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By William F. Lawrence

Due to the immense volumes of municipal wastes that are generated in this nation every year, it has become imperative that effective measures be developed to utilize all or portions of these materials in a productive and efficient manner. Nany larger municipalities employ incineration to reduce the bulk of material to be removed to disposal sites: but, even after incineration, approximately one-fourth of the original waste material in the form of non-combustible residue and flyash remain. It has been found that the flyash fraction of the incinerator residue contains varying amounts of copper, silver, gold, aluminum, iron and titanium along with many other important metals. The serious loss of these natural resources contained within the flyash and the monetary expense of current disposal methods present a serious and rapidly increasing problem to the public and to the Federal Government. The growing emphasis upon environmental control, rapidly increasing urban population and "convenience packaging" 'with its attendant increased waste production has led to the initiation of a broad range of studies concerning municipal waste. For example, the U. S. Bureau of I-lines is investigating all aspects of the incineration of municipal waste under its solid waste disposal program.l "In-house" studies range from improved incinerator design to potential recovery and recycling of the various components within the incinerator residue. A major portion of the latter research is being performed at the U. S. Bureau of ;-lines, College Park, Maryland facility while grants for specific areas of study have been issued to several universities and research institutes. In addition, the Department of Health, Education and Welfare has supported programs examining the potential recovery of paper; glass, metals and other components of unincinerated refuse followed by disposal in sanitary landfills.

Jan 1, 1971

By E. A. Schmidt

In 1962, while investigating the tungsten potential in the Thompson Creek area for Union Carbide Corp., Henry T. Eyrich obtained anomalous concentrations of molybdenum (up to 19 ppm Mo) from stream sediment samples collected along Pat Hughes Creek, a tributary to Thompson Creek. Follow-up investigations by Eyrich during the same year disclosed the presence of a hydrothermally altered and molybdenite-bearing granitic exposure in the headwaters of Pat Hughes Creek about 2.4 km (1.5 miles) upstream from the sample point. The altered outcrop measured about 549 by 305 m (1800 by 1000 ft), contained abundant northwest-trending quartz veins locally with molybdenite, and was overlain and surrounded by the Challis volcanic extrusive sequence of mid-Tertiary age. Eyrich's subsequent soil and rock chip geochemical sampling program over the altered outcrop area showed a general increase in the molybdenum content to the northwest toward the postmineral Challis volcanic cover. Over a distance of about 305 m (1000 ft), values in the altered intrusive rock increased from less than 5 ppm Mo to over 600 ppm Mo, with aconcomitant intensity of the hydrothermal alteration pattern. Based on these geochemical results and on preliminar; megascopic hydrothermal alteration studies, Eyrich correctly concluded that the altered outcrop is an erosional window and that molybdenum mineralization of sufficient grade could exist beneath the Challis volcanic cover some distance further to the northwest. His recommendation to test the area of potential molybdenum mineralization with several drill holes was, however, declined by Union Carbide in 1962. The property remained unclaimed until 1967, when Eyrich joined Cyprus Mines Corp. Acting on Eyrich's recommendation, Cyprus immediately staked 26 lode mining claims to cover not only the altered intrusive outcrop, but also the northwest trend of mineralization where it projected beneath Challis volcanic rocks. Eleven of these claims are now patented.

Jan 1, 1991

By Klaus Grebe

Recent investigations by Mannesmann Research have shown that disintegration of sinter with widely varying quality levels is subject to a much more complex set of influences than was previously assumed. It was confirmed that disintegration at temperatures between 300 and 600 °C is triggered by the degradation of hematite. The specific tendency of hematite to disintegrate is, however, not constant: sinters containing the same amount of hematite can show widely differing disintegration tendencies depending upon structural homogeneity, porosity and gangue quantity. Especially sinters with heterogeneous structures transmit stress to the hematite phase and this favours disintegration. Another factor is the brittleness of the glassy silicate phase, which can be influenced even by small modifications of minor gangue constituents.

Jan 1, 1981

By Xinyang Wang

It is believed that sulfur and volatile matter contents in coals are the main intrinsic properties to cause the self-heating of coal. Their oxidation at lower temperatures than that of fixed carbon to initiate coal's self-heating should be quantified. This study is aimed to improve the previous mathematical model developed by the authors for studying the coal's propensity for spontaneous combustion. It enhances the model's ability to consider the effects of sulfur, volatile matter and moisture contents in the coal - three important factors affecting the coal's self-heating process. Sulfur exists in coal primarily in the form of pyrite which will be oxidized rapidly under suitable conditions. Volatile matters, higher in low rank coals, are more easily to be oxidized. Heat of water condensation provides initial energy for low temperature oxidation. Heat release rates for pyrite oxidation and moisture condensation are built in the model. Finally, the simulation results in terms of time versus temperature on pyrite oxidation and moisture condensation are validated with the adiabatic testing data respectively.

Jan 1, 2013

By D. N. Madge

A description will be given of an approach to plant process automation based on the strategy of first developing an information system followed by later development of process models. The advantages of this approach over the unit process control will be outlined. Some insight will be given into the risks associated with this method. These risks can be shown to be primarily economic.' Some reference will be made to two Cominco installations in the interests of illustrating the generality of the information systems approach.

Jan 1, 1972

By Herry Poernomo, Agus Prasetya, Iga Trisnawati, Himawan Tri Bayu Murti Petrus, Gyan Prameswara, Panut Mulyono

We studied yttrium extraction from untreated zircon sand processing waste tailings. Xenotime mineral as the rare earth element source with an abundance of yttrium (REY) was detected in sufficient grade for extraction. As much as 9.03% yttrium exists in zircon tailings (analysed by X-ray fluorescence spectrometry). The presence of yttrium was confirmed by X-ray diffractometry in xenotime minerals, which are yttrium carriers (Y-PO4). The purpose of this research was to determine the effect of leaching conditions on yttrium recovery from zircon sand after alkaline fusion treatment. Alkaline fusion was chosen to decompose phosphate into hydroxide in the xenotimemineral,which will reduce further required hydrometallurgical processing.Alkaline fusionwas carried out for 3 h at 450 °C, with a ratio of sodium hydroxide to zircon tailings sand of ~ 1:1. The alkaline fusion product was leached with water, followed by hydrochloric acid treatment to leach the yttrium. Yttrium recovery reached 87% under optimum conditions (60 °C, 1MHCl, and solid-to-liquid ratio = 1/10 for 7.5 min). A suitable model for yttrium dissolution with hydrochloric acid was diffusion through a solid particle ash layer. The calculated activation energy (EA) for this model was 20.21 kJ/mol.

May 26, 2020

By John T. Atkins

Utah International's Shirley Basin Mine is located in the Southeastern High Plains of Wyoming at an elevation of about 7,100 feet. The Shirley Basin area is a gently rolling semi-arid land. The uranium ore zone occurs at depths ranging from 270 to 350 feet below the surface. Overburden consists of poorly consolidated sandstones, siltstones, and clays. Underground mining was begun in 1960. Conditions were made difficult by high rates of water inflow and ground support problems. The development of a solution mining I or in-situ leaching technique allowed underground operation to be abandoned in 1964. Solution mining provided limited production from 1963 to July 1970. In January, 1969, large scale development of open-pit mining was begun. A 1,200 ton per day mill went "on stream" in March, 1971.

Jan 1, 1973

By R. Owens, C. Roos, B. Hill, S. Rosenthal

It is common practice, and even legally required in many jurisdictions, to design two-lane haul roads in open pit mines to three and a half times the widest dimension of the haul trucks operating on the road. In open pit gold mines with high strip ratios, the road width has a significant impact on the economics of a design. It is possible to minimize the flattening of the highwall if the road width is reduced, assuming the width used is not needed to flatten the slopes for geotechnical purposes. With the use of autonomous haulage and pullouts, it may be possible to operate safely and efficiently on a reduced road width of two times the width of the hauling equipment.

Mar 2, 2022

By John H. Bassarear

Large scale equipment can reduce both capital and operating costs for concentrators. Large semiautogenous mills can be effectively utilized in big mills but they must be kept fully loaded to take advantage of their great potential. With varying hardness ores, controlling throughput is difficult and careful design must take this into consideration. A variable speed drive on large primary semi- autogenous mills can overcome this problem to a considerable degree. This paper provides a base case using standard accepted technology of two wound rotor single speed 4,476 kW (6,000 hp) motors to drive the primary SAG mills of a hypothetical 36,364 mtpd (40,000 stpd) typical porphyry copper mill. One other fixed speed drive will be described and five variable speed drive configurations will be reviewed. Capital costs for the entire grinding circuit, from crusher product to flotation feed, were developed for the base case and each of the alternatives. A technical evaluation comparing the advantages and disadvantages of each drive is presented.

Jan 1, 1985

By Donald R. Williamson

The Cyprus Pima Mining Company's open pit operation was started in 1957 with the normal assumption that the ore reserve estimate was accurate within an acceptable range of values. At that tine reserves were law - volume (+10 million tons) and high grade (+1.5% Cu). Several periods of growth have brought the operation to its present production level of 55,000 tons per day of ore and +125,000 tons of waste. Over the years, a number of attempts to refine and polish the ore reserve figures have been made using a variety of established standard procedures. None of these attempts have produced results which are totally satisfactory, particularly when comparisons are made between ore reserve estimates and short to medium range production figures. Other refinements were tried using adaptations of established methods combined with computer modeling techniques. The resulting mine-model ore blocks were an improvement but they still tended to show too match tonnage of mill feed, often at too low a grade. One attempt at Kriging had only limited success. Statistical studies of operations data revealed relationships which could be used to subdivide ore blocks in much the same way as they are actually mined. Comparisons thus far between these newly developed ore blocks and production data show that the new method of evaluation yields excellent results. A new, better procedure for estimation of ore reserves apparently has been developed for the Cyprus Pima deposits. The specific procedures seem likely to have direct application to other, similar operations. The methods used to develop the procedures could be of value to still other operations.

Jan 1, 1976

By James N. Orlich

Meridian Gold Company's Royal Mt. King mine (RMK), a subsidiary of FMC Gold Company, is located approximately 40 miles east of Stockton, California on the western edge of the historic mother lode belt. Gold ore at RMK is associated with a series of northwest trending fault zones hosted in either meta-volcanic or meta-sedimentary rock types commonly referred to as greenstone sulfide and carbonaceous phyllite. Mineralogy studies show gold in both ore types to occur primarily as electrum in the form of 1 -20 um inclusions in pyrite with varying amounts of free milling gold. In carbonaceous ores, gold was not found to be closely associated with carbon. At the start of the current mining operation, three major ore resources were identified. North Pit and Gold Knoll pits were the major greenstone reserves while Skyrocket Pit contained over 60% of the total reserves as carbonaceous ore. Since North Pit and Gold Knoll were to be the first areas mined, the mill began operation in February 1989 with a flowsheet designed for greenstone ores with a carbonaceous expansion anticipated after depletion of the greenstone reserves.

Jan 1, 1992

By David G. Kidd

The principles of Ore Sorting in general and photometric sorting in particular are discussed with detailed reference to the Model 16 machine. The operating principles and design of the Model 16 are discussed and a case study presented to demonstrate the use of sorters to process an industrial mineral. In addition, an economic study is presented, to highlight the benefits of ore sorting, by evaluating the increased revenue that can be achieved using an ore sorter to process a particular ore body.

Jan 1, 1983

By John C. Nichols

The Ouachita, Ozark-St. Francis, and Mark Twain National Forests are located within the Ozark and the Ouachita Mountains regions of southern Missouri, central and northern Arkansas, and eastern Oklahoma- an area generally referred to as the Ozark-Ouachita or Interior Highlands. These national forests share many similarities in geography, biotic resources, and areas of shared public, market, and user interests. The three national forests have joined together to evaluate the natural resources within the Highlands and to study the impacts of the forests on local and regional economies and on the people who use these national forests and their resources. Other federal, state, and local government agencies and universities have joined with the Forest Service in this project, broadening the assessment area beyond the national forest boundaries to include 107 counties in Arkansas, Missouri, and Oklahoma. The diverse, resource-related data for the as- sessment area address social, economic, terrestrial, and aquatic issues and conditions throughout the Highlands. The assessment report currently is being developed by the Forest Service and will include an accoMpanying CD-ROM. The Ozark-Ouachita Highlands assessment has provided a forum for minerals- and geology-related resources, industries, social significance, and potential to be profiled and discussed along with other critical assessment resources. The state geological agencies in Arkansas, Missouri, and Oklahoma were active contributors to this effort. Their reports and other key data bases developed and used for the minerals sections of the published assessment report will be available either as a supplemental CD-ROM or on the Internet. Updated information on the progress of the assessment report can be found at the Ouachita National Forest Website, at <http://www.fs.fed.us/oonf/ouachita.htm>. This will also include a matrix displaying the minerals in the assessment area, their occurrence by assessment-area counties, and a potential ranking, as well as information on interesting geologic features. The state and federal reports, resources, and data bases, have helped answer key questions concerning mineral resources and related issues in the Highlands:

Jan 1, 1999

By J. Philip Bromberg

Any effectuation of a viable program of air pollution control will&apos; require a basic understanding of the chemistry of atmospheric processes, the fate and migration of SO2, and the ultimate health effects of the sulfur oxides themselves. The migration and ultimate fate of Sop discharged into the atmosphere will be discussed with particular emphasis on the various conflicting views and how these relate to our present emission control regulations. Our present lack of knowledge is a severe detriment, and consideration will be given to those particular areas where future research may best be directed to provide the answers to a number of perplexing questions.

Jan 1, 1976

By James M. Barker

The borate deposits of Death Valley, first exploited in 1882, were a major world source of boron-bearing products until shortly after very large deposits near Boron, California (Kramer) came on stream in 1926. Borate mining in Death Valley then became moribund, with most extant deposits held as reserves backing Kramer. Tenneco began borate operations following acquisition of claims in 1967 and is the only active borate producer inside Death Valley National Mon¬ument. Borate ores, confined to the lower Furnace Creek Formation (early to middle Pliocene), initially precipitated in a shallow lake fed by boron-rich volcanic hot springs associated with the Basin and Range province. Later diagenesis and remobilization plus regional folding, faulting, and erosion altered and complicated originally simple borate mineralogy and distribution. Dominant ore minerals are ulexite (NaCaB509.8H20), probertite (NaCaB509.5H20), and colemanite (Ca2B6011.5H20). Tenneco currently operates two open pits near Ryan, California. Mining is by bulldozer-pushed scrapers hauling waste and front-end loader/off-highway truck combinations for ore and waste haulage. Archeological, environmental, and aesthetic considerations influence both mine plans. Boraxo Pit, opened in 1970, has deep ore (stripping ratio 10 to 13:1) which occurs along a fault and as beds in shale on the hanging wall. Problems include pit wall instability (adverse bedding), non¬ripable calcareous shale (11,000 fps), and mixed ore separation. Sigma Pit, opened in 1975, has shallow (stripping ratio 1 to 1). low dipping, bedded ore in shale. Massive ulexite and "granular" un- cohesive colemanite present problems in mining and milling. Colemanite is crushed to -6 inches, washed, dried at 350°F, then calcined at 800°F with 100 tpd capacity. Crushed (-6 inch) ulexite or probertite is Raymond milled to customer specifications with a capacity of 1000 tpd.

Jan 1, 1976

By VIVEK KASHYAP, PATRICK TAYLOR

The presence of iron in sphalerite induces the formation of zinc ferrite (ZF) during roasting. Zinc refinery residues (ZRR) and electric arc furnace dust (EAFD) contain a significant amount of zinc ferrite and are difficult to treat. According to the literature, up to 71% ZF has been found to be present in ZRR. The dissolution of zinc ferrite is inefficient even in aqua regia. The Waelz process is an existing process to recover zinc from electric arc furnace dust by reduction using carbon monoxide followed by the volatilization of zinc. In this study, zinc was selectively extracted from zinc ferrite by partial reduction of zinc ferrite at relatively low temperature followed by leaching in dilute sulfuric acid (20 g/L). Pure zinc ferrite samples were partially reduced to zinc oxide and magnetite by using H2 gas as a reductant. The presence of magnetite and zinc oxide was confirmed by X-ray diffraction (XRD) technique. A maximum of 95%zinc extraction was achieved after leaching with sulfuric acid on the reacted samples. The paper also discusses the effect of time, temperature, and concentration of H2 in the reducing gas on the extraction of zinc and iron from ZF.

Sep 16, 2020

By D. J. Boron, R. Kollrack

Introduction The US has the largest total (364 Gt, or 401 billion st) and recoverable (182 Gt, or 201 billion st) coal reserves in the world (Steam/Its Generation and Use, 1978). However, progressive coal use in this country has been hampered by environmental constraints, such as restrictions on sulfur dioxide, nitrogen oxides, and particulate emissions. Since the promulgation of the Environmental Protection Agency (EPA) New Source Performance Standards (NSPS), it has become apparent that all US coals will require sulfur removal to comply with the new sulfur dioxide emission levels. If more stringent emission standards are imposed, through pending acid rain legislation or amendments to the Clean Air Act, it is likely that the coal industry, and the industrial and utility sectors will look increasingly toward the development of new and advanced coal cleaning technologies to continue and promote the use of coal. Table 1 summarizes the EPA NSPS for sulfur dioxide emissions. Fortunately, the basis for determining sulfur dioxide removal is the Btu and total sulfur content of the run-of-mine coal. Thus, credit is given for any sulfur removed before combustion, e.g., through physical coal cleaning. Unfortunately, physical coal cleaning alone may be unable to produce a compliance coal with respect to the EPA NSPS. Therefore, all industrial and utility coal-fired boilers will require post combustion gas cleanup, e.g., flue gas desulfurization (FGD), to comply with the new regulations. Table 1 - EPA NSPS for SO2 Emission Category SO2 Content of ROM Coal SO2 Reduction Required 1 0-2 Ib/M Btu 70 2 2-6 Ib/M Btu To 0.6 Ib/M Btu (70% to 90%) 3 6-12 Ib/M Btu 90% 4 12 Ib/M Btu To 1.2 Ib/M Btu Alternatives to post combustion gas cleanup, such as staged combustion, lime and coal injection, fluidized-bed combustion, and chemical coal cleaning, are currently under development. Of these, chemical coal cleaning offers the advantages of providing (1) high quality fuel for combustion (low sulfur and ash, high Btu, and attenuated product variability); (2) a solid fuel product for use in retrofitted gas and oil boilers; and (3) a solid fuel product with multiple end uses, e.g., internal combustion engines, carbon electrodes, coke, or chemical feedstock. Gravimelt process In the Gravimelt process, 1.4 mm x 0 (14 mesh x 0), physically beneficiated coal of about 5% moisture is fed into a 3401 to 3901 C (644 ° to 734 ° F) molten bath of sodium hydroxide and potassium hydroxide. Organic and pyritic sulfur and most of the mineral matter in coal react with and are dissolved in a molten mixture of potassium hydroxide and sodium hydroxide. The reacted coal sulfur is converted to alkali sulfides and polysulfides, whereas the mineral matter is apparently converted to alkali-aluminum silicates and other byproducts that are water and acid soluble. Coal residence time is 60 to 240 minutes. The sodium hydroxide to potassium hydroxide ratio is about 9:1. The total hydroxide-to-coal ratio is no less than 10:1. The specific gravity of the caustic melt is about 1.8 g/cm3 . Thus, coal is buoyant in the bath, and insoluble mineral matter with density greater than 1.8 g/cm3 would sink. Reacted coal is first water-washed with an equal weight volume of water in a countercurrent mode of operation. The coal is filtered and the filtrate sent for regeneration. The coal is then acid-washed in 10% sulfuric acid. The acid-water/coal mixture is filtered to recover the final coal product, and the filtrate treated with lime to precipitate dissolved mineral matter. The precipitate is filtered from the mixture and discarded. The water is recycled to the plant. Filtrate from water-washing is treated to remove sodium sulfide. The sodium sulfide is chemically treated to produce hydrogen sulfide, calcium car¬bonate, sodium hydroxide, and other products. The hydrogen sulfide is treated to produce sulfuric acid for the acid-washing step. The calcium carbonate is treated further or discarded, and the sodium hydroxide is concentrated by evaporation and recycled to the reactor. The Gravimelt process is currently at the bench-scale stage of development. A 0.9-kg (2-lb) continuous reactor is in operation with a continuous water- and acid-washing section adjacent. The regeneration section is not developed to a meaningful extent, although key reactions have been tested in the laboratory. Process conditions, product recovery efficiencies, etc., however, have not been determined adequately and continuous regeneration operation has not been attempted. Work is now in progress to develop each section to a continuous bench-scale operation in processing coal at 9.1 kg/h (20 lb per hr). Microwave process The carbon and hydrogen in coal are relatively transparent to microwave radiation. However, water, caustic, pyrite, and other mineral matter components are strong conductors of microwave radiation energy at the proper frequency. As a result, selective heating of the noncarbonaceous materials occurs that enables the chemical reaction of sulfur and ash with caustic. Although very high temperatures may result from

Jan 1, 1987

By E. Knell

COM Energy, Inc., a subsidiary of Occidental Petroleum Corporation, has constructed a 1.6 million hectoliter-per-year COM plant in Jacksonville, Florida. COM will be distributed to customers in the region as a substitute for fuel oil presently burned in boilers and kilns. As the market for COM grows, COM Energy, Inc. expects to build additional units both in this country and overseas. This paper addresses the overall status and technical aspects of the project, focusing on the facilities for COM production and utilization.

Jan 1, 1982

By K. Marutiram, T. Umadevi, R. Sah, K. Abhishek

Low-grade iron ore received from the West Singhbhum area of Jharkhand state in India was used to develop a beneficiation flow sheet to remove alumina (Al2O3) and silica (SiO2 ) from the ore to get iron (Fe)-rich product for steel plants. Mineral characterization showed that the main mineral constituents present in the head sample are goethite, limonite, hematite, quartz, clay and gibbsite. The head sample contained about 49.2 percent Fe, 8.79 percent alumina and 12.01 percent silica. The received iron ore sample is amenable to upgradation through beneficiation to get pellet grade concentrate. The beneficiation process was established, involving crushing, screening, jigging, spiral separation, size reduction, desliming and magnetic separation. The process showed significant enrichment in Fe values to 63.5 percent Fe in concentrate, from 49.2 percent Fe. From the developed flow sheet, it is possible to obtain pellet grade fines with 63.5 percent Fe, 2.92 percent SiO2 and 2.61 percent Al2O3 with 50.6 percent weight recovery. The overall tailing loss is 49.4 percent with 34.3 percent Fe, 21.28 percent SiO2 and 14.64 percent Al2O3

By S. A. McGhee

Introduction ASARCO&apos;s Ray Operations include an open-pit copper mine in the Mineral Creek Mining District of Pinel County. The district is located in east central Arizona, 132 km (82 miles) southeast of Phoenix, AZ, and 126 km (78 miles) north of Tucson, AZ. The mine lies 8 km (5 miles) north of the Gila River in Mineral Creek Valley, between the Tortilla Mountains to the west and the Dripping Springs Range to the east. The rich copper deposits of the Mineral Creek mining district were first noted by a US Army Officer in 1847. The Ray Copper Company was the first major operation in the area. The company was named after the Ray mining claim, which was one of the earliest recorded claims. After decades of sporadic and unsuccessful mining ventures, the Ray Consolidated Mining Company was formed in 1910. The company introduced one of the first successful block caving operations in the world. In 1933, the Kennecott Copper Corporation acquired the Ray Consolidated Mining Company and continued operation of the mine under the same name until 1943, when the property was renamed the Ray Mines Division of Kennecott. Underground mining continued until 1955, when the mine was converted to an open pit operation. In November 1986, ASARCO Inc. purchased the Ray Mines Division and related facilities from Kennecott. The purchase also included the Hayden concentrator and the Ray Smelter, located approximately 32 km (20 miles) from the Ray Mine. This paper summarizes the 54,400 tpd (60,000 stpd) in-pit crusher and overland-conveying system and the 24,200 tpd (30.000 stpd) copper SAG concentrator and tailings disposal system constructed at ASARCO&apos;s Ray Complex - Ray Operations. General description of facilities The Ray concentrator includes a sulfide SAG concentrator, a "relocatable" in-pit primary-crushing plant and overland-conveyor system, a concentrator tailings-disposal system and plant ancillary facilities, The major elements of the project include: • A relocatable Fuller-Traylor in-pit 1524-mm (60-in) x 2261-mm (89-in) primary gyratory crusher at the 442 m (1450 ft) mine elevation. The crushing plant has a nominal capacity of 54,400 tpd (60,000 stpd) of sulfide ore. • A 1524-mm (60-in) wide x 1253-m (4110-ft) long overland belt conveyor, which runs from the in-pit crushing facility to a transfer station serving both the Ray Concentrator and the existing Hayden concentrator railroad load-out facility. • A 1524-mm (60-in) wide x 152-m (500-ft) long stackerbelt conveyor, which delivers a nominal 27,200 tpd (30,000 stpd) of ore from the transfer station to a 23,600 t (26,000 st) live-capacity coarse-ore stockpile serving the Ray Concentrator. • A nominal 27,200-tpd (30,000-stpd) concentrator, including semiautogenous grinding, flotation, copper concentrate filtration and associated plant ancillary facilities. • A tailings disposal system. Flotation tailings are pumped from the Ray Concentrator to a 122-m (400-ft) diam. thickener, located 4.8 km (3 miles) southeast of the Ray Concentrator site. The thickener overflow is pumped to two 76.2-m (250-ft.) diam. concrete process-water reservoirs. The thickener underflow, at 45% solids, is pumped to a new tailings disposal site located at Elder Gulch. • An HDPE-lined temporary-containment area to collect material from Ray Concentrator plant upsets and local storm-water run-off from a 100-year, 24-hour rain event. •A reclaim-water system from the Elder Gulch tailing pond area to the Ray Concentrator. • A 610-mm (24-in.) diam. x 31.9-km (19.2 mile) long fresh-water pipeline from the Hayden fresh-water-well field system to the Ray Concentrator. The overall process design criteria for the new Ray Concentrator is based on a nominal throughput of 27,200 tpd (30,000 stpd) of ore at 90% plant availability. However, based on the maximum performance capacity of the installed grinding equipment, it may be possible to achieve a sustained operating level in excess of 36,300 tpd (40,000 stpd) when processing soft ores. The balance of this paper will cover the major process systems including:

Jan 1, 1995