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By T. Patching

"COAL MINING at Canmore, Alberta, began in 1886 and continued there until July 1979. A comprehensive review of the Canmore operations is not available, but the following notes have been prepared to accompany references where more complete information about the operations and problems encountered may be found. The coal measures at Canmore occur in the Kootenay formation of Lower Cretaceous age, and at least twelve seams of varying thickness and horizontal extent have been recognized. The coal, of high quality, ranges from semi-anthracite in the lower seams to low-volatile bituminous in the upper seams. The strata of the Canmore area form an asymmetric syncline, the western limb being overturned to vertical by the Palaeozoic strata of the Mount Rundle Range. The eastern limb is more gently sloping, its dip ranging from IO to about 40 degrees. Tectonic movements have created a troubled and complex coal field. The seams were initially not regularly or uniformly deposited, and mountain-building activity folded and faulted the strata and in many places sheared and crushed the original structures of the coal(l,2,3,4,5). The lack of continuity of the seams, and changes in thickness and inclination, made longrange mine planning difficult and necessitated flexible methods of work, and the local disturbances sometimes created problems in roof support and contributed to the occurrence of sudden outbursts of coal and gas. Almost all of the coal was produced from underground operations but a few relatively small V-cut mines were worked along surface outcrops. These have all been filled in and revegetated. The basic method of underground work was by room-and-pillar, but many variations in layout and sequences were employed. Gangways were driven along the strike and pillars were formed by raises or slopes and crosscuts. Pillars were recovered to the extent that ground conditions permitted by splitting or taking 'skips' from the sides of the raises or splits or crosscuts. Because of the variety of conditions that existed in the field a considerable range of cutting, loading and support methods were employed, including hand-loading, slushers and scrapers, duckbills and shaker conveyors, mobile loaders and continuous miners. Except for the coal obtained by continuous miners most of the coal was broken out by blasting when driving narrow working places or by air picks when cutting from a face. In the early years hand loading onto cars or conveyors predominated. A continuous miner was introduced into Canmore in 1959 on a trial basis, and another was subsequently acquired, in order to minimize blasting and hand loading, but problems in operating these machines in inclining seams and difficulties in providing safe roof support for the machines and workers restricted their efficient use. Another major difficulty with the use of continuous miners was the problem of ventilating development headings in the very gassy conditions. In the early 1970s they were replaced by electric and compressed air slushers and scrapers, resulting in a general increase in productivity, roof control, safety, and flexibility of operations(1.6,7,8.9)."

Jan 1, 1985

By B. J. Morash

"The six mineable coal seams at Quintette Coal Limited (QCL) vary in thickness from 0.75 m to 9.5 m and include waste partings varying from 0.2 m to 4 m thick. They are highly faulted and folded in some areas and dip angles range from horizontal to 90 degrees. Due to this variation in thickness and orientation of the seams, several different mining methods are used to economically recover the coal. The key to QCL's success depends on the timely collection of detailed geology data and on mine planning which allows selection of the best method of mining to achieve maximum coal recovery with minimum dilution and cost. This paper will describe the various methods of mining employed and the criteria used to select them. IntroductionThe unique challenge in mining at Quintette Coal Limited (QCL) is to extract large quantities of coal and waste, from a complex orebody, with constricted areas for mining and waste disposal. This challenge dictates the mining of as many as four 10 m benches per year, through complicated coal structures, from three separate open pits with a total of 19coal seams. This total is made up of 5 seams at Mesa, 8 seams at Wolverine and 6 seams at Shikano (Fig. 1).Computer-aided mine planning and scheduling is carried out on a routine basis to ensure the prudent selection of methods of mining, to produce plans which ensure delivery of appropriately blended raw coals to the plant, and to coordinate disposal of large quantities of waste.Several methods of mining are applied simultaneously, and as a result coal and waste are released at different rates from twelve to fifteen mining faces. In order to ensure uninterrupted flow of material from excavators and to utilize the truck fleet effectively, a computerized dispatch system is employed.GeologyGenerally, geological conditions can be related to structural (faults and dip angle) and/or stratigraphic (coal seam and interseam waste thickness) sources"

Jan 1, 1988

By M. B. McClenaghan

"Abstract- The Ontario Geological Survey completed a regional till sampling project between 1984 and 1988 in the Black River Matheson (BRiM) area of northeastern Ontario. The project was designed to aid mineral exploration in the area by documenting the Quatemary stratigraphy and by providing a database of till geochemistry. This paper summarizes the results from the sonic overburden drilling component of the BRiM project carried out in areas of thick drift (> 5 m). Results from till sampling in areas of thinner drift (<5 m) have been presented in a separate paper. Approximately 1100 till samples were collected from 225 sonic drill holes. For each sample, the non-magnetic heavy mineral (> 3.3 S.G.) and fine (-250 mesh) till fractions were analyzed geochemically and visible gold grains were recovered by panning.Gold and other multielement geochemical data were plotted using proportional symbols to high­ light anomalous areas and to demonstrate how to evaluate and interpret a very large geochemical dataset. Thresholds for identifying anomalous concentrations of Au in till samples, defined using the 95th percentile, are 1500 ppb and 12 ppb for the heavy mineral and fine fractions, respectively. The heavy mineral fraction of older tills and to a lesser extent, Matheson Till, display a number of significant gold anomalies in the study area, which in a few cases, are associated with base metal anomalies. The vertical position of gold anomalies in Matheson and older tills varies from directly on bedrock to several meters above the bedrock surface and can provide information of relative ice transport distances."

Jan 1, 1994

By G. Simard, M. J. Gagnon, A. Charette, G. Péloquin

"In the Bayer process, flocculants are introduced in the feed well to enhance the aggregation of red mud particles. To gain a better understanding of the flocculation process taking place in the feed well, shearinduced flocculation was achieved using concentric rotating cylinders (Couette system). Our previous work has shown the effect of the shear rate on the aggregate’s settling rate and on residual turbidity of the Bayer liquor. In this study, the data demonstrated that the aggregates at selected shear levels appear as a population with a widespread size distribution. By using an imagery apparatus, it was possible to measure the size, the shape, and the sedimentation velocity of the larger red mud aggregates. Marked differences were noticed in the characteristics of the aggregate populations obtained under various shear conditions. IntroductionIn the Bayer process, red mud particles are separated from the liquor in primary clarifiers and washing circuits under the action of synthetic polymers. The agitation in the feed pipe and in the feed well affects the flocculation since the aggregates are formed by the interaction between the red mud particles and the polymers. A minimum shear rate is needed to initiate the aggregation process of fine particles (Warren, 1975).Different systems have been used to study the effects of shear on the flocculation process (Tampo and Watanabe, 1979; Klimpel and Hogg, 1986; Scott et al., 1996; Biggs and Lant, 2000). The Couette flow system has been used in some studies on shear-flocculation (Oles, 1992; Farrow and Swift, 1996a, 1996b). The importance of the physical properties of the aggregates in determining the efficiency of the solid-liquid separation process within a given gravity thickener has been demonstrated (Farrow and Warren, 1989a)"

Jan 1, 2002

By D. M. Baird

ABSTRACT A revised stratigraphic section of Carboniferous rocks which occupy about 700 square miles in southwestern Newfoundland is presented. Gypsum, which outcrops at thirty-five different localities, is found to belong to four principal zones in the section. The newly mapped structure of the region makes the otherwise random occurrences of gypsum appear as systematic outcroppings of the four zones which follow the regional structure. Details of several occurrences are given and those most advantageous for exploitation are pointed out. INTRODUCTION ROCKS of Carboniferous age occupy an area of about 700 square miles in southwestern Newfoundland. They occur in a belt, from 5 to 15 miles wide, which extends from the head of St. Georges bay southwestward for 65 miles to the mouth of Little Codroy river (see Figure 1). Several hundred feet of Pennsylvanian rocks overlie rocks of Mississippian age, which are at least 12,000 feet thick. The rocks of the Carboniferous section of Newfoundland are very similar to those of Nova Scotia and represent a northeastward continuation of the same palaeophysiographic conditions. Mississippian measures of Windsor equivalence include thick gypsum deposits which outcrop at many places and which extend under large areas as indicated by sink-holes and by structural considerations.

Jan 1, 1951

By Geo. D. Furse

"The information that has accumulated regarding the geology of the McIntyre mine is the result of work by several geologists and, as might be anticipated, their conclusions have varied. The writer has been at the mine only a few years and due to wartime conditions has not been able to study and review the mine's geology as thoroughly as his predecessors. Where the writer's conclusions differ from those of previous authors, it should be noted that he has been unable to re-map all the workings in which flow contacts are exposed, but has endeavoured to describe the structure as indicated by the evidence he has been able to observe. He wishes to thank H. G. Skavlem and H. S. Robinson for their constructive criticism of this paper, and McIntyre Porcupine Mines, Limited, for permission to present it. A list of references is appended.The McIntyre mine is in the central part of the Pearl Lake section of the Porcupine gold area. It includes the ""bed of Pearl lake, twenty-two claims to the north of the lake, and the Platt Vet lot to the southeast. It is bordered on the west by the Hollinger and on the east by the Coniaurum property. The chief ore-bearing veins on the upper levels lie very close to the north shore of the lake, but some occur under the lake, and others of lesser importance are found to the southeast. The most important veins are Nos. 3, 5, 7, 10, and 25. Some of the veins extend into the Hollinger. on the west and others cross the east boundary into the Coniaurum. The total production since the commencement of milling in 1912 to the end of March 1947 has been $168,448,218 in gold and silver from the milling of 17,690,963 tons of ore."

Jan 1, 1954

By I. Clauzier

"A new method of managing electricity to reduce costs andimprove profits is discussed. Before taking steps to better manage electricity, it is essential to understand the terms peak demand and energy and their effect on an electrical bill. Knowing how electricity is presently used in the plant is the important first step. This goes beyond knowing how much electricity is used at each time of day, the individual contribution of every plant process to the total plant load must be known. At this point, each process is ranked according to importance to production in order to identify possible savings opportunities. These improvements can then be implemented by simply improving operations or in some cases, through the use of an automatic control system. IntroductionIn today&apos;s competitive environment, controlling costs is becoming increasingly important. All costs associated with operation must be controlled in an effort to improve competitive position. Electricity is a cost component that can and should be managed.Although the term management can mean different things to different people, the intent is always the same: to make the best possible use of existing resources, whether labour, equipment or electricity.The simple approach to managing electricity presented in this paper consists of four steps.I. Understand Peak Demand and Energy. It is essential to understand the terms, peak demand and energy and their effect on your electrical bill.2. Know Your Present Electrical Use. The present use of electricity in the plant must be known. This goes beyond simply knowing how much is used at each time of day, to accurately knowing how each process is contributing to the total plant electrical load, and ultimately to the electrical bill.3. Identification of Savings Opportunities. With this information, energy savings opportunities for reducing the electrical bill and improving profits can be identified.4. Implementing Improvements. Having identified areas of potential savings, improvements can then be implemented in order of their significance. This could be a matter of simply improving operations, or in some eases may mean the use of an automatic control system.These four steps can result in significant dollar savings."

Jan 1, 1987

By H. Clare Horwood

"Fractures and fracture zones containing veins and stringers of auriferous quartz in a quartz porphyry dyke are the principal structural features at the Howey and Hasaga gold mines. The 'break' containing the dyke was produced by a force from the northwest, the fractures in the dyke by compressive stress from the west.Both mines are in the north-central part of Heyson township, in the Red Lake area. They lie close to the southeastern side of the Dome granite stock. The Howey mine exhausted its ore reserves and ceased operations in November 1941. The Hasaga mine is producing at the present time (June, 1946).Mineralization was first discovered in the dyke in 1925 by the Howey brothers. Lorne Howey's discovery, which became the Howey mine, came into production in April, 1930; Ray Howey's find, which is now the Hasaga mine, started production in November, 1938.The total production from the Howey mine was 421,324 oz. gold and 144,253 oz. silver from 5,158,376 tons of ore hoisted ( 4,319,331 tons milled) during its 11 years of operation.The total production from the Hasaga mine from November, 1938, to December 31, 1945, was 124,434 oz. gold and 51,009 oz. silver from a total of 901,098 tons hoisted (721,361 tons milled).GENERAL GEOLOGYThe porphyry dyke containing the orebodies at the Howey and Hasaga mines lies a short distance southeast of the contact of a granite-granodiorite mass known as the Dome stock. The dyke fills a fracture trending N.65°E. in a tongue of Keewatin volcanic breccia that extends in a rather irregular manner into an elongate intrusive stock of dioritic rock known as the Howey diorite. The accompanying table presents the geological succession."

Jan 1, 1954

By L. Flett, S. Woods, C. Deredin

"Following a significant reduction in world Imperial Smelting Furnace (ISF) capacity from 2003 to 2008, a major optimization initiative was successfully carried out on the lead upgrading and bulk (zinclead) flotation circuits in the Brunswick Concentrator.This paper describes the process improvements that have taken place in order to reduce the quantity of bulk concentrate:•,Increase in lead upgrading rougher flotation circuit capacity•,Installation of a closed-circuit bulk regrind utilizing MillPebs and 150 mm Cavex hydrocyclones•,Installation of two additional cleaner flotation stages on the bulk circuit•,Optimization of the manipulated variables in the lead upgrading and bulk circuitsINTRODUCTIONThe Brunswick Mine (Xstrata Zinc Canada) is located near Bathurst, New Brunswick and is one of the world’s largest lead/zinc mines. Milling of ore commenced in 1964 from the No. 12 orebody, with an expansion in 1966 to accommodate ore from the No. 6 open pit mine. The planned closure of the mine will take place at the end of the first quarter in 2013. The ore is a complex massive sulphide, consisting of approximately 20% valuable minerals, 60% pyrite and 20% silicates.A simplified flowsheet for the concentrator is given in Figure 1. The overall flowsheet involves SAG milling, secondary ball milling, selective flotation, and dewatering by thickening, filtering and drying. The complexity of the flowsheet evolves from the requirements for fine grinding and the production of four concentrates: zinc (sphalerite), lead (galena), copper (chalcopyrite) and bulk (sphalerite plus galena).Primary hydrocyclone overflow feeds the copper-lead aerator-roughers for flotation at pH 9.0 with soda ash. Plant xanthate (80-20 sodium isopropyl and potassium amyl) and AP407 are used as collectors in the copper-lead circuit. The rougher concentrate is directed to the copper-lead regrind circuit and the tailings are fed to the secondary grinding circuit. The secondary hydrocyclone overflow from the secondary grinding circuit feeds the copper-lead rougher-scavenger banks. Rougher concentrates are reground to a P80 of 28 µm prior to three stages of copper-lead cleaning. The first stage of cleaning is done in an open circuit with the tail reporting to the cleaner scavenger. The cleaner scavenger concentrate reports to the regrind circuit and the cleaner scavenger tail combines with the rougher/scavenger tail to feed the zinc circuit. The second and third stages of cleaning are performed in closed circuit with the cleaner tail being recycled to the copper-lead regrind circuit. The average final concentrate grades 23% Pb, 3% Cu and 704 g/t Ag, and is fed to the copper separation circuit."

Jan 1, 2013

As part of the Australian Mineral Industries Research Association (AMIRA) P9 project, methods of characterising both the ore and the machines in flotation circuits have matured considerably over the past 15 years. Machine characteristics include gas dispersion measurements, such as bubble size, gas hold-up and superficial gas velocity. Ores can be characterised in terms of floatability parameters on a size and mineral basis. Measurements have been performed on various copper plants in many countries around the world over the last 5 years. This paper outlines the cell and ore floatability characteristics of each copper plant and provides a discussion on the differences between these plants. In general, most porphyry copper ore bodies are processed at high tonnage through several parallel banks of cells at each stage of flotation. The floatability parameters of these copper ores are very similar in terms of rate of flotation and proportion of fast, slow and non-floating material. The cell characteristics are also similar, with superficial gas velocity, bubble sizes, and gas hold-up values being of similar magnitude. Investigation of the similarities and differences of the modem copper flotation circuits has allowed an increase in understanding of the particularities of copper flotation. Comparison against other copper circuits provides a benchmark and opportunities for future optimisation.

Jan 1, 2007

By J. George Gregr

"This paper gives a brief review of grinding mills installation procedures with special emphasis on future maintenance. The importance of cooperation between the customer, construction contractor and manufacturer&apos;s representative is discussed, as well as the imparlance of complying with standards, specifications and the keeping comprehensive installation records.Fundamentals of mill maintenance are explained with the aim being to maximize the equipment availability and to minimize unscheduled downtime. The paper also includes development of wear items design, based on operating conditions, drive components and bearings, and how to keep large gears healthy thus eliminating surprise failures.Experience with installation of a large gear on a 34-foot diameter mill in Lornex and successful application of a computer program for the gear alignment on the mill are given.IntroductionSuccessful operation of a concentrator begins with the engineering of the project and continues through selecting the right equipment, its installation , start-up period, and ends with final tuning of the process. machinery and the human factor into an efficient and viable industrial complex.The following is a brief review of grinding mill installation procedures with special emphasis on future maintenance. Fundamentals of mill maintenance are discussed with the aim being to maximize the equipment availability and to minimize unscheduled downtime, with some specific points on key grinding mill components such as bearings and drives.Finally, the experience is recounted with the installation of a large ring gear on a 34-foot diameter mill at Lornex, and the successful application of a computer program for the gear alignment on the mill."

Jan 1, 1985

"Mineral separation by exploiting the differences in mineral or metal specific gravities (sg) has been used since antiquity. In the modern era, this is most commonly seen in the recovery of gold by the use of centrifugal concentrators and the separation of coal (low sg) from waste rock (high sg) using heavy media cyclones and jigs. Other metals, notably oxides of tin and tungsten, are successfully recovered using some combination of jigs, spirals and shaking tables. However, many conventional gravity separation systems are limited to fairly coarse (>1 mm) to medium (>100 micron) particle separations due to the limitations of the gravitational separation forces in conventional equipment.Many of the high specific gravity minerals of interest are friable (e.g. cassiterite, wolframite, etc.) and excessive fines are generated in the grinding/comminution circuits which cannot be recovered with conventional equipment and may be lost directly to the tailings stream. Recent developments in centrifugal gravity separation equipment specifically designed for fine or ultra-fine particle separation and recovery have resulted in heavy mineral particle recoveries as fine as 10-20 microns. Ongoing research work has also identified the benefit of closed-circuit, multi-stage gravity recovery circuits using some combination of centrifugal concentrators and conventional equipment (tables, spirals, flotation, etc.) which can substantially increase overall plant recovery and final concentrate grade.This paper will provide details of the evolution of non-fluidized bed gravity concentrators for high sg mineral recoveries, correlate laboratory scale development work with plant results to confirm scale-up, and show how recent flowsheet developments have contributed to significantly improved plant performance.IntroductionGravity concentration is one of the oldest and most cost effective metal and/or mineral recovery systems, ranging from the classic hand panning of alluvial gold, through to the large scale production of a wide range of metals and minerals, including gold, silver, PGMs, black sands, coal and iron oxides. Conventional gravity recovery systems, comprised of some combination of sluices, spirals, jigs and shaking tables have limited application in fine particle (i.e. <100 microns) recovery, or where there are limited differences in particle specific gravities (sg), e.g. the separation of high sg, high value metal oxides (e.g. those of tin, tantalum or tungsten), from medium sg, lower value metal sulfides such as pyrite and pyrrhotite. These unit operations tend to have limited capacity, and while double and triple turn spirals and multi-deck tables have been developed, extensive floor space is often required in order to process the tonnages incorporated in the modern mineral processing plant."

Jan 1, 2017

By Charles S. Ney

"THE WRITER'S acquaintance with Kootenay King mine is limited to a brief geological excursion in the vicinity. This paper has been compiled largely from maps of J. Sullivan, mine geologist in 1952, kindly made available to the writer by Transcontinental Resources; a 1942 private report by W. L. Brown; and personal communications with J. Sullivan. Remarks on regional geology, and the comparison with Sullivan mine, are specifically the writer's.The Kootenay King mine is situated in the westernmost range of the southern Canadian Rockies at an elevation of 7,300 feet. From the Sullivan Mine at Kimberley, B. C., it lies 19 miles due east. Six miles of this distance are represented by the pronounced topographical and structural gap of the Rocky Mountain Trench. Mineral claims were locat-ed in the area prior to 1898. The deposit was first seriously explored in the late 1920s, and a small orebody was then outlined and developed. Mining and milling were undertaken in 1951.52. The orebody, amounting to about 14,000 tons, was mined out, and no significant new ore was revealed.The orebody is a replacement of dolomitic argillite by fine-grained galena, sphalerite, and pyrite. The sulphides are finely laminated in the direction of bedding, following out minor folds faithfully. The grade of ore estimated from diamond drill holes (W. L. Brown, 1942) was 11 ·I per cent zinc, 11 ·I per cent lead, and 3·8 ounces per ton of silver. The mineralized zone extends about 250 feet horizontally, 160 feet vertically, with a width of six to twenty feet.General GeologyThe country rocks are sediments of the Aldridge formation of late Precambrian age. For several thousand feet stratigraphically above and below the mine, the formation consists mainly of thinly laminated argillite and siltstone. Several horizons contain appreciable carbonate. About 2,800 feet stratigraphically below the mine, predominantly dolomitic beds mark the top of the underlying Fort Steele formation. The medium-bedded fine quartzites, which characterize about 10,000 feet of the Middle Aldridge in the Purcell range, have a restricted development amounting to only 2,000 feet in the Rocky Mountains. This member is about 2,200 feet above the mine horizon."

Jan 1, 1949

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 G. W. H. Norman

"The La Corne molybdenite-bearing veins, discovered in 1915, lie halfway between Amos and Val d'Or near the southwest corner of La Corne township, Abitibi county, Quebec. The Molybdenite Reduction Company, which carried on operations during 1927 to 1930, sank a 200-foot inclined shaft to develop the property and produced about 22 tons of concentrntes from a 30-ton mill during 1929. Hollinger Consolidated Gold Mines, Limited, carried on further underground exploration in 1930. The Molybdenite Corporation of Canada drilled a part of the property in 1938 to 1939 and produced 1 ton of concentrates in 1938. Wartime Metals operated the property from 1942 to July 15, 1945. At this time the mill was enlarged to treat 200 tons daily. Molybdenite Corporation of Canada since July 15th, 1945, have continued the operation of this property and have revised the mill flow-sheet in order to produce both a bismuth and a molybdenite concentrate.The production for the period 1942-1946 is as follows:Total molybdenite production from the mine in the 1942 to 1946 period was about 4,127,400 lb.The La Corne deposit occurs at the northwest contact of a small mass of biotite granite with biotite schist. The form and extent of the biotite granite mass is concealed by a wide sand plain that extends northwest and west from the deposit and overlies bed-rock except in four small outcrop areas. These are spaced 1,000 and 2,000 feet apart along a narrow strip extending 5,000 feet south from the mine. The outcrops suggest that the granite must be at least 1,500 feet wide and 3,000 feet long northeastward. Where exposed in the mine workings the contact dips 70°N.W., but is not a sharply defineable feature due to the presence of numerous granite dykes and tongues in a zone of biotite schist 100 feet wide along the margin of the granite. The biotite schist in the few outcrops that flank the north and south sides of the granite strikes northeast and dips 60° to 70° S.E."

Jan 1, 1954

By R. M. de Carvalho

Empirical and phenomenological models of SAG and AG mills have dominated the scene for the last few decades, being used nearly every time a mill is scale-up or an attempt is made to improve its performance. However, in spite of their success, these models fail to describe the details of both the grinding environment and the breakage mechanisms. The result is a generally good track record, with a few ? but very important ? instances when these models fall short in their predictions, which can be very costly. Further, nearly all of these models require detailed calibration to industrial and pilot plant data, being costly in their application, although still with limited predictive capabilities. Recognizing these challenges, a new model framework has been proposed which overcomes the limitations of currently used models by coupling the microscale population balance model, the discrete element model and functions that describe internal classification at the mill grate. This model framework has already been applied to describe breakage in ball mills, with showing very promising results. However, in the case of SAG and AG mills a much greater level of complexity is involved in the model, given the interaction between the mechanical environment that is formed totally or, in part, by the ore itself, and the breakage response in the mill. The paper describes the model framework, demonstrating how it can address appropriately effects such as mill speed, mill filling and size distribution, on the response from the mill. Some of the advantages of the model are related to its intrinsic multi-component and dynamic nature, which allows simulating appropriately mixtures of ores of different competences in a transient manner.

Jan 1, 2012

By R. G. Helberg, S. M. Rebelo

"Copper and cadmium are removed from the zinc sulphate solution in primary purification by cementation reaction followed by filtration. The solids are sent to the cake leach section where zinc and cadmium are recovered from the copper rich slurry. The copper solids from cake leach are filtered, bagged and sold. In order to achieve a lower cadmium copper cake, an altered process flow is suggested. It is proposed the primary purification be split into two steps for individual removal of copper and cadmium. INTRODUCTIONCopper cake is a product of primary purification, also known as copper and cadmium removal. Copper and cadmium are removed by cementation reaction in four stirred reaction tanks. The slurry is then filtered through Shriver manual plate and frame presses with the solution advancing to the next stage of purification, hot zinc purification (HZP). The solids from primary purification are further processed at the cake leach section where zinc and cadmium are separated from the copper rich slurry. Cadmium and zinc are dissolved in the cold cake leach tanks using spent electrolyte. Cadmium is recovered from this cadmium rich solution by cadmium sponging, leaching and electro-winning circuits. The residue from this cold leach step is further leached using spent electrolyte at a temperature of 85°C. The copper rich solids from hot cake leach are filtered on a drum filter and bagged in 1 ton bulk bags for shipping to buyers.Current assays of copper cake are shown in Table 1. The goal is to produce a copper cake product that is below 0.5% cadmium."

Jan 1, 2012

By Gilles Carrière

"The Suffield mine is situated five miles southwest of the city of Sherbrooke, in the Eastern Townships of Quebec. lt is the property of Suffield Metals Corporation Limited, a wholly owned subsidiary of Ascot Metals Corporation Limited.The old Suffield or Griffith mine had been worked intermittently from 1863 to 1906, without ever producing much ore. In 1951, diamond drilling done by Ascot Metals proved over one million tons of ore in three new bodies. The average grade of this ore was 0·018 ounces of gold per ton, 2-40 ounces of silver, l ·28 per cent copper, 0·59 per cent lead, and 6-45 per cent zinc. Since then the mine has been active, producing between 300 and 700 tons daily. More recent exploration has found continuations of the ore zones.General GeologyThe ore of the Suffield mine is found in porphyritic rhyolites of the Sherbrooke group (Silurian), at or close to its contact with sediments of the Beauceville group (Ordovician). The contact shows pronounced folds as well as irregularities due to erosion of the sediments prior to the outpour-ing of the lavas. On surface the general strike is north, and the dip is steeply east.The footwall sediments consist of interbedded black graphitic slates and. dark-coloured chloritic quartzite, overlain in many places by a thin bed of magnetite-hearing quartzite.The rhyolite host rock is a light grey rock, containing very abundant quartz and feldspar phenocrysts in a microcrystalline matrix of the same minerals. Some sericite is always present, especially in the vicinity of the ore zones, where the rock shows a pronounced schistosity.Several narrow basic dykes, with an easterly strike and vertical dip, cut all other formations. Most are lamprophyres, of either the monchiquite or camptonite type, but one has been identified as a diabase."

Jan 1, 1949

By Louis J. Cabri

"Mineralogical evaluation of ores containing platinum-group elements (PGE) is fraught with difficulties because of the low grades (often 1-3 g/t PGE), dispersal of PGE (usually as finegrained platinum-group minerals [PGM]), as well as occurring in solid solution in some base metal sulphides (BMS) such as pentlandite and pyrrhotite. Mineral processors not only need to know how the PGE are distributed, that is, their mineralogical distribution, but the grain sizes of the PGM, and especially the associated minerals.After obtaining a representative ground sample, common approaches include studying a large number of polished or polished thin sections, concentration of the PGM and sulphides by different gravity methods (e.g., heavy liquids, elutriation, micro-panning) and study of the produced concentrates. Unfortunately, these methods often do not produce concentrations that are sufficiently representative of a particular ore. The newly developed patented technology of hydroseparation (Rudashevsky et al., 2002) is especially effective at making concentrations of PGM over a wide range of concentrations (down to <300 ppb Pt+Pd+Au) and size fractions, including <45µm (down to a few micrometres in size), a size range not possible by other techniques. This water-based environmentally friendly laboratory technique provides a large amount of data from samples, which can range from 5 to 1000g, giving confidence to extrapolation of results to bench-scale and larger test work, especially when combined with quantitative image analysis to determine modal percent of major minerals and mineral liberation data, needed to predict the minimum (K80) and optimum (K40) grinds. INTRODUCTIONThe invention of the electron microprobe in the early 1960’s and scanning electron microscopes (SEM) with energy dispersive spectrometers (EDS) in the 1970’s presented mineralogists with powerful tools permitting them to obtain chemical and surface area data on minute amounts of minerals in situ, in a non-destructive manner. An area of process mineralogy that has very much benefited from this new technology was the characterization and understanding of PGM species, resulting in rapid growth of newly characterized species (Figure 1). Recognition of the first PGM may be ascribed to the Spanish explorer Ulloa in 1748, who described native platinum from South America. The second PGM (laurite, RuS2) was only recognised in 1866 by Wöhler in placer samples from Borneo (now Kalimantan) and the third (sperrylite, PtAs2) by Wells in 1889 from the Sudbury Ni-Cu sulphide deposits. The discovery of the important Pt-Pd deposits in the Bushveld Igneous Complex led to the characterization of the important Pt-Pd sulphides (cooperite, PtS2 in 1928 and braggite, (Pt,Pd)S2 in 1932). By the early 1960’s, only about 20 PGM were recognised, and many later proved to be poorly or incompletely characterized. Thus, the introduction of electron probe microanalysis (EPMA) led to a 50% increase of known PGM, as shown in a 1972 review on PGM (Cabri, 1972). At the time of the next review in 1981 (Cabri, 1981) the number of known PGM had grown to 75 and as of 2002, it had increased to 109 (Cabri, 2002). However, it is the author’s experience that less than about 25 PGM may be considered to be “common”, i.e., found in many different deposit types, as shown in Table 1, but usually only some of these occur in any one particular deposit."

Jan 1, 2004

By Geoffrey Gilbert

Rossland is in the West Kootenay district; 5 miles north of the International boundary and 4 miles west of Columbia River at Trail. The first claim in the district was located in 1887; but there was little activity until 1890; when the showings that later became important were staked and development began. Production increased rapidly from 1; 800 tons in 1894 to a peak of 360; 000 tons in 1903; and averaged 286; 000 tons for the next 13 years. It declined sharply in 1917; and continued to decline until the closing of the main mines in 1928. Between 1933 and 1941 leasers mined a further 137; 000 tons. Total output has been about 6; 200; 000 tons; with average recovery of 0.47 oz. gold; 0.6 oz. silver; and 1 per cent copper. The total value was $82; 500; 000; of which aboµt three-quarters was in gold; the camp still leads the province in total gold produced. Ninety-eight per cent of the tonnage came from a small compact area worked by four inter-connected mines : the Le Roi (39 per cent); Centre Star (25 per cent); War Eagle (24 per cent); and Josie (10 per cent); all now owned by the Consolidated Mining and Smdting Company of Canada; Limited.

Jan 1, 1954