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"The Pierre Beauchemin Mine attained commercial production in January 1989, and went on to produce 23 968 ounces of gold for the year, 3% higher than forecast. Underground development was accelerated to prepare the way for the production of an estimated 40 000 ounces in 1989.IntroductionThe Pierre Beauchemin Mine is located nine miles west of the city of Rouyn-Noranda (Fig. I).As a result of the amalgamation with Sullivan Mines in October 1987, the mine is now entirely owned by Cambior Inc.The Eldrich deposit was discovered in 1953and began production in 1955, closing in 1962 following the decrease in gold prices and the simultaneous decrease in silica content in the ore, which was sold to the Noranda smelter for flux. A total of 717 000 tons grading 0.14 oz Au/t were produced during this period.From July 1984 to October 1987, an underground exploration program was carried out by Sullivan on the Eldrich mining concession and the nearby Flavel property. The first phase of the program included reinstallation of surface facilities and dewatering of the Eldrich Mine. Following the completion of the first phase, extensive exploration was carried out, including a major program of diamond drilling. It was then decided to deepen the shaft to a depth of 1800 ft to access a new zone called the No.5 Zone .Following the amalgamation with Sullivan in October 1987, Cambior assumed the role of operator for the project, and continued the shaft deepening work that had already been started. A feasibility study was completed in April 1988 and the Cambior board of directors approved a production decision for the Eldrich-Flavel project.The preproduction program extended from May to December 1988 and was aimed at preparing the stopes in the upper part of the mine and at developing the mine infrastructure at the bottom of the mine.The mine was officially opened in August 1988. Cambior renamed the project the Pierre Beauchemin Mine in recognition of Pierre Beauchemin's pioneer-work."

Jan 1, 1990

By J. Y. Liu

In Comparison to the application of HPGR (high-pressure grinding rolls) in the cement industry and in iron ore processing, the introduction of this energy-efficient comminution technology to processing of nonferrous base-metal ores has progressed slowly as yet. In the field of minerals industry, challenges lies not only in the large variety of the feed material with respect to particle size, density, hardness, moisture and mineralogy of the material but also in complexity of the processing flowsheet, which need to be addressed from case to case based on result of HPGR comminution test work. Whereas conventional comminution test with HPGR in lab-scale or pilot scale needs quite a large amount of feed material, a compressive stressing test of particle-bed in piston-die device provides an effective way to study interparticle breakage behavior of ore material with much less feed. In the present study the effects of interparticle breakage on comminution of a copper ore has been investigated by compressive stressing of particle bed in a piston-die device. The feeds to the particle-bed stressing tests were four narrow size fractions in the size range of 0.3~3.35 mm and a broad size fraction of 0/3.35 mm. The pressures applied were in the range between 30 and 180 MPa. The results showed that the specific energy consumption increases linearly with the pressure; the product fineness increases also with the pressure, but the magnitude of this increase declines as the pressure goes up. Under the same stressing intensity, the product size distribution curve becomes steeper as the particle size decreases, so there is no self-similarity between product size distributions of different feed sizes. The suitability of a material-characterization model for interparticle breakage was verified for this ore. In addition, the effect of interparticle breakage on comminution of this ore with respect to its influence on the downstream flotation were analyzed in terms of size-metal distribution and mineral liberation.

Jan 1, 2016

By T. Yalcin

"This paper describes the development of a process flowsheet designed to provide high dewatering rates and to bring about savings in tailing disposal costs. The flowsheet comprises a hydrocyclone, a settling cone and a thickener arranged in closed circuit. The mill tailing, after pulp density adjustment, is fed to the hydrocyclone which yields two products. The underflow product is sent to the settling cone for further dewatering. The overflow product, together with the settling cone overflow, goes to the thickener which, with the aid of flocculants, gives a clear effluent and a relatively dilute underflow. The latter combines with the original mill tailing to form the hydrocyclone feed. Solids finally leave the system in the settling cone underflow which can reach pulp densities of up to 78%, making it suitable for direct disposal or for use as mine backfill.Some discussion is included for the selection and sizing of equipment required for this process. INTRODUCTIONDewatering is an essential process for the disposal of mill tailings which forms an added burden in milling practice, often with little or no return. It is, therefore, desirable to conduct this process as economically as possible. One way of reducing costs would be to increase the rate of dewatering so that savings can be made in the size of dewatering unit s which would lead to lower capital and operating expenditures.A common method of dewatering mill tailings is gravity sedimentation which is frequently carried out with the aid of flocculation. In most cases, flocculation greatly enhances the sedimentation process, but only until a pulp density of 50% to 55% is reached in the settled material.Centrifugal sedimentation in hydrocyclones can provide high pulp densities rapidly, but does not result in a clear effluent.This paper outlines a suitable combination of gravity and centrifugal sedimentations that results in an over-all high rate of dewatering. The system developed produces a material of high pulp density while at the same time giving clear effluent."

Jan 1, 1989

By G. R. Tostengard

"The paper identifies functional areas which frequently can yield significant productivity improvement. Emphasis is placed on the repair and maintenance, material supply and training functions. Successful approaches are discussed in some detail, and typical results which have been achieved are reviewed.IntroductionA challenge which at all times confronts the manager o f any mineral operation is the potential to improve his operation no matter how effective it might be a t the present time. The challenge has been present through my entire lifetime and I suspect a good many years prior to that and, I expect, will continue many year s and decades in the future. The competitive and economic situation in which we in the minerals industry find ourselves today makes the challenge eve n more important during the 1980s. In many cases it’s a question of survival, not simply pr of it improvement.There are, of course, many avenues to improving profitability; technological changes, new equipment and facilities, improved products, improved mine plans, etc. All of these are important and must be addressed by each manager. This paper focusses on just one avenue to improvement and that is ""improved management"", not so much of management at the boardroom level but of improved management o f the facilities and workforce as the y currently exist. Management implies control. There are three areas where control is often found lacking:1, Repair and maintenance of equipment and facilities;2. Material supply including procurement, inventory control and warehousing; and3. Training.These three areas are all complex and difficult to manage and they are all but ignored in the formal education which our industry's manager s normally receive. They are areas which can be improved with relatively small investments and therefore, they represent especially attractive improvement targets during the 80s when, for most of us, investment dollars are very difficult to come by."

Jan 1, 1983

By W. Petruk, D. R. Morris

"The lead concentrate produced from the volcanogenic base metal ore of Brunswick Mining and Smelting in New Brunswick contains major to trace amounts of iron, zinc, cadmium, arsenic, antimony, copper, silver, selenium, tellurium, bismuth, sulphur, tin, indium, mercury and manganese as impurities. When the concentrate is sintered some of these impurities enter a variety of oxide phases, whereas others precipitate as metals. In the lead blast furnace the oxide phases decompose and the impurities recombine 10 form new phases, most are eventually discharged to the slag. The metals of the sinter remain as metals in the lead blast furnace where even more are produced, and are discharged largely with the lead bullion,IntroductionMany lead concentrates are produced from volcanogenic base metal ores . Such ores contain a wide variety of trace minerals and trace elements , therefore, concentrate s obtained from them will contain the trace elements as impurities . The impurities in the lead concentrate from one such ore, the Brunswick No. 12 deposit of Brunswick Mining and Smelting Corporation (BM S) in the Bathurst area in New Brunswick , were studied as pan of a larger investigation that was conducted to determine the extent of chemical reactions of the sinter within the lead blast furnace(1,2) The lead concentrate was investigated to identify the host mineral s for the impurities, the lead sinter was studied to identify the host phases, and frozen products from different depths within the BMS blast furnace at Belledune, New Brunswick were studied to determine the behaviour of the impurities.T he investigation was performed by pre pa ring polished sections o f the samples and studying them with a reflecting microscope, a scanning electron microscope (SE M) equipped with a Kevex energy dispersive X-ra y analysis system, a microprobe, and a Quantimet 720 image analyzer. The samples were also analyzed by X-ray diffractometer, and individual grains were analyzed by the Debye-Scherrer X-ray powder film method."

Jan 1, 1985

By Frank D. Shepherd

THE Gunnar Gold Mines property is on the west shore of Beresford lake, in the Rice Lake mining division, southeastern Manitoba. Gold was first discovered in place in the Rice-Beresford Lakes area in 1911 on the Gabrielle claim, now included in the San Antonio Mines property at Rice lake. Following an eastward trend of favourable prospecting ground, promising showings were next staked in the vicinity of Long lake during the years 1915 and 1916. At that time, William Walton and associates, among them Gilbert LaBine, staked the Eldorado vein, north of Long lake, and Eldorado Gold Mines, Limited, was formed to develop the property a Company that, later, was to become widely known through successful operation of its radium-silver mine at Great Bear lake, Northwest Territories. Continuing eastward from Long lake, Walton and his partners, Gunnar Berg, Leo Seaberg, and Wm. Quesnel, staked claims during the years 1920-24 in the vicinity of Beresford lake which, later, were to become incorporated into the Gunnar Gold Mines property. The presence of gold-bearing quartz on these claims had been known from the time they were staked, but an opportunity to consolidate them into a group was not forth-coming until early in the year 1933. Although promising gold showings had been disclosed during the performance of assessment work on individual claims, no continuous mineralization having mine-making possibilities was discovered until after the consolidation of the entire group. From this stage onward, progress was rapid. Active exploration was begun on August 1st, 1933, and late in October James Houston reported favourably on the property. On October 27th, Gunnar Gold Mines, Limited, was incorporated in the Province of Ontario to develop and operate the property. The Company is also registered in Manitoba. The authorized capitalization is 3,000,000 shares of $1.00 par value. Gilbert LaBine was appointed president of the Company, with James Houston as mine manager. Following completion of a diamond-drilling programme, planned in 1933, it was decided to install a mine plant capable of carrying development to a depth of 1,000 feet. A transmission line, 5 miles in length, was built to the Great Falls-Central Manitoba line of the Manitoba Power Company, Limited.

Jan 1, 1939

By A. S. Robb

SAMPLING of an ore-body is one of the fundamentals in its exploration, development, and exploitation. It is carried out to determine the gold value per ton of an ore-shoot and the probable gold output of the refinery. Sampling also places any broken ground in one of two classes, ore or waste. The value and importance of correct sampling has been shown repeatedly in the past. On the other hand, instances could be given where those in charge of mines have not believed in sampling, and the result has been the same in every case-disaster. This has in all probability been due to assuming that good-looking, well-mineralized rock was ore. Sampling in some form might have averted this misfortune. In this essay, only the method of sampling and the estimation of ore reserves is treated. All titles, figures, and values are fictitious. None of the illustrations can be directly applied to any portion of the Hollinger mine, except Figures 1 and 2, which are included with the special permission of the Canadian Mining Journal. METHODS OF SAMPLING Hollinger Consolidated Gold Mines, Limited, has, from the very beginning of operations, carried out an extensive system of sampling, and the results of this sampling have completely governed the whole mining programme. The Hollinger discovery was made in 1909 and prospecting and exploration of 'the surface began at once. Four men were given jobs as samplers and they sampled thoroughly all the outcrops and all bed-rock which could be exposed to view by removing the overburden. When shaft sinking began, samples were taken at two-foot intervals. In 1912, an underground gang of six samplers was organized, and the number was increased at a rate of about two men per year as the size and extent of the workings grew. No permanent record of sampling was kept until 1919, in which year an average of 11.6 men took 28,458 samples, cutting 113,894 feet of sample. From that time until the end of 1934, the average number of samplers varied, but the total number of samples cut, and the total footage sampled, increased.

Jan 1, 1937

By R. N. W. DiLabio

"In a continuing program of mapping glacial dispersal trains, near-surface till samples were collected down-ice from a copper-silver-gold prospect at Onaman River, Ontario, and down-ice from the past-producing gold district at Oldham, Nova Scotia. The program is designed to aid exploration by illustrating glacial dispersal from various types of known occurrences and by identifying postglacial processes that control drift composition at a detailed scale.Native gold was mined at Oldham from arsenopyrite-bearing saddle reef quartz veins in a dome in Early Paleozoic slates of the Meguma Group. Gold abundances greater than 10 ppb in till form a dispersal train at least 1900 m long. Gold abundances between 40 and 100 ppb occur in clusters up to 600 m down-ice from gold-bearing bedrock. Anomalous arsenic abundances in till form a broad train 1100 m long that is significantly different in shape from the gold train.At Onaman River, subeconomic amounts of Cu-Ag-Au occur in quartz lenses in Archean mafic volcanic rocks. A dispersal train defined by the distribution of mineralized boulders and increased abundances of Cu, Ag and Zn in till extends 600 m down-ice from mineralized bedrock. Gold abundances between 10 and 70 ppb do not define this dispersal train, but occur instead in clusters 300 to 600 m down-ice from mineralized bedrock. In vertical sections in the upper 2 m of copper-rich till, Au abundances vary sympathetically with Ag and Cu levels; however, there seems to be no preferential site of Au enrichment in the mineral portion of the soil profile.Several of the mineralized boulders at Onaman River are richer in Au and Ag than the mineralized bedrock from which they were derived. Although some of the high Au and Ag assays were caused by biased sampling of sulphide-rich portions of the boulders, many were caused by leaching of sulphide minerals from 'the boulders, which left a limonitic boxwork that is enriched in Au and Ag. A few of the Au-rich boulders may have been derived from an unknown bedrock source."

Jan 1, 1982

By A. Hasselbring

Sub-Level stoping was used in this country at least as early as 1912, at the Magpie mine in Michipicoten. The fundamental principles of the system are the same wherever it is used, but modifications have been adopted in the Michigan iron mines and in several Canadian mines. It is a system whereby a suitable body of ore may be developed quickly and worked cheaply, with a low initial expense, and by miners of quite ordinary ability. The basic method, described in the following pages, may be modified to meet conditions where one or more factors are not in accord with normal sub-level stoping requirements. In suitable ground, sub-level stoping is probably the most satisfactory method yet devised for underground hard-rock mining. It is so simple, straightforward, and safe that no special training of a crew is necessary. Any miner capable of drifting and raising can carry on the work necessary for sub-level stoping after the shaft has been sunk. The method is suitable, however, only for reasonably wide ore-bodies, the minimum width of ore to which it can be economically applied being somewhere in the neighbour-hood of thirty feet. The development programme is simple. Assuming, for purposes of illuslustration, an ideal ore-body, two thousand feet long, dipping nearly vertical, and varying in width from forty to a hundred feet, this programme would be as follows. The shaft should be sunk on the footwall side about sixty feet outside of the ore-body and near its centre. Levels are then established at suitable intervals, usually 125 feet. On each level, a cross-cut is run to the ore-body, and there a main-haulage drift is driven in either direction to the end of the ore. Cross-cuts are then run at 25-foot intervals, so that ore-chutes can be spaced about 25 feet centre to centre each way. It is not necessary that these cross-cuts be extended to the walls of the ore-body, if the position of the walls is known; they are extended just far enough to make room for the correct location of the chutes.

Jan 1, 1934

By J. G. Ross

A DESCRIPTION of the block-caving system of mining as practised at the King mine of Asbestos Corporation, Limited, Thetford Mines, Que., was presented. at the Annual Meeting of the Institute held in Quebec city, in April, 1934. The paper was published in the Transactions of the Institute for 1934. There has now been recovered, by this method, one million tons of ore, and a review of experience indicates that the method was well planned and has been executed smoothly. In approaching the problem of selecting the type of mining best adapted to this deposit, the method had to be one which would be low in cost, use no wood, and not disturb the surface immediately bounding the ore-body on two sides. Prior to 1931, block-caving had been considered but rejected, as it en-tailed disturbance of the surface area which lay within a 45° angle of the lowest contemplated workings, if the practice of such mines as Ray, Miami, Inspiration, and Morency, where this type of mining has been used extensively, was followed. At these properties there have been mined large tannages at outstanding low costs, but an extensive surface area has been disturbed in each case. In some instances, townsites in the mining area have had to be abandoned, and dwellings, not considered worth moving, have been allowed to fall into the depression. A study was made of the principles involved in the various types of block-caving, the existing information on ground movement, subsidence, etc., also being carefully analysed. Toward the end of 1931, a plan was conceived whereby the country adjoining ground being caved might be held permanently in place and undisturbed. The success of such a scheme was essential to the application of block-caving at the King mine on account of property boundaries, the Quebec Central Railway right-of-way, and expense of moving buildings. Study of existing data on caving and movement of rock, caused by sub-surface excavations of sufficient unsupported area to result in failure of the backs, indicated, in practically every instance, that such movements followed certain definite laws of mechanics.

Jan 1, 1936

By Jerry Ran

When multiple levels are mined in open stope mining with delayed backfill, mining underneath fill is inevitable to achieve higher rate of recovery if no sill pillars are left for the purpose of fill separation or partial ground support. This is particularly true when the orebody has low reserves per unit vertical distance. Instability of the exposed fill back causes stope dilution, production delay and possible equipment damage and consequently increases the mining cost. Previous studies have concentrated on fill mat stability in cut-and-fill and stability of fill walls in open stope mining. Workers are exposed to the fill back in cut-and-fill and therefore smaller open spans are normally adopted and the fill mat is reinforced. In open stope mining openings with large spans are designed, workers are not exposed to the fill back and its stability has to be maintained by fill of sufficient strength. However, the required fill strength is normally determined by considering different modes of fill wall failure. The fill back stability is usually not analyzed in the fill design process. The lack of analysis on fill back stability for open stope mining is mainly due to difficulties in relating the fill properties to the fill and stope geometry. In this paper a simplified wedge model is adopted to analyze stability of the exposed fill back in open stope mining with consolidated backfill. The proposed method calculates the factor of safety of the fill back against shear failure under known fill and stope geometry and fill strength parameters in order to assess the fill back stability. It can also be used to determine the required fill strength with known fill and stope geometry. The proposed approach is indirectly validated by mining experience and observations at two operating mines. In addition there are many factors affecting the fill quality and stope configurations in an operating mine. These factors have to be considered in the analysis of fill back stability. This situation is highlighted by presenting an actual mining case where the fill back failed due to factors commonly encountered in mining operations.

May 1, 2003

By K. J. Aglias, M. Jeffery, J. V. Simmons, J. A. Johnston, S. G. Fityus

"Shales and mudrocks undergo weathering in the near-surface geological environment, and degradation when directly exposed at the surface. Despite having superficially similar appearances, many shales have very different degradation behaviours. For example, some shales break down readily upon exposure whilst others remain more or less intact. Further, for those shales which break down, the style of degradation can vary greatly. A phenomenon is described where fresh shale blocks that are exposed at the surface undergo rapid breakdown to a fragmented texture which then persists for relatively long times, without further breakdown to finer soil particles. This is illustrated for some coal mine waste rocks. Furthermore, this metastable fragmented texture seems to be particular to each particular type of shale. The paper presents a conceptual model for this physical breakdown of shales which accounts for the differences in fragmental texture that are observed. The importance of several factors including strength, permeability and expansive clay content are identified for their contribution to the degradation behaviour of different shales.INTRODUCTIONShales and MudrocksThe term shale is applied widely to a class of mudrocks having a range of distinctive but gradational appearances and mechanical behaviours. Mudrocks are sedimentary rocks primarily composed of silt and clay sized particles (Tucker 1982). Mudrocks form in low to very low energy environments such as flood plains, lakes and deltaic environments, where the finest detrital particles are able to settle out of suspension (Reeves et al., 2006; Stow, 2005; Tucker, 2001).Unlike coarser siliciclastic sedimentary rocks, mudrocks do not present with a wide variety of sedimentary structures and textures. The most common texture is a preferred orientation of clays and micas parallel to the bedding plane. This textural component is related to the property of fissility or the ability for such textured mudrocks to split along smooth planes that are parallel to the bedding. Another common structure is lamination, formed by variations of grain sizes and types during deposition, Massive texture describes the situation where no sedimentary structures or textures are apparent (Boggs Jr., 2009; Tucker, 2001). The term shale varies between sources and contexts, with Reeves et al. (2006) stating that a shale is a fissile mudrock, where “fissility is a property that has developed at a certain stage of weathering”."

Jan 1, 2015

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 R. G. Achter, E. Wilson

"This paper describes the work undertaken to renovate the Pipe Open Pit drainage system, incorporating measures to improve effluent quality. It describes the technical feat of drilling an l l-inch-diameter hole from the bottom of the open pit for a distance of 1100feet to intersect an underground drift. An innovative method of checking the inclination of this hole is also described. The paper outlines the construction required on the 1500-foot level of the underground mine to facilitate water impoundment and slime settling, as well as the instrumentation for control of water flows. IntroductionThe Manitoba Division of Inco Metals Company operates two mines-an underground mine at Thompson and an open pit at Pipe. The Pipe Open Pit is located 20 miles southwest of the City of Thompson, approximately 400 miles due north of Win nipeg(Fig. 1).Surface diamond drilling of the Pipe ore body was sta rted in 1959. In 1961, a shaft was sunk to a depth o f 1,570 feet to carry out further exploration work and obtain bulk ore samples for metallurgical testing. Site preparation to mine the upper portion of the orebody as an open pit was initiated in 1966 and by late 1967, the removal of silt and clay overburden was in progress. The contractor completed the dredging in the fall of 1969. Over 12 million cubic yards of muskeg, clay and silt was removed, which in some places was 150 feet deep over the orebody. Production began in early 1971, by which time the top three benches of rock had been excavated and the orebody exposed.The Pipe No. 2 Mine main shaft was started in 1970 and completed in 1973 at a depth of 3,160 feet. Development for exploration purposes was carried out on the 1500 and 2460 levels, and permanent pumping facilities were installed.The surface dimensions of the open pit are 2,200 by 1,600 feet, with the long axis striking northeast. The walls are designed with a 45- to 50-degree slope and the ultimate design depth is 720 feet. The present depth of the pit is 520 feet and production capacity is 6,800 tons of ore per mine day."

Jan 1, 1980

By C. D. Smith

Numerous studies have shown the potential of air rate profiling to significantly improve the flotation performance of a bank of cells. To date many profiles have been implemented; balanced, increasing, decreasing and combinations of these. Previous work has typically shown using an increasing air rate profile, where the air rates are increased proportionally to each cell down the bank, yields a higher grade for a given recovery compared to other profiles. Air rate profiling can then be thought of as having two distinct components; determining the correct distribution of a given volume of air and, perhaps more importantly determining the correct total volume of air to add to the bank. To date no method for this second component has been established. In an alternative technique the air rate to each cell is adjusted in order to optimise the air recovery from that cell. Air recovery is the fraction of the inlet air which overflows the weir of a cell as unburst bubbles. The peak in air recovery (PAR) is a robust measurement that indicates optimal performance for the operating conditions in the cell, and sets the PAR air rate. By doing this for each cell in the bank, the optimal total air rate to the bank and the optimal distribution of that air can be simultaneously determined. In this study plant trials evaluated different air rate profiles, and profiling based on PAR. The total air rates to the bank for all profiles tested were approximately equal, allowing direct comparison. The results showed that the highest cumulative mineral recovery was achieved by operating at air rates which gave the PAR for each cell down the bank. Different distributions of the same total air rate to the bank gave the same cumulative upgrade ratio, but lower cumulative recoveries. In this study, the increase in cumulative mineral recovery, at the same upgrade ratio, was more than 33%. Peak in air recovery (PAR) based optimisation of air rates to banks of cells and to the individual cells has been shown to significantly improve flotation performance as it simultaneously determines the optimal air rate to the bank, and the optimal distribution of that air.

Jan 1, 2009

By Lin Li, Ahmad Ghahremaninezhad

"In this paper, the electrochemical behavior of ferric and ferrous couple on the pyrite surface was studied by application of the chronoamperometry technique in an acidic solution at room temperatures and atmospheric pressure. The effects of the pyrite origin, ferric and ferrous ions concentration on the electron transfer kinetics of Fe3+/Fe2+ reaction were probed. Tafel behavior of all the test conditions were developed. The results showed that the ferric-ferrous reaction on pyrite electrodes with different origins have a common mechanism, despite different reaction kinetics. The exchange current densities of the redox reaction have the same magnitude of about 10-5 A·cm-2, which is far lower than the value reported for the gold and platinum electrodes, but larger than that of the chalcopyrite electrode. The cathodic transfer coefficients are between 0.4 and 0.5, while the anodic transfer coefficients appear unusual with the lowest value being 0.23. It was also demonstrated that an increase of one order of magnitude of iron concentration has no clear effect on the exchange current density and cathodic transfer coefficients of Fe3+/Fe2+ redox reaction.INTRODUCTIONPyrite, FeS2, is mostly considered as a gangue mineral which frequently coexists with valuable minerals and metals, such as gold, copper and nickel. Its separation from valuable minerals can be a costly process (Chandra & Gerson, 2010). In refractory gold ores, the fine gold particles are locked inside the pyrite structure. Oxidative leaching of pyrite is a hydrometallurgical method to destroy the structure of pyrite and expose the fine gold particles to the cyanide solution in the downstream cyanidation circuit. Pyrite oxidative leaching can be achieved by oxidation of pyrite via oxygen and ferric ion(McKibben & Barnes, 1986). In fact, pyrite oxidation by oxygen and ferric is a significant source of acid mine drainage(AMD), which adversely causes serious environmental problems(V. P. Evangelou, 1995).The pyrite oxidation mechanism is accomplished in three consecutive steps: (i) pyrite is oxidized by oxygen and ferrous ion formation (Eq. 1); (ii) ferrous oxidation by oxygen and ferric ion formation (Eq. 2); (iii) pyrite oxidation by means of ferric ions (Eq. 3). Hence, both oxygen and ferric participate in the pyrite oxidation process (Moses & Herman, 1991):"

Jan 1, 2016

By Roderick C. McDonald

Great Bear lake, the fourth largest body of water on this continent, comprises an area of approximately 12,000 square miles, and is set astride the Arctic circle, with a large portion of it witnessing both the Arctic winter night, and the Arctic summer day with its midnight sun. The very name itself, Great Bear, suggests an Arctic region. The seasonal contrasts at Great Bear lake are very marked by the cold and darkness of winter, and the warmth and brightness of summer. Signs of winter such as low clouds, snow, sleet, fog, and night frosts, generally appear the latter part of August, while ice forming on bays during the last two weeks in September ends boat travel on the lake about September 30th.Early in June the ice leaves the river mouths, during the last ten days of the month it disappears from small bays, and at some time during the last two weeks of July the main body of the lake becomes navigable. In 1916, the ice disappeared earlier than usual, and the lake was navigable on July 5th. Great Bear river generally breaks up about June 1st, but, on account of ice cliffs and floes, it is not navigable with safety until early in July. Even then, the traveller is confronted with the Charles rapids, but these might be overcome by a wagon or motor road, such as the one from Fitzgerald to Fort Smith: However, in 1931, Mr. A. W. Boland, Fort Norman, N.W.T., carried with scow and motorboat practically all freight between Fort Norman and LaBine point and Hunter bay, at the east end of the lake. During July and August, two types of clothing are required, for, although the days are mainly ones of warm sunshine, the nights are cloudy and cold. In the winter season, fur clothing is used exclusively. Considerable timber, mainly white spruce, suitable for fuel and building purposes, borders the lake and extends eastward to the Coppermine river. North of the lake, no timber is found except in the river valleys, but with oil and coal possibilities near Fort Norman, the fuel situation might be solved.

Jan 1, 1932

By Charles A. Schacht

"Relining, lining maintenance, and ladle repair can represent a significant cost in steelmaking. Therefore, the objective of all steelmaking shops is to minimize this cost and to maximize the lining life. However, the thermomechanical behaviour of the ladle lining is highly complicated and sometimes misunderstood. The ladle design also has a significant influence on lining life. Other factors which have a direct influence on lining life are the shop lining installation practices and manner in which the ladle is used. The objective of this paper is to provide an understanding of the impact of these factors as related to improving the integrity of the lining and increasing lining life.Fundamentals of Steelmaking Ladle Lining Expansion ForcesNumerous studies have been conducted by the author on steelmaking ladles with molten metal capacities ranging from 110 ton s to 350 tons. The steelmaking ladle typically has a sidewall plate with a slight taper and for analytical purposes can be classified as a cylindrical vessel.A typical ladle is described in Figure I. The ladle sidewall lining usually consists of a working lining and a safety lining. Although the type of refractory used in the working lining varies based on the type of steel produced and other metallurgical requirements, the refractories typically used are: 70% alumina brick; direct bond mag-chrome brick; resin bond dolomite brick; and direct bond dolomite brick. Fireclay brick, high alumina brick, or a similar strength brick is used as the safety lining.Normally, the ladle lining is exposed to a through-thickness temperature gradient and restrained by the ladle sidewall plate, is described in Figure 2. The hot face side of the refractory lining develops a compressive thermal force and the ladle sidewall plate develops an equilibrating tensile force. The radial joints open on the cold side of the working lining as the lining joints have little or no tensile strength. A radial compressive load (P) develops between the working lining and the vessel shell due to the thermal restraint imposed on the lining by the shell. The radial compressive load is transferred from the working lining through the safety lining onto the ladle sidewall. Therefore, the safety lining must have the strength to accommodate the radial compressive load (P)."

Jan 1, 1991

By Euler De Souza, Kyle Penner

"Booster fans play an important role in assisting main fans by boosting the air pressure of the ventilation air passing through it. The correct placement of underground booster fans in a mine is a critical decision since it not only influences the mine’s operational capacity in terms of costs and efficiency, but also affects safety and environmental conditions. There are many booster fan installations around mines that are inappropriate, in some cases resulting in system failure (fan stall) or in a major negative impact on mine production. Inadequate booster fan placement has resulted, in some cases, in extensive air leakage which has compromised the economics of the operation and, in other cases, in recirculation of contaminants which has seriously compromised safety. This paper provides guidelines for proper design and siting of booster fans underground, to not only achieve an economic ventilation system but to create safe work conditions and efficient atmospheric conditions. Procedures for the planning, installation, operation and maintenance of booster fans are also presented. A case study illustrating the design and siting of a major booster fan installation in a potash mine, supported by computational fluid dynamics analysis, is presented.INTRODUCTIONIn large underground operations with complex ventilation networks offering high resistances, the never ending increased required quantities of air can almost only be supplied through use of high pressure main fan systems. These fans may, in effect, induce significant leakage losses in the system, having a negative impact on safety performance. In such cases, booster fans can be designed to decrease the main fan pressures and reduce the leakage flows, improve system efficiency and reduce costs. In well designed systems, booster fans can thus be used to redistribute the airflow and assist the main surface fans.When the pressure supplied by the main fans is insufficient to meet the ventilation circuit pressure requirements, booster fans are placed in strategic locations where pressure boosting is essential. Booster fans are typically sited in sections of the circuit offering high ventilation resistance and requiring this form of ventilating air motivation. Booster fans are thus required when the mine expands, and so the resistance of the ventilation network, and the primary fans no longer can supply the required pressure and flow."

Jan 1, 2019

By R. E. Bou Serhal

Many miners are exposed to dangerously high levels of noise on a daily basis. Over the past 15 years, a continually increasing number of occupational hearing loss has been reported from the mining community in the United States, of which, more than 95% is attributed to prolonged noise exposure. Although the noise exposure levels may differ between coal miners and metal and non-metal miners, in the absence of noise control at the source, the solution is the same: use of personal hearing protection devices (HPD). While protecting the miners? hearing it is also essential to no longer hinder their ability to communicate. With access to an advanced HPD that is customized to the miners? ears we are able to combine these two requirements. Using an intra-aural instantly custom molded HPD miners are protected from high levels of noise. The HPD is equipped with wireless capabilities, and contains both a speaker and an In-Ear Microphone (IEM). Therefore, the miners? speech may be captured from inside the ear and transmitted to the remote listener. This IEM signal is relatively noise-free since it is isolated from the background noise. The IEM speech signal, however, is ?boomy? and is missing some high frequency content, making fricative consonants hard to understand. Nonetheless, the IEM speech signal is correlated with the natural speech signal and may be manipulated through statistical techniques to more closely resemble natural speech. By improving the intelligibility and quality of the IEM signal, numerous applications may be enabled. One use of the enhanced IEM signal will be for radio communication. Using wireless radio communication in a noisy mining environment is sometimes the only practical and affordable solution to allow communication between miners equipped with personal hearing protection devices. Traditionally, one of the weaknesses of such wireless radio communication lies in the lack of designating receivers: whether they are the intended receiver or not, all those carrying a radio receiver are subjected to the broadcasted signal. The current work will detail a new concept of a "radio-acoustical virtual environment" where the radio signal will only be received by miners within a given spatial range, such range depending on the user?s vocal effort as well as the ambient and perceived background noise levels.

Aug 1, 2013