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By D. L. Romero

The Line Creek Mine is operated by Crows Nest Resources Limited, a wholly owned subsidiary of Shell Canada Resources Limited. It is located approximately 25 km north of Sparwood in southeast British Columbia (refer to figure 1). The Line Creek coal deposit is part of the Upper Elk Coalfield in the East Kootenays. The coal is a medium volatile, low sulfur, bituminous coal and is used as a coking coal. Coal close to the outcrop is typically of low F.S.I. and is processed for thermal use. The present pits have a total coal reserve of 30 million clean tonnes and is a multi-bench, open-pit operation. The coal is recovered by front shovel and backhoe combination, and hauled to the plant with a fleet of 120-tonne and 40-tonne capacity trucks. Line Creek is unique in that it can simultaneously treat both metallurgical and thermal coal in separate processing facilities (refer to figure 2). Run-of-mine coal is delivered to the truck dump. The plus 50 mm material is scalped off and reduced to minus 50 mm in two rotary breakers. The raw coal is then conveyed to one of five, 5000-tonne capacity, silos. Typically one silo is used for thermal coal and four for metallurgical coal (refer to figure 3).

Jan 1, 1991

By Alexander N. Winchell

"THE MICHIGAN COPPER DISTRICT.Location. The copper mines of Michigan are all in the north¬ern part of that state on the Keweenaw Peninsula, which extends from the (Michigan) mainland for about seventy miles northeast¬erly toward the middle of Lake Superior. The mines occur in a belt or zone occupying the central portion of the peninsula from end to end.. The most productive mines for many years past have been those located not far from the city of Houghton in' the central part of the zone, in a section extending about ten miles southwest of Houghton and fifteen miles northeast. The most important mines in order of.location, from the northeast to south¬west, are the following: Mohawk, Ahmeek, Allouez, North Kearsarge, Wolverine, South Kearsarge, Centennial, Calumet. and Hecla, Osceola, Franklin, Quincy, Hancock, Isle Royale, Superior, Baltic, Trimountain, and Champion. The first three named are in Keweenaw county and all of the others are in Houghton county. Still farther to the southwest are the Lake, Mass, Adventure, Michigan, Victoria, and White Pine mines. These are in Ontonagon county.History. The exploitation of these copper mines extends back into prehistoric times. There is little evidence that the American Indians carried on mining operations in this region in the period shortly before the advent of white men, but there is abundant evidence that the ancestors of the Indians, •or some earlier inhabitants of the region, carried on primitive mining operations. For example, Mr. S. O. Knapp, Agent of the Minne¬sota Mining Company, in the winter of 1847-48, made the first discovery of prehistoric mining. This event is described as fol ,lows by Foster and Whitney : Location: As shown on the accompanying map, the Mar-quette district extends from the region of Marquette on Lake Superior westward for about forty miles. The main basin of the district is only about three to six miles wide, but outlying areas which are commonly regarded as belonging to the same mining unit, make the north and south dimension of the district about twenty miles. The most productive portion of the area is in the vicinity of Negaunee and Ishpeming, .about ten miles from Mar¬quette, but important mines are found also all along the range as far west as Michigamme and also to the south in the fold known as the Republic Trough. Comparatively recent developments indicate that important ore bodies have been developed in the outlying Swanzy district about eighteen Miles nearly due south of Marquette.The Marquette Iron District.History: The first discovery of iron ore in the Lake Supe¬rior region was probably that made in 1844 near the site of Ne-gaunee by the government surveying party in charge of William A. Burt, sent out under the direction of Douglass Houghton. This surveying party was using a solar compass, invented by Mr. Burt, and noticed variations in the direction of the magnetic needle at various times and places, but. on the 19th of September, while running the east line of Township 47 North, Range 27 West, the needle varied from its normal position until it stood nearly at right angles thereto."

Jan 1, 1920

By I. Iwasaki, K. A. Smith, X. Cheng

A cathodic decomposition model was examined for both pyrrhotite and chalcopyrite in an attempt to analyze the formation of surface coatings in deoxygenated solutions at natural, near neutral pH. Combination potentials established between the mineral-grinding media pair were simulated by applying a potential of -0.455 V (vs. a standard hydrogen electrode) either to a pyrrhotite or to a chalcopyrite electrode. Chronoamperometry was used to study the polarization behaviors and alteration of mineral surfaces through cathodic processes. The kinetic parameters of cathodic decomposition of pyrrhotite and chalcopyrite were determined and compared. It was concluded that pyrrhotite decomposed faster than chalcopyrite in near-neutral pH solutions. Both minerals were then covered by a film consist¬ing mainly of iron hydroxides, which adversely affected their flotation.

Jan 1, 1999

By Felicia F. Peng

Employing the combination of two types of polyelectrolyte might become necessary to enhance both settling rate of the flocs and water clarity of fine particle suspension in high capacity thickening process. The sequence of addition, mixing, type and dosage of polymer affect the overall performance and capacity of a thickening process. For the sedimentation of copper flotation tailings, the cylinder test results showed that two different techniques may be used to achieve the required performance including high settling rate of flocs, underflow density and overflow water clarity, by either adjusting pH prior to anionic polyelectrolyte flocculation, or using a combination of polymer flocculant and coagulant (dual polymer system). The dynamic thickening testing results showed that only the approach using a dual polymer system is able to meet the required performance. The effects of polyelectrolyte dosage, the sequence and numbers of polyelectrolyte addition, feed concentration, bed level and rake speed on the settling rate, underflow solid, and water clarity were studied in cylinder test and a laboratory dynamic thickening unit with various conditions related to the thickening process. The dynamic test is especially valuable in evaluating the addition and mixing of polymers. The results were evaluated quantitatively, using total suspended solids in supernatant, underflow density and capacity per unit settling area per unit time as the measure of effectiveness. The objective of this study is to quantify and explain some of the parameters involved in the sedimentation of copper flotation tailings.

Jan 1, 1997

By Keith G. Wallace, K. M. (Chris) Chmura, Kirk H. McDaniel

The Waste Isolation Pilot Plant (WIPP) performed a feasibility analysis for the purpose of either modifying, supplementing, or replacing its two main mine fans. The WIPP, located near Carlsbad, New Mexico, is a U.S. Department of Energy (DOE) fa¬cility designed to demonstrate the permanent, safe disposal of U.S. defense-generated transuranic waste in a deep bedded salt deposit. Since the centrifugal fans were installed in 1988, multiple op¬erational and performance concerns have been identified. A com¬prehensive engineering study was conducted in 1995 to: 1) Qualify and quantify operational concerns, 2) Evaluate possible alternatives, and 3) Recommend an optimum solution. Multiple system modification and/or replacement scenarios were evaluated with associated cost estimates developed. The study considered replacement with either centrifugal or axial fans. Multiple fan duties are required at the WIPP. Therefore, Variable Frequency Drives and Inlet Vane Controls (IVC) were investigated for centrifugal fans. In-flight adjustable blades were investigated for axial fans. The study indicated that replacing the existing system with two double-width, double-inlet centrifugal fans equipped with IVCs was the best choice. This alternative provided the most desirable combination of: 1) ensuring the required operational readiness, and 2) improving system performance. The WIPP is currently planning to replace the first fan in 1997.

Jan 1, 1997

By Takahide Wakamatsu

In Japan all the complex sulfide ore deposits of base metals have been developed as underground mining. The underground mining cost is extremely high as compared with open pit mining case. Therefore, in milling section of base metal mines, technical advance for maximum recoveries of valuable constitutents in ore has been required all the time to utilize as much ore mined as possible. At present, in mill plants treating "Kuroko", complex sulfide ore, concentrates of copper, lead, zinc, pyrite and barite are produced, with gold and silver contents in the copper, lead, and zinc concentrates being estimated as extra profit. As in the case of Kuroko milling plants, gold and silver contents in the lead and zinc concentrates of lead-zinc ore milling plants are also varuable from an economical viewpoint. Silver content is very high in Kuroko crude ore as well as lead-zinc ore.

Jan 1, 1982

By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By N. S. Tanious

Three water-spray ventilation systems were tested to determine their effectiveness in diluting methane at a room and pillar coal mine face. These systems were: the shrouded nozzle, the spray fan, and the conventional nozzle system that is normally supplied with continuous miners. The testing was carried out in two stages. The air-moving capacities of the nozzles were first evaluated in a duct. The duct tests showed that: (1) the volume of air moved by the nozzle is controlled by the type of nozzle, the space around it, and its location relative to the duct wall; (2) the spray fan nozzles or the shrouded nozzles are both capable of moving air quantities up to 6400 cfm (3.02 m3/s); and (3) when the shroud was removed from around the shrouded nozzle, its bare nozzle moved between 9 and 65 percent more air. In the second stage of testing, a full-scale evaluation of the three systems was conducted. The three spray systems were mounted on a continuous miner which was positioned in an aboveground simulated mine entry. The simulated face tests gave the following results: (1) the standard nozzle system, which is mounted on the front end of the miner, is inadequate for diluting methane; (2) under the best testing conditions (lowest methane emission rate, highest primary ventilation airflow, and closest brattice set back), the slight improvement given by either the shrouded system or the spray fan is the same; and (3) under the worst testing conditions (highest methane emission rate, least primary ventilation airflow, and farthest brattice set back), the spray fan gave superior performance.

Jan 1, 1987

By E. G. Kelly

Many methods of evaluating the efficiency of mineral separations have been described in the literature. In most cases, these present the efficiency as a single number. This paper argues that single-number measures of efficiency are inadequate because they seldom allow reliable comparisons, and are unlikely to correspond to the optimum operating conditions. To realistically make comparisons, it is necessary to distinguish the particle-induced component from the equipment- induced component. In a number of situations where the separating property can be identified (such as in gravity concentration and classification), this can often be done. However, in more complicated systems, such as flotation, this has not been possible. A technique is described whereby particle and equipment contributions to efficiency can be distinguished without the need to define the separation property. Because smelter schedules define a boundary correlation between grade and recovery, they generate some of the better single¬-number measures of efficiency, but even these have limitations.

Jan 1, 1991

By Alphonse C. Van Besien

It is doubtful if anyone knows precisely how many U.S. coal miners have been killed by roof falls. In 1931, 51 miners were killed by falls or rock in every 10-day period (Paul, 1). Roley (2) indicated that from 1882 to 1945, over 3,700 men were killed by roof falls in Illinois alone. During this 63-year period, roof falls accounted for 47.6 percent of the coal mine accident fatalities. Similarly, Price and Nolting (3) reported that during the 10-year period beginning in 1937, roof falls claimed over 50 percent of the lives lost in West Virginia coal mines. Over 150 miners died under fallen rock in West Virginia alone, in each of those years. Certainly, the statistics have improved over the years. The Theodore Barry Report of 1971 (4) indicated that during, the previous 4 years, fewer than 100 miners per year lost their lives to roof falls--in the entire United States. The improvement is less remarkable when it is realized that the accident rate has declined only as the number of miners employed underground has declined. The TRW Report (4) pointed out that although the number of fatalities in underground coal mining (from all causes) had been cut in half from 1956 to 1965, the fatality rate per thousand miners had remained virtually constant. The decreased accident rate was due not to a reduction in hazards but to a reduction in work force. Roof falls still claim approximately 50 percent of the lives lost in underground coal mines. Considering the increased underground production of coal which must result from our energy crisis, the environmentalists' pressure against strip mining, and our inevitably diminishing supply of strippable coal, it is apparent that some means must be found to reduce the number of lives claimed by coal mine roof falls.

Jan 1, 1973

By B. M. Davis

Frequently ore reserve studies utilize samples from a series of drilling or other types of sampling campaigns. Sometimes there is a mix of sample types, say for instance, channel samples from underground workings and drill core from surface drilling. Since drilling, sampling, logging, and sample preparation protocols tend to change over time, there is always a question whether the samples give similar results that vary only by chance alone. The problem of comparing the various campaigns is complicated by the presence of spatial relationships among the samples. Geostatistical tests take the spatial correlation into account, but these tests may be complicated and cumbersome to implement. A comparison method based on familiar statistical graphics is presented which accounts for the spatial relationships. It allows comparisons to be made quickly by using simple statistical graphs such as scatterplots and boxplots.

Jan 1, 1996

By M. A. Trevits

Methane gas is an ever-present problem facing the underground coal mining industry. Mine operators are mandated by law to maintain safe concentrations of methane gas within the mine environment. Ventilation is used by all mine operators to reduce gas concentrations. When gas emissions are such that ventilation alone cannot control the gas, methane drainage methods must be applied. In the 1970's, there was a concern that unprecedented gas emissions would be experienced as mine operators began new developments in the deep and gassy Blue Creek and Pocahontas No. 3 Coalbeds. Such concern was well founded as significant gas volumes are being emitted as these mining operations mature. In 1971, the O.S. Bureau of Mines began to tabulate and track the change in mine emissions. These data were extracted from the underground coal mine quarterly air quality-quantity inspections performed by the Mine Safety and Health Administration. At that time, approximately 6,560 Mm Id (231.7 MMcfd) of gas were emitted from 430 underground coal mines. It was determined however, that the mines emitting less than 2,832 m Id (100 Mcfd) contributed only 2.1 percent of the total daily gas flow. In 1988, underground coal mines emitted 8350 Mm Id (294.9 MMcfd) from 487 mines. This report evaluates the changing nature of U.S. mine gas emissions over the last 18 years.

Jan 1, 1993

By Thomas D. Henderson, Wayne R. Hopkins, Arthur Lynch

Brought on stream in 1975, the 10,000-tpd copper oxide processing plant of Anamax Mining Co. at Twin Buttes, Ariz., is now at design capacity. Based on acid agitation leaching, solvent ion exchange, and electrowinning, the Anamax plant bridges an ore to metal gap by producing mine site copper cathodes from lime-bearing oxidized feed. Until recently, such material was considered submarginal because copper recovery was uneconomic using conventional methods. The new Twin Buttes oxide plant, the largest of its kind in North America, is one stage in the expansion of Anamax, a joint venture owned equally by The Anaconda Co. and Amax Inc. Other phases include an accelerated stripping program in the pit, an expansion of the sulfide concentrator from 32,000 to 40,000 tpd, and construction of a 57 MW diesel generating station for the oxide plant. Once the basic oxide flowsheet had been selected, Anamax awarded a contract to a joint venture of Arthur G. McKee and Co.' (Western Knapp Engineering Div.) and Davy Powergas' (a member of the Davy International group). The contract covered process design, engi¬neering, procurement, and construction. The property, identified by Banner Mining Co. in the 1950s and developed by Anaconda, was put into production in 1965 as an open pit mine equipped with a 32,000-tpd sulfide flotation concentrator. The Twin Buttes mine has since become one of the largest earth¬moving operations in the world, excavating some 10 tons of waste per ton of ore. At the original sulfide mill feed rate, the mining fleet moved 375,000 tpd of material (225,000 tpd alluvium and 150,000 tpd of rock). The most significant copper mineral in the ore body is chalcopyrite, which is scattered in pockets of weak mineralization. Cap oxidation has progressed to depths of several hundred feet, principally as chryso¬colla. Present estimates place minable reserves at 347 million tons of sulfide ore grading 0.6% copper and 55 million tons of oxide ore grading 0.9% recoverable copper from 1.25% total copper. The nonre¬coverable fraction is composed of native copper and insoluble copper oxides and sulfides. The oxide ore has been partially stripped, and 25 million tons having a copper content greater than 0.6% have been stockpiled. The stripping of oxide ore to expose sulfides will continue for several years. The copper oxide studies launched by Banner Mining in the 1950s were continued by Anaconda. Process routes investigated included: acid leaching; cyanide leaching; segregation roasting; chlorination, oxidation, and reduction; flotation in combination with leaching; cal¬cite flotation plus leaching; copper oxide flotation; and sulfidization flotation. By late 1968 test work favored a low temperature roast followed by an ammonia-ammonium carbonate leach of the ore. During this period, acid leaching had been discounted because of excessive con¬sumption of expensive acid by the high limestone ore matrix. Restric¬tive SO2 emission controls in 1971 suggested an increase in acid production and availability at a lower cost. Research then switched to acid leaching, resulting in development of the present flowsheet. Mechanically agitated leaching in a cascaded series of tanks was selected in preference to vat leaching after piloting a stirred-tank system. Preliminary testing of vat leaching indicated that inter-vat solution clarification problems could be caused by gas evolution from leaching of high limestone ore. Lack of research time for resolution of this problem, plus success of the pilot agitation system, dictated the final flowsheet selection. In 1973 Anamax decided to proceed with agitation acid leaching to produce liquor for a solvent ion exchange (SIX) and electrowinning (EW) installation. At that time, such systems were proving their eco¬nomic viability, and they allowed direct recovery of high quality cath odes while bypassing the smelting and refining of lower grade cement copper precipitates. The Anamax oxide plant processes a grade of about 1% recovera¬ble copper from stockpiled sources and new mine production. Unit operations include size reduction, agitation leaching, countercurrent decantation washing of residue, pH adjustment of solution, clarifica¬tion, solvent ion exchange, and electrowinning of copper, with a mini¬mum of environmental impact. The feed to the primary crushers is composed of: 1) -10-in. stockpiled oxide ore, which has passed through the mine crusher and has been grade-segregated to allow blending for a consistent feed material. 2) -4-ft ROM oxide ore stockpiled in the early mining operation. 3) -10-in. ore currently being mined. Ore is delivered in 100-ton mine trucks to a dump pocket, where an apron feeder transfers the ore to an inclined 6-in. fixed grizzly. The grizzly oversize passes through a 48 x 60-in. primary jaw crusher, where it is reduced to -6 in. The fixed grizzly undersize and the crusher product are conveyed to a 4-in. vibrating grizzly. Oversize from this grizzly reports to a 7-ft Standard cone crusher, where it is reduced to -1 in. The crusher product and the vibrating grizzly undersize are conveyed to two 8 x 20-ft double-deck vibrating screens that operate in closed circuit with two 7-ft short head cone crushers. Screen undersize (-1/2 in.) is delivered to a covered conical fine ore stockpile with a designed live capacity of 15,000 st. Beneath the stockpile are ten Pioche belt feeders feeding two conveyor belts that supply the milling circuits. Each group of five feeders has two variable speed and three fixed speed drives. Grinding to 95% -48 mesh is done in two parallel lines, each consisting of an 111/1 x 18S -ft rod mill and a 12' x 30-ft ball mill in open circuit. Water is added to maintain a pulp density of 72% solids in the mills. Further water additions are made to the ball mill discharge sumps to create a 60% solids slurry for transfer to the leaching reactors. The nominal grinding rate is 440 tph solids. A splitter box on the product from the mills separates the incoming slurry into two streams, one for the leaching reactors and the other for the pH adjustment reactors. This split is determined by the lime content of the feed material and the pH of solution entering the pH adjustment reactors. Five 30-ft diam by 31-ft high mechanically agitated, rubber-lined leach tanks are arranged in cascade, with gravity transfer from one to another by enclosed launders. Concentrated sulfuric acid (93.2%) from storage is added to the pulp principally at the first leach tank, and the reacted slurry leaving the cascade after a 5-hr residence time at 50% solids is pumped to countercurrent decantation (CCD) wash thickeners.' At this point, most of the soluble copper has been leached from the ore. Up to 250 lb of sulfuric acid is consumed per ton of ore, the acid being supplied in tank cars from a nearby smelter. Leached slurry at 50% solids is washed countercurrently with return raffinate from the SIX plant in a series of four Doff-Oliver center caisson, 400-ft diam thickeners. Solids advance from thickeners No. I through No. 4 to the tailings dam. SIX raffinate advances from thickener No. 4 to thickener No. l, where a pregnant solution overflows to pH adjustment. To optimize solvent extraction, the pH of the pregnant solution from CCD thickener No. I is adjusted from 1.5 to 2.5 by adding unreacted ground slurry (split from the grinding circuit) to the preg¬nant overflow of No. I CCD thickener. The resulting slurry is retained for 45 min in three mechanically agitated leach tanks. CO, generated during leach and pH adjustment is ducted to a demister, where entrained acid droplets are removed. The cleaned CO, and the No. 4 CCD underflow are the only effluents from the plant. The underflow of No. 4 thickener mixes with tailings from the sulfide concentrator and is neutralized. Slurry exits the pH adjustment vessels at about 10% solids and

Jan 1, 1985

By John W. Trammell

The White Pine Mine is situated in Ontonagon County, approximately six miles south of Lake Superior, in the Upper Peninsula of Michigan (Figs. 1 and 2). The ore body mined by White Pine Copper Company occurs in the lower - most 20 to 25 feet of the Nonesuch formation, a series of middle - to upper Keweenawan shales, siltstones, and sandstones. This paper elaborates on the solution of problems in ore grade control and mine planning arising from the presence of an essentially barren sandstone stratum within the ore body.

Jan 1, 1964

By John D. Abbatt, H. B. Newcombe

The Eldorado Epidemiology Project formally began in late 1978. It consists of a retrospective cohort study, to be followed by and to leave in place the mechanism for a prospective cohort monitoring program. These present and future activities are intended to merge into one another. The current study includes all Eldorado employees past and present for whom records are available. The objectives of the undertaking are - 1) to obtain cause-and-effect, dose-response data with which to evaluate the risks to workers in radon and radon daughter containing atmospheres, and to provide additional quantitative information on which to base possible improvements in working conditions; 2) to identify any dead employees of E.N.L. whose cause of death suggests that a potential compensation claim right exists for their survivors, and to similarly identify living ex-employees of E.N.L. whose work histories and states of health suggest potential compensation claim rights. The nature of the study has been determined by the history of Eldorado, which will be described briefly. Eldorado began (as Eldorado Gold Mines) in the late 1920's, operating relatively unsuccessfully in Manitoba. In 1930-31 the emphasis shifted from gold to uranium, and what became the Port Radium mine was first staked. From the very early 1930's until 1942 the company's purpose was almost entirely devoted to the mining of pitchblende - uranium ore - from the Port Radium site, and the extraction of radium from this ore at the Port Hope refinery in Ontario, a process from which the uranium was a largely stockpiled byproduct. In 1942 the company was requested by the Canadian government to accelerate the production of ore, and of uranium. At the same time, all its stockpile of uranium (then being stored in silos as a "waste" product which might later "come in") was acquired for the Manhattan Project. In 1944 the company shares were bought by the Canadian government and the company itself became Eldorado Mining and Refining (1944) Limited; it has operated as a Crown corporation ever since. The name was changed to Eldorado Nuclear Limited in 1968. The Port Radium mine was closed in 1960, and since 1953 the Beaverlodge uranium mine complex has operated on the northern shores of Lake Athabasca. Radium production was discontinued in 1954 when the radium circuit was taken out of the refinery, and since that time the refinery's products have been various types of refined uranium (uranium dioxide, uranium trioxide and uranium hexafluoride) plus metallic uranium. The refinery, of course, processes the company's own yellowcake but as it is one of only five refineries in the western world, most of its work consists of custom refining for a large variety of customers. In addition to the mining and the refining operations, there have been three other divisions involved, and these are, respectively, Aviation, R. & D., and Exploration, in addition to the financial, corporate, administrative and support activities. MATERIAL AND METHODS The human study population (Abbatt, J.D. et al., 1980) consists of all Eldorado employees who have ever worked for E.N.L. and for whom records are available. The total nominal roll to December 31, 1980 is approximately 21,000 and consists of employees of the Mining, Refining, R. & D., Aviation and Exploration Divisions, as well as a relatively small number of other employees (Head Office, etc.). The basic epidemiologic design, as has been mentioned, is for a retrospective cohort study merging into a prospective cohort study, and depending very heavily upon automated record linkage and file matching. The original design called for an initial assembly of the bulk of the nominal roll, prior to the first of two matches with the National Mortality File. The first of these two matches is now complete and the second is to be separated from it in time by 18 months to two years. The second match will be the definitive match prior to the analysis, the interpretation of the data, and the publication of the results. The period between the two matches was designed to permit completion of the nominal roll, including: - addition of previously missing individuals and groups, elimination of duplicates, and amplification of identifying information, - assembly of work histories and exposure information for each individual, - digestion and application of lessons and modifications resulting from the first match with the National Mortality File at Statistics Canada, - preparation of files for the definitive second match. The analysis of data will be carried out by the National Cancer Institute of Canada Epidemiology Unit based at the University of Toronto. This group, with whom Eldorado Nuclear has a memorandum of understanding, have for a variety of reasons, including their familiarity with record linkage and their expertise, been members of the Project Team since

Jan 1, 1981

By John J. Simmons

Federal and state regulations, in an effort to reduce acid rain, are becoming increasing more stringent regarding sulfur content of fossil fuels. Whereas coal from Illinois and Ohio contains between 1.5 and 37. sulfur, Wyoming has billions of tons of coal reserves containing less than O.5/% sulfur. However, much of Wyoming's coal is high in moisture, 25-30%, and low in energy, 8,200 to 8,900 BTU, vs. 10,500+ BTU for Illinois and Ohio Basin coals. Wyoming currently ships in excess of 130 million tons of coal to utilities as distant as Michigan, Wisconsin, Minnesota, Iowa, Nebraska, Louisiana, Oklahoma, and Texas. With freight charges ranging from $12.00-$20.00/ton, more than $1.5 billion a year is spent to transport water. A test shipment of more than 50,000 tons of Wyoming coal has recently been made from Mobil's Cabello Rojo Mine, south of Gillette, by Burlington Northern to the Robert's Bank British Columbia coal terminal for transloading and shipment to Taiwan Power Company. This coal assayed approximately 8,400 BTU/lb. and 277. moisture. Carbontec Energy Corporation feels the future availability of dried coal containing 11,000+ BTU could substantially enhance the potential export as well as the domestic market for Wyoming coal. Large and small utilities, along with industrial furnaces require tens of millions of tons of Wyoming, West Virginia, and other Eastern The federal government is encouraging the expenditure of hundreds of million of dollars by utilities to retrofit their boilers to allow them to burn higher sulfur coal. This program, if carried to completion, would cost many tens of billions of dollars, with the net result being high power costs sustained by the utilities. These costs will be passed on and paid for by the American consumer. A sensible alternative would be the burning of low sulfur coal by the utilities, without costly capital and operating expenditures. Carbontec Energy feels the successful operation of a Carbondry Coal Drying Plant in Wyoming will be the forerunner to other similar and larger operations in the State. CARBONTEC RESEARCH The activities of Carbontec's personnel to develop a viable process to dry sub-bituminous and lignite coal goes back some 14 years; in fact, for some of us it goes back to 1962 when we were involved in a program under which lignite coal from Truax Trayer's Velva mine was placed in low sulfur coal, which are burned alone or blended with high sulfur, Midwestern coal. Carbontec believes that a high BTU. reasonably priced, shippable coal fuel produced in Wyoming and delivered at costs competitive with Eastern low sulfur coal, could offer substantial benefits to Wyoming coal producers, utility customers, and to the State of Wyoming.

Jan 1, 1989

By F. Villanueva Araya

Compania Minera Disputada de las Condes (CMD), an Exxon Affiliate stewarded by Exxon Coal and Minerals Company, operates two' copper mines and concentrators in Chile. Pilot plant studies were initiated in 1984 using a 0.46 meter and a 0.91 meter diameter column for the cleaning of copper concentrate. This eventually resulted in the design and operation of a rectangular column with an area of 6.8 m2, increasing the final copper concentrate grade from about 30% to almost 32'1. Cu without any loss in copper recovery. Evaluation studies of a commercial bubble generator as against the previously used filter cloth air sparger are now in progress.

Jan 1, 1990

By R. J. Roman

Several research groups have spent considerable time investigating coarse ore leaching. Coarse ore leaching includes both heap and dump leaching. The term "heap leaching" will be used in this paper to mean any type of coarse ore leaching. The results of their work are generally in agreement: the progress of leaching can be described by the shrinking core model and its associated mathematics. The shrinking core model is schematically represented in Figure 1. The particle is contacted by the leach solution on its outside surface and the solution migrates inward by filling the pores on the particle. As the solution filling the pores comes in contact with grains of ore mineral the reagent chemically reacts with the ore mineral to solubilize the contained metal. The solubilized metal diffuses outward to the particle surface and is washed away by the leach solution flowing past the particle. The leaching reagent in the particle interior is replenished by the diffusion of reagent from the surface of the particle to the interior of the particle. If the chemical reaction between the leaching reagent and the ore mineral is relatively fast compared to the diffusion rate of the reagent in the solution in the pores of the particle then leaching proceeds from the outside of the particle inward resulting in a "shrinking core" of unreacted ore surrounded by a "leached shell". The rate controlling step in the reaction can initially be a chemical reaction between the reagent and the ore mineral grain, however, once the reaction has proceeded a short period of time diffusion of the leaching reagent through the leached shell will be the slow step and hence the rate controlling step in the production of metal from an ore particle.

Jan 1, 1992

By H. F. King

In 1982, the result of a 3-year study was published as "A Guide to the Understanding of Ore Reserve Estimation". The principal conclusions were that computation formed only part, and perhaps not the most important part, of an estimate and that the estimate should recognize that geological, mining engineering and metallurgical judgements formed an essential part of the eventual result. Subsequent expectation/realization studies of several mines have produced case histories which form the principal basis of the paper. These imply that geology is even a larger factor than thought earlier.

Jan 1, 1985

By S. M. Hsiung

Chain pillar design formula under weak roof condition was developed by statistically analyzing the results from the three dimensional finite element parametric analyses. The parameters such as mechanical properties of the roof and floor strata, overburden depth, panel width and length, and coal strength were incorporated in the formula. A conversion formula which transfers a rectangular chain pillar into a square chain pillar of equal strength is proposed. The influence of high in-situ horizontal stresses, which are often encountered in the coalfield, on chain pillar stability is also discussed.

Jan 1, 1984