out in White Papers or other documents to be implemented into legislation only after further consultation. Canadians will accordingly have to wait for such promised benefits as increased RRSP and other pension contribution limits."

Jan 1, 1984