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By D. A. Nelson, Wedding. R. E.

"A planned production increase of 1.25 million tonnes per year at Domtar's salt mine in Goderich, Ontario required an increase in the underground mine ventilation system from 320,000 to 700,000 scfm, The design of the expanded facility included the addition of a 22-foot-diameter shaft to augment the existing production and service shafts.A preliminary review of the ventilation requirement s indicated that in order to pr ovide good economic design for the expanded system, a comprehensive ventilation survey of the existing system was required. The survey was completed in the latter part of 1980, and subsequent engineering design and equipment selection proceeded throughout 1981.This paper presents the methods used to carry out the survey. Particular attention is given to the numerous challenges presented in determining air flows in the large 45-ftwide by 43-ft-high airways. Emphasis is placed on practical field survey techniques.A mathematical model of the ventilation system using a computer program is described, and the application of this program to provide design data and assist in equipment selection is discussed. Economic sizing of airways and fan selection are analyzed considering horsepower and operating costs.IntroductionThe Goderich Mine of Sifto Salt Division, Domtar Inc., is located on the eastern shore of Lake Huron in the town of Goderich, Ontario, and has been operating since November 1959. The salt is extracted by a modified room-and-pillar method, and the workings advance at a rate of 1.5 miles per year by undercutting, drilling and blasting. The surface mine area consists of a 12-acre site in the Goderich harbour, at the mouth of the Maitland River, and the underground property consists of approximately 2,700 acres lying mainly under Lake Huron."

Jan 1, 1983

By J. F. Wright

INTRODUCTION Eighteen years ago about half a dozen experienced prospectors moved from northern Ontario to northwest Manitoba to prospect 'greenstone' areas shown on geological maps of .parts of northern Manitoba prepared by J. B. Tyrrell, D. B. Dowling, and Wm. Mclnnes. Gold-bearing quartz was soon discovered by these men in this new field, and in succeeding years many more prospectors entered the district. A number of the pioneers are still prospecting, and since 1913 mineral deposits have been discovered each year in widely separated areas of the large district extending 150 miles eastward from the 102nd meridian and 60 miles north of latitude 54°30'. The majority of the discoveries, however, have been made near Schist, Elbow, Wekusko, and Kississing lakes. One hundred or more mineral occurrences of some merit are known, the main types being sulphide replacement bodies containing nickel-copper or copper-zinc and gold-bearing quartz veins. Some underground exploration or diamond drilling has been completed on 25 or more of the prospects. Up to the present, three ore-bodies, namely, Flin Flon, Sherritt-Gordon, and Mandy, have been proved. Of these, Flin Flon, the largest, is estimated to contain from 16,000,000 o 19,000,000 tons of copper-zinc ore carrying some gold, silver, and lead. Next in importance in regard to tonnage is Sherritt-Gordon, with a calculated reserve of nearly 5,250,000 tons of copper-zinc ore carrying less than $1.00 combined gold and silver.

Jan 1, 1931

By P. R. A. Andrews, R. K. Collings

"There are more than thirty-five deposits or occurrences of celestite in Canada, chiefly in Newfoundland, Nova Scotia, New Brunswick, Ontario and British Columbia. Deposits of economic interest are the Lake Enon and MacRae deposits in the Loch Lomond district of Cape Breton Island, Nova Scotia. Two Cape Breton deposits were operated from 1969 to 1976 by Kaiser Celestite Mining Limited, which supplied an upgraded celestite concentrate to an associated strontium carbonate plant at Point Edward, near Sydney.The MacRae deposit is of current interest because of its possible development for strontium carbonate production using the ""black ash"" process. This process involves chemical leaching, rotary kiln calcining with petroleum coke, washing and treatment with carbon dioxide.Celestite is the basic material used in the manufacture of strontium chemicals, principally strontium oxide, carbonate, nitrate, chromate, chlorate, chloride, sulphide and phosphate. Strontium chemicals, in turn, are used in the manufacture of glass and ceramics, in pyrotechnic devices, in the production of strontium metal and in other metallurgical applications. Recently, strontium ferrites have been introduced into permanent ceramic magnets."

Jan 1, 1991

By H. Soto, P. Lacombe, R. del Villar

"During the summer of1989, Les Mines Selbaie, with the cooperation of Université Laval initiated a research project to evaluate the column flotation technology for its concentrator near Joutel, Quebec. The experimental work involved both a 5 cm by 11 m laboratory column and a 45 cm by 10 m home-made pilot unit Amenability tests were conducted using both units on various streams of the plant. Results showed that columns are not suitable for the copper cleaning circuit because the loss of precious metals, associated with the rejected pyrite, overcome the improved copper metallurgy. On the other hand, column testing at the zinc circuit showed improved zinc selectivity.Scale-up testing was', therefore, conducted on the first zinc cleaner concentrate using the laboratory-size column. The influence of variables such as porous parger grain size, wash water flow rate, airflow rate and froth depth values on column performance was' first assessed. Scale-up parameters were then determined using the optimal operating conditions from the previous tests. Simulation facilitated the determination of the most convenient plant-scale column arrangement.A number of conclusions concerning the experimental determination of scale-up parameters are also presented."

Jan 1, 1993

By P. Hanniala, S. Harkki, B. Andersson

"The Outokumpu Flash Smelting Process was put into commercial operation in 1949. The original process was based on the use of preheated air. The adoption of oxygen enrichment of the flash smelting process air in 1971 was a crucial improvement, especially for smelter capacity. Oxygen use according to the Outokumpu method means selection of the most suitable oxygen content for the process air on the basis of the concentrate composition, capacity and matte grade. The furnace is not bound to only one fixed degree of process air oxygen enrichment; the oxygen content may vary from 21% to 100%.The degree of oxygen enrichment and the matte grade are the most important process parameters in reducing the operating costs of the whole system, including smelting, converting and acid production. Raising the matte grade allows more efficient use of the reaction heat in the smelting stage. Sulphur is recovered in a high-strength, constant-flow gas stream. The blowing time in the converters is reduced considerably.Oxygen enrichment and high matte grade improve the energy efficiency of the plant complex as long as the surplus heat from the system is not wasted. When the operating costs and investment expenditure of the total plant complex are taken into account, a reasonably high degree of oxygen enrichment (35-80°/0) is the best alternative in practice. Existing operating smelters have achieved single-stream thoughputs of over one million t/yr of concentrates.HistoryLocal capital and operating costs as well as lower energy consumption per unit of metal produced are now the crucial factors in determining the choice of process. In the late I940s, the same reasons prompted Outokumpu to develop the original flash smelting process in Finland. Between 1900 and 1940, very few technical changes in pyrometallurgical copper production took place. However, the commercial use of the flash smelting technique began a new era in metal extraction from sulphide concentrates. The characteristic feature of the method is that the concentrate is smelted in suspension: this offers the most favourable conditions for heat and mass transfer, and thus high reaction velocities. Outokumpu was the first to exploit the known advantages of flash smelting in a commercial installation at Harjavalta in 1949."

Jan 1, 1982

By R. W. Zeindler, D. A. Fawcett

"A summary is presented of the report completed by the Coal Mining Research Centre in 1980. This report was of a cross-pit conveyor system applied as an auxiliary mining complement to the major stripping unit, a dragline. The purpose of the CMRC report was to evaluate selective mining and replacement of the upper horizons of the strata as an aid in reclamation. These strata were the topsoil, subsoil and glacial till.Past utilization of cross-pit conveyor systems and related engineering studies were assessed. The parameters of the study were based on the mining and geological conditions of the Alberta prairie coal mines. The principal excavator for stripping was a dragline. Excavation of the upper horizons was done by a bucket-wheel excavator discharging onto the crosspit conveyor. Alternative equipment applications were economically compared. Four cases or geological sections were evaluated in detail. The economics of the alternative mining systems for each of the cases were determined. In all instances, the most economical solution was a tandem system utilizing a dragline with a bucket-wheel excavator/ cross-pit conveyor system."

Jan 1, 1981

By V. K. Karra

"A simplistic algorithm jar simulating the rod or ball mill Bond grindability test applied to ores and minerals is presented. A performance comparison between the present algorithm and a previously published algorithm by Kapurot, modified in this work jar a fair evaluation, is also given. A worked example is furnished jar leading the reader through the various calculations.IntroductionThe Bond grindability test is widely used for the prediction of the power requirement of an industrial grinding mill. The procedures for conducting this test in either a ball or a rod mill are available(1). In the case of a ball mill, batch grind abilities have been correlated to Bond grindabilities to help monitor day-today variations in ore grindability for the purpose of plant control( 2,3). Of special interest is the algorithm presented by Kapur(3) for the simulation of the Bond test. In practice, the test lasts from 6 to 14 cycles (4 to 10 hours). With this algorithm, one can predict the Bond grind ability using information available from only the first two cycles.Sometimes the prepared feed for the grindability test contains too much fines compared to the total product desired at the mesh of grind under consideration. In such a case, the operator separates the fines from the test feed and adds to the remaining coarse fraction an equal amount of fresh feed. This combined feed then forms the feed to the first cycle of the test. In the present work, Kapur's algorithm is modified to include this special case. A new and relatively simpler algorithm is also presented. Statistical equations relating work index to pertinent variables are also given."

Jan 1, 1981

By R. G. Bradbury

The paper uses two hypothetical, but typical, situations to show that hydraulic rock drilling equipment has very significant economic advantages over pneumatic equipment both in drifting applications (even in the situation where advantage is not taken of the high productive capacity of the hydraulic drill) and production drilling applications. Other factors, including the non-economic advantages and the disadvantages of hydraulic drilling, are discussed. The paper concludes with a brief analysis of why the Canadian mining industry has been slow to take advantage of this apparent broadly advantageous advance in rock drilling technology.

Jan 1, 1979

By w. w. Fraser

"In late 1985, Hudson Bay Mining and Smelting Co., Limited announced the Namew Mine/Mill project. Environmental baseline monitoring and planning began in early 1985 while a major drilling program was under way to delineate the ore deposit. This paper overviews the environmental assessment work conducted for the mine/mill operation, and outlines the tailings and water management practices.Planned environmental works and approach were affected in late 1988 when a massive water problem was encountered in the mine. The requirement to depressurize the rock surrounding the underground workings led to the need to develop, on short notice, an environmentally-acceptable procedure to dispose of a large volume of slightly-saline water.Biological toxicity tests performed on the depressurized effluent proved the effluent to be non-toxic and thus allowed resolution of the disposal problem. This resolution includes the conversion of a large water body from native fish species to a rainbow trout fishery."

Jan 1, 1991

By J. G. Whellock, J. W. Matousek

"In spite of their importance to the economics of many metallurgical projects, the values and technical aspects of precious metals losses in smelting operations are relatively undeveloped subjects. This paper attempts to provide a condensed overview of precious metals losses in pyrometallurgy, using some of the limited production and experimental data available. It is intended for three general groups of metallurgists - operators attempting to improve recoveries, cost-accounting engineers making estimates of metal losses in new or modified smelting plants, and research engineers and students.Distribution coefficients of precious metals, the ratios of metal or matte assays to slag assays, are convenient means of comparing smelting processes and the tendencies of the different metals to enter the slag. For silver in lead blast furnaces this ratio is approximately 250;for gold and the platinum group metals in submerged-arc, electric furnaces it is 900. The ratio for gold and PGMs in converters varies with the degree of oxidation of the charge. The over-all cycle-average for industrial installations is tentatively estimated to be 35.IntroductionThe magnitude of precious metal s losses in pyrometallurgical operation s is difficult to assess. The primary producers are generally reluctant to discuss the subject. The by-product producers (the large sulphide smelters o f copper , nickel, and lead) might be more willing to disclose information, but these have precious metals in their feeds in such dilute quantities that accurate accounting is extremely difficult. It is not until numerous processing steps have been taken t hat the precious metals become sufficiently concentrated to permit acceptably accurate sampling and analytical procedures to be followed. Thus, knowledge of the precious metals contents o f the intermediate mattes and slags may simply be unavailable."

Jan 1, 1990

By C. J. Newman, K. Molnar, A. G. Storey, G. Macfarlane

"The Kidd Creek copper smelter started operation in June 1981 using the Mitsubishi Continuous Smelting Process. A steady increase in plant production rates has been demonstrated, and the annual tonnage of concentrate smelted is now double the original design rate. This paper summarizes recent modifications made to increase concentrate feed rates, increase on-line time, reduce copper losses and control the impurity level in anode copper. It includes improvements made to furnace design, improvements to reduce downtime f or boiler cleaning, and improvements made to the acid plant to allow handling of increased gas flowrates and higher gas strengths while maintaining low acid plant emission rates. IntroductionThe Kidd Creek Division of Fal conbridge Limited operates a base-metal mining and metallurgical complex, located in Timmins, Ontario, which includes an underground mine, a 13 500 tonne/day concentrator, a 130000 tonne/year zinc plant and a copper smelter and refinery, The copper smelter and refinery were commissioned in 1981, with an installed capacity of 59 000 tonnes/ year of cathode. The smelter, utilizing the continuous smelting technology of the Mitsubishi Process, has subsequently been expanded in production capacity, and in 1991 produced 113 500 tonnes of blister copper The capacity of the refinery, which uses Onahama No. 3 tankhouse technology, has been increased to 96 000 tonnes/year of cathode. The capacity increase at the smelter has been realized through the input of additional oxygen to both the smelting and converting furnaces, without increasing the process off-gas flowrate through the associated waste heat boilers.In this paper, a brief review of smelter operation and modifications in the period 1982 to 1987 is presented, together with a more detailed discussion of operation during 1988 to 1991, and a discussion of future planned improvements."

Jan 1, 1992

By Veronique Falmagne

"In 1987 the only knowledge available of the Ansil orebody was that it was high grade, deep and relatively small . High stresses were also anticipated because there had been rockbursts during shaft sinking and destress blasting had been necessary (Fig. 1). From the design, through the development and now in the mining stage, rock mechanics and microseismic monitoring have been used to maximize recovery and safety and minimize dilution. Examples are given of the use of a microseismic monitoring system in the early stages o f mining. This paper also presents a review of the mine's in situ stress measurement campaign, cable bolting practice and backfill procedure.The Design PhaseAfter the initial mapping and classification work, preIiminary numerical modelling was performed using assumed stresses of ov = oH and oH = 3ov• The empirical Mathew's analysis was then used to optimize stope dimensions based on geomechanics data available at the time (Bawden et 01., 1988; Falmagne, 1987; Potvin et 01., 1988). Because of high stresses predicted by numerical modelling and a brittle, strong and relatively unfractured host rock which was believed to be rockburst prone, it was decided to mine without temporary or permanent pillars. However, in order to satisfy production rates, the orebody had to be broken up into two horiwns separated by a triangular pillar (Fig. 2). The final mining sequence is shown in Figure 3.The Development PhaseAt this point, based on work accomplished to date, including mapping and observations from drifting and modelling, the following problems were identified: (1) flat backs in the hangingwall andesite were thought to be highly stressed; (2) flat backs in weak massive sulphide ore would also be subjected to high stress and be difficult to support; (3) the felsic dyke was noted as a "" stress raiser"" , because activity increased as drifts approached it; and (4) chlorite coating of joints in the hangingwall was thought to be an important factor in reducing their stability."

Jan 1, 1991

By W. A. D. Edwards

"The Edmonton region accounted for 40% of Alberta's sand and gravel production in 1981 (about 12 million m3). Most of this production was by commercial operations (75%) for use in essential public works projects (64% in road and asphalt production).Although the Alberta Geological Survey delineated 1157 million m3 of gravel, a number of areas are considered to be in short supply. An econometric model was used for forecast demand in the region. Predictions, based on this forecast and estimated supply, are that 15% of the region's gravel and 3% of the sand will be consumed by the year 2010.The City of Edmonton is the largest consumer of aggregate in the region. Approximately 40% of the gravel in Edmonton's current supply area (about 55 km radius) will be used by 2010. These estimates are considered optimistic because the supply figures do not account for sterilization of deposits - some has already taken place in the region. If 25% to 50% of Edmonton's current supplies (one major deposit) were sterilized it would result in $1,000 million of additional cost due to increased haulage alone.Information on the value of aggregate resources, the location of sand and gravel deposit s and forecast demand has been made available to all municipalities in the region to aid planners and regulators in effective future use of this important resource.INTRODUCTIONAlberta is the second largest producer of sand and gravel in Canada at ab out 50 million tonnes annually. Value of provincial production exceeded $120 million in 1985. In Albert a, aggregate is used primarily in road and building construction (about 50% and 15%, respectively) and is important in such activities as petroleum exploration, oil sands development and the forestry industry.The Edmonton region (Fig. I) accounts for about 40% of provincial production and has a per capita consumption rate of about 13.4 m3 of sand and gravel per annum. Although Edmonton dominates, with approximately 50% of regional production going into the city, aggregate is produced in every county and municipal district in the region.Sand and gravel is essential to the current way of life in the Edmonton region. The resource is valuable and dwindling. The Alberta Geological Survey (a department of the Alberta Research Council) has collected information on aggregate in this region for the last decade which ca n aid government planners and regulators in pre venting sterilization and misallocation of this critical resource. This paper is a summary of that information."

Jan 1, 1989

By R. E. Phelan

Introduction The Flin Flon deposit was discovered by Tom Creighton, an experienced prospector and trapper, who in 1914 made his headquarters at The Pas, Manitoba, and still lives there. The discovery was made in the fall of 1914, and Creighton and his associates spent the next summer in investigating the find. In August of 1915 the first claims were located. The property was optioned to Hayden Stone & Company, and in the spring of 1916 they started a diamond-drill campaign. Hayden Stone spent $50,000 on the property and, among other things, completed over 7,500 feet of diamond drilling. The ore not being as high grade as expected, and also being complex in character, Hayden Stone then abandoned their option. In the spring of 1917, Messrs. Earle, Faskin and associates took an option on the property and started an extensive drilling campaign. This campaign accomplished approximately 18,000 feet of drilling and secured for the Faskin interests a 32 per cent interest in the property. In 1920, W. B. Thompson and The Mining Corporation of Canada secured an option, and, while this option was in force, sunk two shafts in the ore-zone and did a considerable amount of drifting and cross-cutting in the ore to study the manner in which it occurred in place, and, by actual channel sampling, to check the grade indicated by the previously accomplished diamond drilling. All of this work had indicated that an ore-body existed on the property containing at least 18,000,000 tons of ore, assaying as follows: gold, 0.074 oz./ton; silver, 1.06 oz./ton; copper, 1.71 per cent; and zinc, ). 45 per cent (1). Complete analyses of composites from the drill holes showed that the ore was very complex in character. It contained the above values in gold, silver, copper, and zinc, as well as large amounts of sulphur and iron, and traces of lead, arsenic, antimony, cobalt, and various other metals. Although the ore was obviously pyritic in character, pyritic smelting, the usual method of treating this class of ore, could not be used to beneficiate the ore with any reasonable expectation of making such a venture a financial success, since coke and coal landed at Flin Flon would be expensive, and siliceous flux was known to be scarce locally and low in grade. The property therefore lay idle until 1925.

Jan 1, 1935

By M. J. Davie, D. S. Ross

"The successful design of a new mine water pumping system is in part dependent upon the accuracy with which the operating conditions can be forecast. This paper outlines the data which should be obtained on water quantity and quality. Methods of modifying the quality are also discussed. Some pumping systems for disposal of sludge are described, citing operating examples. The various types of mine dewatering pumps are described in terms of design characteristics as they affect operating efficiency and installation and maintenance costs. Piping system parameters are outlined. Finally, a hypothetical case study of pumping from a depth of 3,000 ft is assessed, with particular reference to capital costs.IntroductionMany mine operators would agree that the pumping system usually receives a low engineering priority during the design stage, and as a consequence causes some of the most frustrating problems during production. For a ""dry"" mine (up to 200 gpm), design of the pumping system may indeed be simple, but outside of the potash industry such mines are not common. For ""wet"" mines, the pumping system is important because it significantly affects the initial capital and on-going energy costs of the mine and, in the area of sludge disposal, the working conditions and production costs. Indeed, a failure in the pumping system usually has serious, and sometimes disastrous, consequences.The intent of this paper is to outline the factors which should be considered in the design of a pumping system and to describe the types of pumps used in mine dewatering and relate them to practical operating systems."

Jan 1, 1980

By Pierre Antoine Dhaiti, Gilbert Sergerie, Michel Dagbert, Jean-Maurice Brodeur

"Quality of iron ore is generally determined by several factors which can be correlated or independent. Among these f actors, we generally find per cent iron, silica, magnetite or hardness. Forecasting the short- and long-term fluctuations of these f actors in the future pit is an important feature of mine planning in these deposits. Computer simulation provides a realistic answer to these questions. It allows to generate values for the most important quality features of each shovel load with the proper correlations between them as well as between one load and the next being fully reproduced. Then different sampling schemes f or grade control or blending facilities can be examined on the simulated data. The mechanics of the simulation procedure and its potential applications are described in an example.IntroductionThe purpose of this paper is to illustrate the use of in-situ grade simulation in the planning o f the grade control in an iron ore min e. More precisely, given a production unit of given size, we need to know how many samples have to be collected in that unit so that we ca n recognize its grade with a reasonable precision. Conclusions that we would like to arrive at are of the type: with 4 blasthole samples we ca n get the iron grade of a 10,000 ton bloc k with less than a ± 2% Fe error 95% of the time.The approach t hat we take to solve this problem is simulation. We simulate the grade of samples and the grade of the block that contains these samples. We repeat the simulation for hundred s o f blocks. By comparing the sample grade with the block grade in each case, we determine the actual number o f blocks where the difference sample-bloc k is less than a given limit."

Jan 1, 1983

By H. Fuchs, W. Hilger, E. Prosser

"The Lone Gull discovery of Urangesellschaft Canada Limited was made in 1974 as a result of extensive exploration activities for unconformity related uranium deposits in the Baker Lake area of the Northwest Territories. The structurally controlled mineralization is hosted in Aphebian metasediments and intrusive granitic stocks located approximately 2 km south of the unconformity between the Lower Proterozoic Aphebian and the Middle Proterozoic The/on Formation.The original discovery was made with a helicopter mounted radiometric system during systematic prospection over the unconformity. Diamond drilling was first carried out in 1977 and continued, with an interruption in 1981, to the present. A total of 25,500 metres and 199 holes have been drilled and outlined three ore zones. Preliminary geological reserves are estimated at some 16 million kgs. U30 8 at an average grade of approximately 0.4%.IntroductionUrangesellschaft initially entered exploration in Canada as a participant in joint ventures in 1968, and was incorporated in Canada in 1974 as Urangesellschaft Canada Limited. Subsequently, UG was among the first companies to apply exploration criteria developed from the Athabasca Basin, namely the Middle Proterozoic unconformity over Archean/ Aphebian basement, for selection of exploration targets in the geologically similar Thelon Basin.The Baker Lake area in the Keewatin District of the Northwest Territories was selected from the extensive Thelon Basin because it had the best accessibility (Fig. 1). The hamlet of Baker Lake located close to the geographical centre of Canada, is largely an Inuit settlement with a population of approximately 1,100 inhabitants. The distance from Baker Lake to the Lone Gull property is about 80 km. Baker Lake is connected to the tidewater of Hudson's Bay via Chesterfield inlet and is serviced by scheduled domestic flights several times weekly from Churchill, Manitoba, about 650 km to the south."

Jan 1, 1986

By Sukumar Bandopadhyay, Arunapuram Sundararajan

"Unique production and environmental problems are associated with multiple seam mining where dragfines are used for overburden and interburden removal. On the production side, the problems are decreased dragfine efficiency in chopdown operations, the general need for spoil rehandle, complications in sequencing stripping and coal handling operations, and potential stability problems in highly stacked spoil. Many variations of dragline methods have been reported in the literature to encounter these problems. Because of the physical properties and peculiar hydrology of sub-arctic region mining of multiple seams in such areas probably could not be done with the methods used in temperate regions.The properties of thawed material would make spoil hard to place in a stable manner and thus would influence the mini sequence. For instance, multiple seam mining which requires some rehandling of spoil and prolonged exposure of overburden and interburden material might be precluded by the instability of thawed spoil and therefore mining techniques have to be modified. Since the deployment and operation of a dragline in modes different from conventional schemes will affect productivity, it will be necessary to study the effects on production of changing the dragfine operational schemes. This paper presents a heuristic simulation model of multi-seam dragline operation of a sub-arctic mine that can be used to analyze the handling of overburden and detailed study of the dragfine performance."

Jan 1, 1986

By B. M. Das, F. Claridge, V. K. Garga

Coal, as well as a large part of other minerals in Canada, is produced by surface mining methods. This has resulted in short and long-term accumulation of waste rock in dumps. Two main problems arise from the flow of water through these dumps: stability of dumps and the impact on stream flows. In recent years, a number of mines have started placing waste rock in valleys for general disposal or far constructing mine access road ways. All these have created challenges in the following areas: effective design of dumps and flow-through dams; stability monitoring of the structures; environmental impacts downstream by way of increased amount of fines in suspension, sedimentation, and toxic substances; and long-term status of flow and flood. Similar situations also exist in many civil engineering structures. This paper discusses the status of rock drains and the immediate and future research needs.

Jan 1, 1990

By James K. Gowans

"Cominco Ltd's Polaris Mine, located on Little Cornwallis Island, Northwest Territories (latitude 75degrees 30') went into production in late 1981, becoming the world's most northerly base metal mine. The concentrator, powerhouse, shops, offices, dry and warehouse, were uniquely designed and constructed on a 100f t wide by 400 ft long by 65 ft high barge. This process barge was constructed in southern Canada at Trois Rivières, then towed 3,500 miles into the High Arctic Islands and beached on Little Cornwallis Island.The paper deals with the original design philosophy, flowsheet and the initial year's operation of the Polaris process barge.IntroductionTo bring a mining property into production in the Canadian High Arctic, a company must deal with the harsh environment, the short shipping season and the high cost of labour. In trying to minimize the effect of these fact or s, the concept of a barge-mounted process plant was investigated as an alternative to the conventionally built concentrator. Cominco's previous experience gamed at the Black Angel Min e on Greenland's west coast suggested that this would be a practical alternative.The normal shipping season at Little Cornwallis Island is approximately six weeks long, from mid-August to the end of September. All equipment and materials must be delivered during the ""shipping window"". The logistics problems are enormous; if the conventional approach had been used the major process equipment would have been delivered to Little Cornwallis Island at the same time as the completed barge actually was! Construction would also be starting at the start of the worst climatic conditions, the Arctic winter. Nine months' production was gained by opting for the ""barge"" approach."

Jan 1, 1983