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By John Tribe

"This paper gives the organizational history of Luscar Ltd. and its various coal producing operations. Each 0/ the operations is then described individually. They include the Bienfait Mine, the Boundary Dam Mine, the Diplomat Mine, the Luscar Mine and the Coal Valley Mine. The company's plans/or expansion are also discussed. IntroductionLuscar Ltd. is an Edmonton-based energy company formed in 1967 through an amalgamation of two pioneer coal companies-Mountain Park Collieries and Luscar Coals Ltd. In 1912, Mountain Park Collieries initiated the development of an underground mine in the Coal Branch area of Alberta. This mine necessitated the establishment of a townsite and the provision of rail access. The second operation, Luscar Coals Ltd., was established in 1921 at Luscar, Alberta. This was also an underground operation and required the development of a rail spur and townsite. The product from both mines was used to fuel railway steam locomotives until the 1950's. The market demand for coal decreased as the use of the diesel locomotive increased, so that these operations were finally closed in 1954. Through the amalgamation of the two companies in 1967, however, Luscar Ltd. was formed to become an active participant in the revival of coal operations. Western Canadian Luscar Ltd. is involved in the exploration, development, production and marketing of coal, coal products, oil and gas. Luscar Ltd. is currently participating in five producing mines (Fig. I). The Diplomat Mine, approximately 125 miles southeast of Edmonton, was in production when it was acquired in 1949 by Luscar Coals Ltd. In 1966, the Bienfait Mine and char plant were added. The Luscar Mine, which is a 50/50 joint venture of Luscar Ltd. and the Consolidation Coal Company of Canada, and is operated by Cardinal River Coals Limited, started production in 1969 on the original Luscar Coals Ltd. mine site. The Boundary Dam Mine was opened in 1974. Both the Bienfait and Boundary Dam operations are located in the lignite coal fields of southeastern Saskatchewan. The most recent addition to the group is the Coal Valley Mine, 40 miles southwest of Edson, Alberta. It commenced commercial production in July, 1978 under the operation of Luscar Stereo (1977) Limited."

Jan 1, 1980

By C. Chapman, P. Cowx

"INTRODUCTIONTetronics Research and Development Co. Limited has installed several plasma torches for steel heating in continuous caster tundishes. An extensive demonstration program on a 14 tonne tundish reported the following advantages of plasma heating.• close control of steel temperature to ± 5°C;• a reduction of the furnace tapping temperature of up to 20ºC, leading to savings in energy and refractories costs ;• improvement in steel quality, to produce a slab product with an equi-axed crystal structure and a reduced number of large sub-surface slag inclusions;• the virtual elimination of oxygen pickup in the tundish; and• the capability of making ferroalloy trim additions to the tundish to permit a change of steel grade during casting, greatly decreasing inventory costs.The Need for Tundish HeatingIn the continuous casting of steel, the function of the tundish is to transfer the molten metal between the ladle to one or more continuously fed casting moulds. Figure I shows a typical temperature profile of molten steel in the tundish over the first cast cycle. It is observed that a rapid drop in temperature occurs both at the start of the cast, due to the tundish refractories absorbing heat, and also during ladle changeover. In addition, as a result of thermal losses through the ladle sidewalls, there is a gradual decline in the bulk steel temperature over the entire heat cycle. Consequently, in normal tundish operation, it is necessary to cast with a high superheat (25°C to 30°C + above steel liquidus temperature) in order to prevent freeze-up of the strand. Problems of high temperature variation and the potential of the steel solidifying in the tundish are greater for small caster units of low thermal mass . A high steel superheat increases the steelmaking vessel process time and refractory wear and prevents the formation of a uniform equi-axed solidification structure in the cast steel, resulting in centre line segregation and porosity. The large drop in the casting temperature that occurs during unsteady state operation causes an increase in the number of sub-surface inclusions, reducing the value of the product. Accordingly, plasma heating may be employed to maintain the molten steel at an optimum temperature to produce a slab of consistent quality."

Jan 1, 1989

By Howard J. Handewith

"Tunnel boring machines have been tried in underground mining applications as early as 1957. Although tunneling machines hardly threaten to supplant traditional drill-blast-muck operations for drifting, the latest modern machines offer substantial improvements in progress rates over traditional methods under certain conditions.An evaluation guide is offered which identifies conditions that favour machine boring. Different types of costs related to the use of a tunnel machine are discussed. The advantages of the tunnel machine over the traditional method are reviewed. IntroductionThe first tunnel boring machine ever used in a mine was in Quebec in the year 1957. Although less than successful at the time, this experiment served to illustrate the mining industry' s willingness to try new ideas. Since 1957, the manufacture of tunnel boring machines has developed into a full-size industry. Most contractors in Europe and North America will now consider the use of a tunnel boring machine first on any long tunnel that can have a circular section.The modern tunneling machine offers a great potential to the mining industry. Just as the blasthole and raise drill machines have revolutionized mining in the last decade, it is expected that tunnel boring machines will make an equal impact in the future.This is not to suggest that tunnel boring machines will replace explosives in the foreseeable future, but they can supplement drill-and-blast operations. In areas where long drifts, inclines, declines or cross-cuts are needed in a hurry, the tunnel boring machines offer the fastest way to a production return on investment. Figure 1 shows the range of production that two types of tunnel machines have produced as compared with conventional excavation methods in various ground conditions.Rotary hard-rock tunnel boring machines (Fig. 2) commonly produce progress rates in excess of 20 metres per day in uniform ground from 300 to over 2,000 kgl cm2 in unconfined compression strength . These averages are for the completed tunnels. One 7. I6-metre-diameter tunnel boring machine just completed a 1O.7-km tunnel in hard rock of 700 to 2,800 kg/cm>, The last eight months of production averaged 30 metres per day."

Jan 1, 1980

By Nicola Robjohns, Kenneth A. Grace, Jane Spooner

"The important volume end use sectors for the rare earth elements and yttrium continue to be in steelmaking, petroleum cracking catalysts and glass polishing. The market sectors which may provide future high volume demand, to replace the older sectors where demand has declined over the past decade, include magnets, automobile catalysts, ceramics and superconductors. Increases in production capacity in all major rare earth producing countries are expected to restrain large price increases for high purity rare earth oxides over the short term. A sharp increase in demand for yttrium in advanced ceramics and superconductors would lead to a shortage in separation capacity, however, and to a corresponding rise in prices for separated rare earths.IntroductionWhile visiting the Geology Department of the University of Toronto in 1981, Professor Lu from the Institute of Geochemistry of Academia Sinica at Gweiyang in China introduced one of his seminars with the observation that in China, rare earths are not rare at all.Indeed, while being rare in terms of abundance as measured by, say, the Clarke scale, the rare earth elements, like many other low abundance elements, are subject to geological, geochemical and mineralogical concentration.The rare earth elements comprise the fifteen inner transition elements of Group IIIB, Period 6, of the periodic table, from lanthanum to lutetium (the lanthanides), as set out in Table 1. The rare earths constitute a distinct group as they have extremely similar chemical characteristics due to the uniformity of their outer electron shells. The element yttrium is usually, by convention, included in the rare earth group as it occurs with the rare earths in nature and has properties very similar to them.The rare earths are divided into two subgroups, the ""light"" rare earths, made up of the first seven elements of the series, lanthanum to europium, and the ""heavy"" rare earths, comprising the remaining eight elements, gadolinium to lutetium, plus yttrium. These two subgroups are also referred to as the cerium subgroup and the yttrium subgroup, respectively. The subdivision is based on differences in the atomic radii resulting in somewhat different properties for the two subgroups."

Jan 1, 1991

By Adrianus (Adrian) Halim, Tomas Bolsöy, Stina Klemo

"This paper describes the ventilation system used in underground hard rock mines in Sweden and Finland, which is very different than the system used in other major mining countries such as Australia, Canada, South Africa, and USA. The system utilizes auxiliary fans with Variable Speed Drive to distribute primary airflow to working levels instead of regulators that are commonly used in other countries. Descriptions about Ventilation on Demand in this system and ventilation regulations in Sweden and Finland are presented in this paper. This paper also quantifies a comparison between the Nordic and the system used in other countries. KEYWORDS Auxiliary fan, Regulation, Ventilation on demand (VOD), Variable speed drive (VSD) INTRODUCTION The primary objective of an underground mine ventilation system is to provide sufficient airflow to dilute atmospheric contaminants that are produced by mining activities. Typically, a mine ventilation system consists of one to four primary fans, regulators to control primary airflow distribution to levels or panels, and auxiliary fans that distribute airflow to dead-end workings via ducts. Depending on circuit geometry, underground booster fans are sometimes used. This system is widely employed in many countries such as Australia, Canada, USA, United Kingdom, and South Africa. However, mines in Nordic countries such as in Sweden and Finland employ a different system, in which auxiliary fans instead of regulators are used to control primary airflow distribution, as well as distributing airflow to dead-end workings. This paper describes the general layout of Nordic ventilation system, its VOD and heating system, its comparison with the system that is widely used in other countries which is referred as “Non-Nordic ventilation system” in this paper, and the current ventilation regulations in Sweden and Finland. The comparison was quantified by doing simulations using Ventsim Design 5 software. DESCRIPTION OF NORDIC VENTILATION SYSTEM General System Layout Figure 1 shows a schematic of typical ventilation system used in Swedish and Finnish mines. The schematic was drawn by Franzen, Myran, Larsson, and Rustan from Stiftelsen bergteknik forskning (Swedish Rock Engineering Research Foundation) and the English translation is written next to the Swedish terminologies. The system is generally a push-pull system. None of Swedish and Finnish mines use exhaust system. As shown in Figure 1, the primary airflow is distributed to levels by auxiliary fans that are bolted to a bulkhead located in the access to intake raise/shaft. These intake fans are attached to ventilation ducts that distribute airflow to all working faces. The exhaust auxiliary fans, which are not attached to any ducts and also are bolted to a bulkhead located in the access to exhaust raise/shaft, then flow the exhaust airflow into the exhaust shaft. A small quantity of this exhaust air is directed into the ramp, which acts as another main exhaust airway. The deepest level is also equipped with booster fan(s) that draw air from the intake shaft/raise and direct it into the ramp to provide fresh air in the ramp. There is no free split in this system. Figure 2 shows a schematic of a level ventilation layout in the Nordic system. Figure 3 shows a photo of these auxiliary fans."

Jan 1, 2019

By P. Chryssanthakis

"The aim of this study is to give a better understanding for the behaviour of the composite material RRS and the synergy effect of its components while an excavation face advances during tunneling. The lower part of the Q-system scale (i.e. Q-values of 0.4 or less for a tunnel span of 10 m) is commonly used for tunnel applications in softer rock formations. This is implied in the updated Q-chart, where the dimensioning of rock support by using Reinforced Ribs of Shotcrete (RRS) comprises the use of fiber reinforced shotcrete, S(fr), rock bolts and shotcreted beams reinforced with steel rebars. The newest part of the lower part of the Q-chart gives more information on required support compared to older Q-system versions. The RRS can be described as a three part composite material with the 3rd component being the reinforced concrete beam (with no fibers) and with steel rebars that are placed in groups and sprayed with pneumatically applied shotcrete in layers. The RRS usually have a total thickness varying between 0.20 and 0.50 m and a width between 0.25 and 0.60 m respectively. They are placed at different distances along the tunnel axis depending on the rock mass quality.In this study, an effort has been made to simulate RRS in a numerical code and to study its behaviour in a real case study. A double tube, triple lane tunnel in weathered tuff has been chosen as reference material for testing RRS under changing load. This was done by writing a FISH subroutine which is implemented in the discontinuous numerical code Universal Distinct Element Code (UDEC) that takes into account the characteristics of the reinforced beam (height, width, number of steel rebars, and spacing) in the calculation process. A rather extensive parametric study has been run by using the same key input rock mass parameters and in situ stresses (?H/?v = 0.5), and changing only the RRS key input parameters in order to simulate the changing loads during the face excavation process in tunneling. The critical part of this numerical modelling work that is under investigation in this study, is the period just after the excavation: i.e 4 hours; 16 hours after excavation; and finally after the complete curing (28 days) of the cement based components. The RRS are installed in three consecutive steps as follows: a) the application of S(fr) on the rock surface in order to even it up for the RRS application; b) radial bolting; and c) installation of the reinforcing steel rebars, that form the reinforced beam.The results show that: a) the S(fr) is almost overloaded in the first 4 hours after application in both tunnels; b) on the S(fr) and steel rebars, the stresses, axial and shear loads, as well as moments are decreasing after 16 hours and 28 days; and c) the bolts loads are increasing with time. The observed reduction in stresses and loads in both the S(fr) and reinforced beam is surprising. A possible explanation is that tunnel arches are self-supported to a lesser extent as time progresses due to redistribution of stresses in the nearby rock mass through the radial bolts. The loads in radial bolts are increasing by almost 25%. This can be seen from the overall results as we observe an increase in maximum principal stresses, displacements, shear displacements along joints and hydraulic apertures compared to tunnel baseline that has no support of any kind."

Jan 1, 2015

By W. R. McQuarry, M. E. Wilson

"The structural relations of the ore deposits of the Staclacona mine are of considerable interest, for they not only illustrate the effect of a change in competence of the wall-rock, but also the effect of either the convergence together of two faults, or a change in the clip of a fault, or both of these factors, on ore deposition.The mine is in claim M.L. 1818 near the eastern end of Pelletier lake and lies about 2 miles south of the town of Rouyn. The Company was organized in 1926 to explore claims in Rouyn, Dufresnoy, and Beauchastel townships, but attention was directed mainly to those in Rouyn. Underground work began in 1928 with the sinking of the main shaft to a depth of 300 feet. Except for 1930 and 1931, the mine has been in operation continuously since that time. Production, which began in 1936, up to the end of 1946 has amounted to 1,381,422 tons of ore from which gold and silver having a total value of $7,861,860 were recovered. The average grade of the ore milled in 1946 was $6.68.GEOLOGYThe rocks in which the Stadacona ore deposits occur are mainly volcanics of the (Keewatin) Abitibi series intruded here and there by numerous dykes, sills, or masses of diorite. Single dykes of olivine diabase and lamprophyre are also present in the property and are intersected in the mine workings. The volcanic rocks are part of a wide belt of northward dipping lavas and pyroclastic rocks, having a thickness of over 16,000 feet, that extend from the east-west trending belt of Timiskaming sediments on the south northward to the Horne Creek fault and form the south limb of a large syncline. The volcanic rocks within the property consist of numerous, east-west trending, northward dipping flows of andesite and a few flows of rhyolite interbedded with rhyolite ruff and agglomerate. The angle of dip throughout most of the property is 60° to 80°N., but in the main cross-cut on the 850- and 975-foot levels stratification in rhyolite tuff was noted dipping 35°N. The tuff at this point lies above an intrusive mass of diorite and it is probable that the change in dip results from the way in which the tuff and agglomerate have been thrust aside by the intrusive."

Jan 1, 1954

By Eric D. Jamieson

"How well is the Canadian coal industry able to expand in the tough competition with other fuels in Canada and with other suppliers in the international, metallurgical and thermal coal trade? Coal production as a whole decreased between 1950and 1960 and, although there has been a tripling of production since then, underground production continued to decrease until 1975. This trend has been reversed to a certain extent as a result of the modernization of mines in Cape Breton, but in 1978 over 85% of coal production came from surface mining.Against this background, the paper reviews current Canadian mining operations as described in papers presented at the 1979 CIM Annual General Meeting. Conclusions are then drawn regarding some of the longer-term mining applications in relation to the latest coal demand forecast, which indicates a potential Canadian production requirement of 160 million tonnes per year by 2000. IntroductionThe Canadian 'coal industry is experiencing a period of expansion and a new record for production is set with each passing year. This growth pattern over the last IS years has been based on two main markets-the generation of electricity in Canada and the export of coking coal to Japan. Supply to those markets has entailed either the development of mining, coal preparation and transportation systems for the delivery of large volumes of products with predetermined specifications or the establishment of minesite plants designed specifically to suit the characteristics of the coal in the adjacent deposits (Fig. I).Coal is a low-value commodity in relation to its bulk, and the economics of coal use become progressively less attractive with increasing distance from the producing mine. The transportation component in the export trade and in the movement of western coal to Ontario is a major cost factor. In the latter case, it approaches the cost of coal produced at the mine. In addition, the .international competition among coking coal producers chasing diminishing markets in a worldwide steel recession as well as the competition from U.S. coals and other fuel s for Canadian thermal coal have introduced a tight set of parameters to govern mining practices in all parts of the country. Some appreciation of the required changes in mining techniques brought about by declining traditional markets and the rigorous demands of the new bulk markets can be gained by comparing production from surface and underground mines over the past 25 years (Fig . 2)."

Jan 1, 1980

By Yves Breau, François Robichaud, Jacques McMullen, Pierre Pelletie

"The Expert System implemented at the East-Malartic mill controls a copper flotation circuit while taking into account the profitability of the operation. Treating a copper gold ore, the flow sheet integrates a flotation step followed by cyanide leach circuit. The Expert System calculates the actual net smelter revenue of the flotation circuit, accounts for the operating cost of the cyanidation circuit and defines optimal metallurgical operating targets to optimize the overall economics. To be realistic, probability factors and circuit diagnostic analysis are integrated in the operating target selection. The Expert System has proved to be successful, robust and flexible. It resulted in a copper concentrate grade increase of 1% while maintaining recovery. The use of the expert system also gives plant personnel good training procedures and promotes interaction with mill operator.INTRODUCTIONThe East-Malartic mill is located close to the town of Malartic in North Western Quebec, Canada. The mill was first commissioned in 1938. The flowsheet at that time was a straight cyanidation circuit. The original mill operated in that mode up to the discovery of the Bousquet 2 orebody in 1989 which consisted of a massive copper sulphide ore averaging 0.7% copper and 7 g/mt of gold. Bousquet 2 ore processing required a mill retrofit. Grinding capacity, gravity and copper flotation circuits were inserted prior to the cyanidation circuit in order to achieve 3,200 mtpd capacity at expected recoveries.In 1993, control of the flotation circuit was deemed critical in order to maintain the flotation performance and to reduce cyanide consumption since the copper sulphide contained in the Bousquet 2 ore is a high cyanide consumer (50% bornite). (McMullen, 1999) As a result, the implementation of advanced control strategies in order to increase the metallurgical profitability of the plant and to reduce the operating variability between operators was the most rational alternative. In 1994, Corem was chosen to perform the plant audit and to write the first set of rules. The on-line control strategy was started up in 1995. In 1996 and 1997 the Expert System was shutdown because of an extreme change in the ore coming from the milling of Bousquet #1 ore and the departure of the projects engineer. In 1998, the system was restarted and the results were impressive. Furthermore in May of 2001 a patent was issued for the control strategy. (Breau, 2001)"

Jan 1, 2002

By N. Wachgama, J. P. T. Kapusta, R. U. Pagador

The Thai Copper Industries smelter (Thai Copper) is located some 200 km southeast of Bangkok near the MapTaPhut deep sea port in one of the industrial centres of Thailand. Although construction of the complex started in the mid 1990s, the smelter, acid plant and refinery were completed in 2004. Thai Copper is a state-of-the-art copper smelter processing imported concentrates with a total yearly capacity of 165,000 tonnes of cathodes. It employs the El Teniente Technology where concentrates are blended and smelted continuously in a primary smelting reactor to produce high grade matte or white metal. The white metal is converted in batches to blister copper in Hoboken siphon converters equipped with Air Liquide Shrouded Injector (ALS I) Technology for sonic injection of oxygen enriched air. Final refining is carried out in two anode furnaces prior to anode casting and electrorefining. Thai Copper and Air Liquide Canada executed a technical assistance agreement in late 2005. Pre-commissioning of the shrouded injectors took place in July and November 2006 while hot commissioning occurred in December 2006. This paper presents the major technical aspects of the joint commissioning, including the establishment of operating practices specific to shrouded injection converting. Preliminary results of this first commercial implementation of ALSI Technology in copper converting are also provided.

Jan 1, 2007

By B. D. Thompson

Many mines adopt necessarily conservative backfilling strategies due to a lack of understanding of their backfill behaviour. In the past five years, extensive instrumentation campaigns have focused on measuring in-situ cemented paste backfill behaviour. These tests have resulted in optimised backfilling efficiency through elimination or reduction of cure periods where; i) controls on backfill behaviour are understood, and ii) barricade pressures are demonstrated to be within a safe range using routine monitoring as a required QA/QC procedure. Recently, this work has been extended to consider hydraulic backfill operations. Two stopes have been instrumented at Vale?s Coleman and Garson Ramp mines. The instrumentation and deployment methods are discussed, and representative total earth pressure, pore pressure, and temperature data are presented. In one case, pressures in the stope are relatively low as drainage causes strength gain in the fill and so pressures are arched. In a second example, again, drainage causes strength gain in the fill, but increasing temperatures correlate with pressure increases even during downtimes in filling. These pressure increases are thought to be caused by thermal expansion of the backfill in confined locations. Generally, the instrumentation approach has proven successful in measuring hydraulic backfill behaviour and offers good potential for incorporation into a procedure to safely maximise backfilling efficiency through routine barricade pressure monitoring, when combined with safeguards to ensure adequate fill drainage.

Aug 1, 2013

By J. S. Godard

The mining developments in western Quebec have shown the ores of this district to be more complex in character than was at first expected. When the initial discoveries were made it was thought that the ores would be similar in character to those of Kirkland Lake and Porcupine areas. Developments have since shown that although the ores along the southern margin of the belt are somewhat similar, those within the belt, constituting by far the greater proportion, are vastly different, being complex mixtures of base-metal sulphides carrying values in the precious metals. These mixtures of base-metal sulphides of copper, zinc, and iron, present several problems to the metallurgist, especially in their concentration into marketable products. For the purpose .of this paper the following classification may be made of the ores of the district. Group I. - Milling Ores: (a) Ores in which~ arsenopyrite is the predominating sulphide. In this class of ore the gold occurs in fractures in the quartz and is associated to a lesser extent with the arsenopyrite and pyrite. Examples of this class of ore are the Granada, south of Pelletier lake, and the O'Brien and the Thompson in Cadillac township. (b) Ores in which iron pyrite is the predominating sulphide and gold is the valuable constituent. This may also be classed as a concentrating ore-Group III.

Jan 1, 1928

By N. Erarslan, M. Serati

"This paper attempts to provide a simple analytical solution for the calculation of the stress tensor induced in a two-dimensional elastic ring of a brittle material subjected to general loading on either its internal or external curved surfaces. Airy stress functions and Michell’s expression in polar coordinates are used, and parametric studies are performed to investigate the influence of loading angle and the geometrical aspect ratio of the ring on the maximum induced tensile stress. INTRODUCTIONA proper and accurate estimation of the tensile strength of a given material is a crucial step in many engineering projects; in particular in rock cutting, since rocks are much weaker in tension than in compression or shear. Both direct and indirect laboratory techniques such as the direct uniaxial pulling test, the Brazilian indirect tensile test (also known as the split cylinder test), and the ring test have been proposed and developed to estimate rock tensile strength. Among these, the ring test, initially suggested by Ripperger and Davids (1946), has become particularly appealing for its ease of sample preparation and its unique breakage mechanism in almost a pure tensile mode. Unlike the Brazilian test in which high shear stresses developed in the vicinity of the contact could interfere with the tensile breakage of the disc specimen, a tensile failure starts away from the loading platens in a ring test (Hobbs, 1964)."

Jan 1, 2015

By Fred A. Webber

STEEL ROPE WIRE is one of the more specialized commodities among the many grades of highcarbon steel wires. In its production there are many metallurgical aspects of interest from the steelmaking on through the final stages of wire drawing. We produce rope wire in an integrated mill com prising all the operations from the blast furnaces to finished wire drawing in the wire mills. Steelmaking is carried out with close control on hot metal, selection of scrap, and fluxing materials. The basic open hearths are of 170 tons capacity. Steel is teemed into 19 1/2 in by 19 1/2 in big-end-down molds and is hot-topped with the use of clay refractory hot-tops. The weight of this ingot is 6,900 pounds. All high -carbon steel ingots are charged hot into the soaking pits and are then rolled through the blooming mill in 16 p asses to a bloom ax~ in. by 41/2 in. and approximately 65 feet long. Necessary bottom and top crops are taken on the bloom as it passes from the blooming? mill to the billet mill. The still hot bloom is continuously rolled through the billet mill in . 5 passes to a 2 3/8 in. square. Billets are cut to 32 feet in length by a flying shear and then allowed to cool on the hot bed.

Jan 1, 1957

By Walter Valery, Lauren Hafla, Megan Tibbals, Adrian Dance, Jeff Snyder

"Newmont Mining Corporation’s Carlin (“Carlin”) operation processes roaster feed through a double rotator, dry grinding mill following secondary crushing. Metso Process Technology & Innovation (“PTI”) was involved in a review of crushing and dry grinding circuit operations in August 2009. A number of opportunities for increased throughput were identified in the PTI study.Extended plant trials were conducted in July 2010 to provide justification for possible circuit modifications and if needed, additional equipment. The trials confirmed that a finer crusher product would reduce the build-up of material ahead of the coarse grinding chamber in the double-rotator. Also revealed were material transport and airflow restrictions in the dry grinding circuit that resulted in a high circulating load.The secondary crushing plant demonstrated it was capable of reducing the amount of +¾” material in the mill feed thus increasing dry grinding throughput. However, wet and sticky feed affected screening efficiency resulting in low utilisation of the existing crusher capacity. The open circuit configuration also made it difficult to control crushing circuit product size.For the dry grinding circuit, changes in grinding media size, transfer grate design and increased airflow were recommended to reduce circulating load issues and increase circuit capacity. This paper summarises the crushing and grinding circuit review, the extended site trials performed by PTI and the improvements in dry grinding operation that resulted from this study."

Jan 1, 2012

By C. Diaz, S. W. Marcuson, K. I. Burgess

Deep roasting in fluid bed roasters is the first unit operation in Inco&apos;s "roast-reduction smelting" process for the treatment of nickel sulphide concentrates. The roaster product, a low sulphur calcine, is reduction smelted with coke in an electric furnace to yield a sulphur-deficient matte which is then converted to low (<1%) iron matte in Peirce-Smith converters. Due to the deep roasting, a minimal amount of sulphur dioxide is evolved in the smelting and converting stages. Sulphur is eliminated from the concentrate almost exclusively in the roasters in a single, continuous, strong, gas stream suitable for treatment in an acid plant. This paper discusses the commercial scale testing of deep fluid bed roasting of nickel concentrate at Inco&apos;s Thompson Smelter as part of two commercial scale campaigns on roast reduction smelting conducted in 1981-1982. Slurry feeding, practised during the second campaign, permitted excellent control of the air/concentrate ratio in the roaster and, in turn, good control of sulphur elimination (about 80% of input). Smooth autogenous operation with high oxygen efficiencies (about 95%) was promoted by operating at bed temperatures of830°C to 850°C with calcine recycle. High throughputs and the desired metallurgy were achieved. The extensive testwork conducted during the two experimental RRS campaigns demonstrated thot deep fluid bed roasting of nickel concentrate at high oxygen efficiency is a technically viable operation.

Jan 1, 1994

By J. J. Mather

NEPHELINE SYENITE is a feldspathic rock resembling granite in texture and hardness but it is white in colour and contains appreciable quantities of the mineral nepheline, the result of a silica deficiency. It is relatively common, with known occurrences in Russia, Scandinavia, India, Canada, United States, and other parts of the world. However, the particular deposit being mined at Nephton, Ontario, by American Nepheline Limited is unique in its remarkable uniformity and low content of iron bearing minerals, both qualities essential in a ceramic raw material. The orebody is a pear-shaped mass over five miles in length, up to a mile in width, rising in three prominent ridges as much as 300 feet above the surrounding lakestudded terrain. American Nepheline holdings, consisting of 2,200 acres, comprise some 85 per cent of the total mass.

Jan 1, 1957

By T. S. Wanne

Tunnel sealing experiment (TSX) was conducted between 1997 and 2004 in the Canadian URL located in Manitoba. In the experiment, a tunnel was excavated parallel to the maximum principal stress direction, filled with sand and sealed with bulkheads at both ends. It was then pressurized and heated by injecting hot water. The rock mass response was monitored with seismic methods. The main interest in the numerical simulation was the damage formed during the heating phase. The large-scale PFC2D models replicated a central plane- strain section of the TSX chamber. The models captured the first-order phenomena observed in-situ displaying the difference in the damage in the roof and floor regions, respectively. This difference was due to the filling material confinement of about 100 kPa on the tunnel floor, which occurred in the course of the heated experiment.

May 1, 2009

By E. Cauda, K. Phillips

"Diesel particulate matter (DPM) exposures are a serious occupational health hazard in underground mines. While the complex nature of DPM makes direct analysis difficult, its primary component, elemental carbon (EC), can be used as surrogate. The near real-time FLIR Airtec monitor measures EC mass based on laser extinction. It was developed for use in underground metal/non-metal mines, and calibrated to the NIOSH Method 5040. Its effectiveness has not been demonstrated in coal mines, however, where coal-sourced EC could produce analytical interference. To gain preliminary insights regarding the Airtec’s performance in the presence of coal dust, experiments were conducted in a controlled laboratory chamber. The Airtec EC results consistently underestimated Method 5040 EC for both respirable-sized and submicron coal dusts. While underestimation of coal-sourced EC bodes well for limiting coal interference for DPM measurement, one key reason for the observed trend was non-uniform coal particle deposition on the Airtec filter. This may be problematic in coal mines where DPM and coal dust occur together in the instance that DPM deposition is also non-uniform. INTRODUCTION Diesel particulate matter (DPM) is an occupational health hazard, and underground miners have the potential for relatively high exposures due to their work in confined environments with large diesel equipment (NIOSH, 2011). DPM is heterogenous and complex in terms of its chemical composition (Birch & Cary, 1996; Noll et al., 2006), but elemental (EC) and total carbon (TC) are often used as analytical surrogates (Birch & Noll, 2004; Noll et al, 2015). EC is the solid “soot” component of DPM, whereas TC also includes particulate organic carbon (OC) (Noll et al., 2007). EC and OC can be measured in filter samples using the standard NIOSH Method 5040 (NIOSH, 2003), and summed to obtain TC. In US mines, TC has been adopted as the analytical surrogate for DPM in regulatory activities. In metal/non-metal mines, personal exposures are limited under 30 CFR Part 57. In coal mines, limitations are made on emissions from each piece of equipment (30 CFR Part 72) rather than for personal exposures. This is because coal dust has the potential to interfere with TC measurements in personal samples. In some other countries, and for non-regulatory purposes in the US, EC is also commonly used as a surrogate for DPM in mine monitoring applications [e.g., see AIOSH, 2013). EC generally exhibits a linear correlation with TC (Birch & Cary, 1996; Noll et al., 2006 and 2007) but unlike TC measurements, which are analytically intensive, EC can be determined more simply using a laser absorption/extinction principle on filter samples (Noll et al., 2013)."

Jan 1, 2019

"In conjunction with Banick Mercur , Wright Engineers Limited designed the first multi-stage aqueous pressure oxidation system in an alkaline environment in North America. A three-stage slurry heating and slurry pressure let-down with heat recovery was an innovative approach to the problem of high energy cost with the low sulphur content of the ore. The design of this system required solving a number of problems associated with heating and pumping of slurries at pressures up to 3230 kPa and temperatures in excess of 220°C. Wright received an Award of Merit in October 1989, from the Canadian Consulting Engineering Awards Committee in the Process Category for its design of the Mercur pressure oxidation system.The Mercur Mine is located at the south end of the Oquirrh Mountains in Tooele County about 65 miles by road from Salt Lake City, Utah, U.S.A. The original mill was designed to process 5000 tpd of oxidized gold ore in a carbon-in-leach (CIL) circuit. A new ore treatment system was required to treat the substantial quantities of refractory ore reserves which could not be treated economically in the existing mill. The new pressure oxidation circuit makes the refractory ore amenable to gold recovery by oxidation of the sulphides contained in the ore .Wright Engineers Limited was commissioned by Barrick Mercur Gold Mine Ltd ., a wholly-owned subsidiary of American Barrick Resources, to design the pressure oxidation facility. This followed completion of a feasibility study by Wright which showed that the project was technically and economically viable.Work on the feasibility study began in 1986, detail design commenced in February 1987, and was completed in December 1987. Mechanical installation was completed in January 1988, and slurry feed began in February 1988. The plant has operated since June 1988 without major interruption . Wright was responsible for engineering, procurement and construction management for the addition of the 750 tpd pressure oxidation circuit to the existing mill and also had commissioning responsibility.Major unit processes involved in the pressure oxidation facility are ore thickening, three-stage slurry preheating, pressure oxidation and three-stage slurry cooling with heat recovery, stearn generation, propane and oxygen storage and a video graphic display process control system were included in the scope as process anciliaries.The pressure oxidation increases gold recovery from the refractory ore by conventional CIL from as low as 35% to more than 86070. A patent has been granted for the process."

Jan 1, 1989