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By P. R. Taylor

This work is concerned with the extraction of refractory metals using an electro-reduction process, in which a molten oxide melt, instead of halide, is used as the electrolyte. The process may provide an alternative method of producing refractory metals from their ores through use of a reverse-polarity direct current (DC) plasma arc heating process. In such a process, the plasma torch itself acts as an anode, and resulting liquid or solid product metals work as a cathode, whereas the molten oxide melt plays the role of an electron transfer layer. The focus of this study is: the electro-reduction of scheelite; vanadium pentoxide; and chromite. For the vanadium pentoxide, the proposed melt is mainly composed of CaO-V2O5/Li2O-B2O3-V2O5or FeO-CaO-SiO2-V2O5-Li2O with the addition of other alkali and alkaline metal oxides. In the case of scheelite ore, the SiO2-CaO-Li2O-WO3system is used as melt and metal aluminum or nickel could be used as a collecting media. As for the reduction of chromite ore, SiO2-CaO-Cr2O3-Li2O system with trace amount of CaF2is used as the melt. As a demonstration, chromium was successfully produced at a laboratory scale through the plasma driven electro-reduction process. One of the studied slags had a composition of 13.8% CaO; 4.9%CaF2; 36.4% Cr2O3; 6.9% Li2O; and 38.0% SiO2. Two observations are significant in this laboratory scale study. The first one was that the amount of Cr2O3and FeO, which were predominantly electronic conductors in the oxide melt, greatly affected the conducting mechanism of the melt. With high concentration of Cr2O3and FeO in the melt, the current efficiency was fairly low. In this case, chromium was found to be barely reduced because of electronic conducting mechanism. For example, with initial slag of 60% Cr2O3, no chromium was found reduced. With low concentration of Cr2O3and FeO in the studied melt, typically no more than 30% Cr2O3, it was observed that chromium was readily reduced due to the dominant ionic conducting mechanism. The second observation was that the addition of SiO2tothe melt helped to make the oxide melt more ionic, which was apparently desirable to the molten oxide electrolysis. However, high content of silica had a negative effect on the fluidity of the melt. This was resolved by adding trace amounts of CaF2to the melt. Therefore, by judicious selection of oxide melt, refractory metal oxides might be dissolved into it and yet the melt itself retains the required ionicity. A reverse-polarity DC plasma driven molten oxide electrolysis may prove to be a viable alternative route for the extractive metallurgy of refractory metals.

Jan 1, 2003

By T. Good

Twin roll sizers are neither compactly designed machines nor are they configured to accommodate underground mobility; in fact, twin roll sizers leave room for improvement in many aspects of the design. The DRC1200 capitalizes on the weaknesses of these machines to make a more productive, easily maintained, and compact twin roll sizer than current market offerings. These strides were made through the inception of the Hard Rock Extraction System (HRXS), which requires a mobile sizer as part of the system. After use of a third party sizer on the HRXS, potential improvements were evident. These improvements include improved structural design to reduce overall footprint and an optimized design to reduce operating cost.

Feb 23, 2014

By William Bogan, Mark Vaivoda, Richard Daniele

Many factors must be considered when performing engineering and construction work for a major minerals project. Dravo's Afton copper smelter in British Columbia illustrates some of the E&C factors considered when setting up this new operation. In March 1978, Afton Mines Ltd., a Tech Corp. subsidiary, began blister copper production at the world's first integrated copper ore mining, milling, and smelting operation to use Top Blow Rotary Converter (TBRC) process smelting technology. This key TBRC smelting element is licensed by INCO Ltd. to Dravo Corp. The TBRC process uses a single vessel of modular design. It rotates for mixing and provides complete processing capability, from concentrate through blister copper. The Afton facility is about 16 km west of Kamloops, a town in south-central British Columbia. To make its operation as economic as possible, Afton officials considered several options for the copper smelter, the first in modern times in the province. Inability to extract precious metals-primarily gold-from the concentrates ruled out hydro-metallurgical processes. The final choice was between a conventional electric furnace/ Peirce-Smith converter combination and the TBRC process. The TBRC process was chosen for its versatility in processing the various Afton concentrates and grades and its 40% lower capital cost compared to the electric furnace converter system. The TBRC smelter -including concentrate slurry receiving, drying, storage, smelting, off-gas cleaning, and copper casting-had an installed cost of about $25 million. In December 1974, a pilot program was conducted by Teck, Dravo, and INCO at INCO's Port Colborne, Ontario, research facility. The pilot program demonstrated the TBRC's ability to smelt, convert, and clean copper slags. After the successful pilot effort, Dravo performed additional research to find an effective fixation system for high grade, intermittent SO2 gases. Once the go-ahead was given, design and procurement for the Afton copper smelter was provided by the Denver Division, Dravo Engineers and Constructors. Commonwealth Construction, Vancouver, constructed all Afton facilities. Afton's single furnace smelter is designed to produce 25-kt/a blister copper in 363-408 kg cake form. The product is sold to England for electrolytic refining. The TBRC process is only one of several innovations proposed by Dravo and chosen by Afton to insure the most modern smelter in North America and meet all British Columbia's environmental restrictions. Conventional smelting methods use a separate furnace for each process step. The TBRC process employs but one furnace. The Afton TBRC vessel measures 6.4 m in length and 4.3 m in diameter. The furnace is supported on a set of motor-driven rollers. These permit the vessel to rotate at variable speeds to 40 rpm. Rotation provides good mixing, high reaction rates, and effective heat transfer from the refractory sides to the molten bath resulting in a low thermal gradient. The furnace can tilt through a full 360° for charging, firing, and pouring. Smelting at Afton is accomplished in sequential fashion. Flotation concentrate and limestone flux are initially charged to the hot, empty vessel. Using the water-cooled, oxy-fuel lance, the initial charge is smelted. Once sufficient molten bath is obtained and slagging completed, the metallic concentrate is charged. Materials are charged through the water-cooled, variable position additive chute mounted on the TBRC

Jan 10, 1980

By T. Schrauf, J. Scott

"INTRODUCTION Cyanide heap leaching of gold ores in areas with extremely low winter temperatures requires special considerations to prevent freezing of the leach solution. Factors that impact the solution temperature include ore temperature at the time of stacking and during operation, burial of drip irrigation lines, use of geomembrane covers, and heating of the leach solution. Thermodynamic modeling was performed to simulate expected ore and solution temperatures under different operating strategies for Kaminak Gold’s Coffee Gold project in the Yukon Territory. Mean monthly temperatures at the site ranged from a low of -22.5oC to a high of 13.6oC, while in situ ore temperatures ranged from -0.5 to about -1.5oC. The modeling results were used to optimize the operational parameters and indicated that a combination of drip line burial and heating of the leach solution during the winter months was required to maintain solution temperatures above freezing during the initial years of stacking, although as the size of the heap increased, heating was no longer necessary. MODELING OBJECTIVES The successful operation of a cyanide heap leach under extremely cold winter temperatures requires the implementation of operational strategies to prevent freezing of the leach solution and the ore within the heap. Thermodynamic modeling of the heap provides the means to quantify the effectiveness of these different operational parameters on maintaining the required operating temperatures to avoid freezing within the heap. The operational parameters that were evaluated with the model included: • Burial depth of irrigation lines to prevent freezing during the winter months; • Variation in the temperature of the applied leach solution (to simulate heating of recirculated solution); • Variation in the rate of irrigation including seasonal variations between winter and summer months; • Seasonal versus year-round stacking of ore (to simulate the impact of the initial temperature of each ore layer as it is stacked); • Seasonal versus year-round crushing and stockpiling of ore, including the option to stockpile ore on the leach pad in the winter (i.e. stack but not leach fresh ore in winter); • Start date for seasonal stacking; • Variation of the heap geometry over time (both vertical and lateral expansion) and the associated heat storage within the heap; • Use of thermal covers (gravel or geomembrane) over the ore during the winter months; and • Effect of crush size on heat loss during crushing, stockpiling and stacking."

Jan 1, 2016

By Kenneth S. Johnson

Iodine, a grayish-black nonmetallic element, with a density of 4.9 g/cm3, is a solid at ordinary temperatures. It is a member of the halogen family, along with fluorine, chlorine, bromine, and astatine. Iodine melts at 114°C. and at 184°C it is volatilized to a blue-violet gas that has an irritating odor. It does not occur as an element in nature, but occurs as iodates, iodides, or other combined forms. It is the 47th most abundant element in the earth's crust. Iodine was discovered by Bernard Courtois in 1811. He observed an unknown substance in the crude soda ash that results from the burning of seaweed. Samples of this unknown substance were identified to be a new element, and in 1813 Gay-Lussac named the substance iode, from the Greek word for violet color. The world production during 1989 and 1990 [(Table 1)] is estimated to be about 16 million kg per year (Lyday, 1991), of which about 30% is consumed in the United States. GEOLOGY AND MINERALOGY Compounds of iodine are minor constituents in seawater and brines, in certain marine organisms, and in minerals of the Chilean nitrate deposits. Seawater contains approximately 0.05 ppm iodine, and certain marine organisms, such as seaweed, sponges, fish, and some brown algae, are able to further concentrate iodine (Lyday, 1989a). Some seaweed can extract and accumulate iodine up to 0.45% of their weight, on a dry basis. The northern Chilean nitrate deposits, in the Atacama Desert, contain the following iodine minerals: lautarite, Ca(I03), (calcium iodate); dietzeite, CaJIO,), (CrO,) (calcium iodate-chromate); and bruggenite, Ca(I03), . H20 (Erickson, 1981). Various subsurface brines also contain iodine compounds. Some gas-field brines in the United States and Japan locally contain 30 to 1 300 ppm iodine. Several coals in Germany also contain iodine compounds. Iodine has been recovered from brines mainly in Japan and the United States, but also in Java, Indonesia, Italy, England, and the former USSR. Iodine has also been recovered from seaweed in Ireland, Scotland, France, Japan, Norway, and the USSR. Seaweed was a major source of iodine for the world in the first half of this century, and it remains as a large resource. The reserves and future resources of iodine are large, even excluding the resources in seaweed and seawater, and are shown in [Table 1]. Analysis Iodine as the free element can be detected by the characteristic blue color it gives to a starch solution. Quantitatively it is determined as the free element by titration with standard thiosulfate solutions using starch as an indicator. Colorimetric methods are also applicable. PRINCIPAL PRODUCING COUNTRIES Major iodine-producing nations are Japan, Chile, the USSR, and the United States, with lesser amounts being produced in China and Indonesia ([Fig. 1; Table 1]). Annual world production in 1989 and 1990, respectively, is estimated at 15.6 and 16 million kg. In Japan and Chile, the production of iodine depends on production of other materials, such as natural gas or nitrates, respectively, whereas in US operations (in Oklahoma) iodine is the major product recovered from natural brines. Chile was for a long time the principal world producer of iodine from its nitrate-fertilizer operations, but in recent years Japan has become the world leader with increased production of natural gas and associated iodine-rich brines.

Jan 1, 1994

Introduction One hundred years ago, a recent college graduate created the US aluminum industry. He found an economic way to crack the primeval bond that makes the metal so special. Aluminum doesn't "rust" - oxidize progressively. Rather, when exposed to air, it instantly forms a gem-hard but microscopic oxide coating. This seals off oxygen from the metal underneath. Scratch it and the seal reforms instantly. That tight chemical bond is the reason the metal is not found in nature; it exists only as a compound. The bond had to be broken to free the shiny metal. But once it is free, that self-sealing protection, alongside a remarkable strength-to-weight ratio, offers a marvelous material. Until 100 years ago, finding a way to do that economically was a challenge that defied science. Then it was solved, in a remarkable coincidence, on both sides of the Atlantic. And the way was paved for a material that would become the second most widely used metal in the world. Until that time, aluminum was a rare and precious metal. But then two young scientists independently invented the process that took aluminum outside the laboratory and into applications ranging from pots and pans to spacecraft. Discovery of aluminum Other major metals - copper, iron, lead, gold, and silver - had been used for thousands of years. But aluminum was entirely unknown until 1782, when French chemist Antoine Lavoisier suspected its presence. It was not actually discovered until Sir Humphry Davy, in England, detected it in clay in 1807. It was first isolated as a free metal only in 1825, by Denmark's Hans Christian Oersted. For years, laboratory methods yielded aluminum only in tiny amounts costing about $1200/kg ($545 per lb) - $13,333/kg ($6000 per lb) in today's dollars. Methods developed by French scientist Henri Sainte-Claire Deville in 1854 cut the cost to about $200/kg ($90 per lb). King Frederick VII of Denmark had a helmet of aluminum and gold. And Emperor Napoleon III of France reportedly served his most exalted guests at a state dinner with aluminum plates. He left the lower ranks to dine from pure gold. Deville and others continued to reduce production costs. But the price of aluminum still rivaled silver - $31/kg ($14 per lb) versus about $35/kg ($16 per lb) for silver. Then in 1884, a 2.8-kg (6.25-1b), 230-mm-tall (9-in.-tall) aluminum pyramid was cast as the tip of the Washington Monument. It was probably the first architectural use of aluminum. And it is still there after more than a century. The attractive properties of aluminum were obvious: it is lightweight but strong, conducts heat and electricity, reflects light and other electromagnetic radiation, and resists corrosion. But its commercial promise was blocked by its cost. In 1885, the price of aluminum was $17.70/kg ($8 per lb). It stayed there for the next couple of years until the Hall-Heroult process brought it down sharply. Hall-Heroult process lowers aluminum costs The characteristic that kept aluminum hidden so long continued to thwart chemists. Aluminum clings to oxygen so tightly that it is never found as a free metal in nature, and it is hard to break that bond chemically. The best chemical methods of the 1880s required sodium, and sodium itself was expensive. At Oberlin College in Ohio, chemistry professor Frank Fanning Jewett advised his students that anyone who could make aluminum inexpensive would do humanity a favor and get rich as well.

Jan 3, 1987

By K. N. Han, X. Xie

The dissolution behavior of palladium powder in thiourea solutions was investigated. The variables studied were temperature, stirring speed and the concentrations of thiourea, acid and ferric ion. Palladium was found to be soluble in thiourea solutions in the presence of an oxidant such as ferric ion. The dissolution of palladium in thiourea solutions was believed to be chemical-reaction controlled, with an apparent activation energy of about 25.3 kJ/mol.

Jan 1, 1997

By B. Hornsby

At the Mount Isa mine, Mount Isa, Qld., Australia, sublevel open-stope mining is used for copper ore while both sublevel open-stope and cut-and-fill methods are used for the silver-lead-zinc ores. Production in 1979¬80 was 4 730 000 t (5,200,000 st) copper ore and 2 590 000 t (2,850,000 st) of lead ore. The deepest mine level currently is 960 m below the surface. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size, Shape, and Dip The mine is located in the valley of the Leichhardt River in Lower Proterozoic sediments of the Mount Isa group. The regional strike of the sediments is north¬south and they have a persistent westerly dip of about 1.1 rad (65°). In the mine area, the Mount Isa group comprises a sequence of alternating bands of dolomitic shales and siltstones some 3.2 km (2 miles) in width. It is bounded on the east by greenstones and quartzose sediments of the Eastern Creek volcanics, and on the west by the Mount Isa fault, west of which occur schists and other metamorphic rocks. The ore bodies as currently known at Mount Isa are restricted to the Urquhart shale, a formation of thinly bedded pyritic, dolomitic, and volcanic shales some 1070 m (3500 ft) thick. Towards the top of the Mount Isa group sediments, copper and silver-lead-zinc ore bod¬ies occur as contiguous but discrete entities and are mined and treated separately (see Figs. I and 2). Argentiferous galena, sphalerite, pyrite, and minor associated sulfides occur in distinct concordant bands throughout the Urquhart shales. Wherever these bands are grouped together in sufficient density they constitute silver-lead-zinc ore bodies with widths varying up to 36.6 m (120 ft). Other mineralized bands are common in intervening beds. Individual ore bodies persist for many hundreds of meters concordant with the strike and dip of the enclosing shales. Intersections of these beds with fold axes provide local zones of enrichment but otherwise the mineral content of the ore bodies varies only slightly throughout the extent of the dip and strike. Chalcopyrite is the only important primary copper mineral. It occurs in association with pyrite and pyrrhotite as disseminated flocks and as vein fillings throughout irregularly shaped zones of folded, brecciated, and re¬crystallized Urquhart shale. These zones, which are lo¬cally termed "silica-dolomite" bodies, are broadly transgressive to the shale bedding and confine all significant copper mineralization to the virtual exclusion of silver¬lead-zinc mineralization. In some areas, the downward percolation of acidic water through faults and shears has caused extensive leaching of the coarse-grained carbon¬ates to depths in excess of 610 m (2000 ft), but most of the silica-dolomite remains unleached and the ground is competent. The majority of lead ore being mined is north of the main shaft pillar (R62, U62) area. The major lead¬producing areas for many years have been the Black Star ore bodies (2 and 5 ore bodies) and the Racecourse ore

Jan 1, 1982

By C. J. Dungey

INTRODUCTION The County of Cornwall is situated in the South Western extremity of England, see Fig. 1, and has for hundreds of years been an historical centre of mining activity. In the last century many thousands of Cornish miners emigrated to Canada, U.S.A., Australia and numerous other countries throughout the world to form the nucleii of mining industries in those lands. At present there are four active underground mines within the county, all mining tin, together with a relatively large china clay industry. All the operations are directly or indirectly associated with the three main granite intrusions, namely the Land's End Granite, the Carnmenellis Granite and the Bodmin Moor Granite. The report which follows relates to two mines, owned by the South Crafty Group of Companies, namely the South Crofty Mine itself and its smaller subsidiary, Wheal Pendarves. Both these mines are adjacent to the Carn Brea Granite ridge, see Fig. 2, which is an off-shoot of the main Carnmenellis Granite mass. The Carn Brea Granite has provided one of the most intensive areas of copper and tin mineralisation anywhere in the world. Many of the mines worked during the last two centuries and the early part of this century were situated close to the metamorphic aureole surrounding the granite intrusion. Both South Crofty and Wheal Pendarves have their shafts collared in the overlying slate, locally known as killas, but the major part of their workings are in the granite beneath. The ore bodies, or lodes, of both mines are generally steeply dipping and approximately one to four metres in width. They have been, and continue to be, worked by either back or under-hand stoping with the major level interval being approximately 37 metres. A feature of both stoping systems is that they require no back filling and in many cases the completed working areas have remained open for many years due to the competent wall rock conditions. The stoping operation necessitates having box-holes or draw points at regular intervals along the strike length at 4 to 7 metres centres; thus a stope some 50 metres long may typically have ten box-holes or draw points through which air can percolate into the empty stopes above. The granite country rock contains traces of uranium minerals, namely pitchblende or uraninite, and it is the decay of this uranium mineral which leads to the formation of the mobile radon gas which, together with its short lived daughters, may create a hazard. The presence of uranium minerals has long been known, but the quantities being so minute, it was not appreciated until some ten to fifteen years ago that these might present a hazard. The Mines and Quarries Inspectorate approached Management approximately ten years ago with a view to taking samples in areas of the mine that they felt might contain the highest concentrations of radon. Samples were thus taken in several areas of both mines, normally well away from working areas and where there was no through ventilation. The results of these initial tests were published by the National Radiological Protection Board in a paper entitled "Radon and Its Daughters in Various British Mines". The results of the initial surveys made it apparent that improved ventilation was the most important factor in reducing radon daughter concentrations within the mine atmosphere, with increased flow rates having three beneficial functions, viz. (1) To dilute radon daughter concentrations. (2) To prevent radon and its daughters reaching a state of equilibrium. It is not uncommon for radon and its daughters to have a disequilibrium factor of 1,000 in some uranium mines.

Jan 1, 1981

By Deepak Malhotra, Linda Battich

INTRODUCTION Mineral processing plants generally operate efficiently within the technical, economic and management constraints applied to them. They have no choice, because the dynamic and competitive nature of the mineral business dictates that survival and/or prosperity of the company depends on being a low cost producer. Progressive mineral companies realize that market downturns will occur, and they must continuously strive for lower unit production costs rather than be satisfied with the status quo. Mineral companies generally tend to reduce costs during market downturns. This short-term solution is achieved by high grading of ores, and cut-backs In waste stripping, exploration, and capital expenditures.(1) However, to remain a low-cost producer in the long term, companies must enhance revenue and reduce cost simultaneously. This Involves continuous evaluation of technologies and reagents, periodic audits to locate revenue losses; and gathering and analyzing statistical plant data to clearly understand the Impact of process variables on efficiency. Mineral companies are caught in a Catch-22. They need resources to accomplish the objectives outlined above. Most mill stuffs have been cut to the minimum level necessary to maintain production. Resources such as time, manpower and money are scarce, and mill metallurglsts struggle to use these resources productively. Is your flotation plant optimized? History Indicates that successful operations constantly tune the flotation circuits to Improve metallurgy or reduce cost. This Is the essence of evolutionary change In mineral processing operations. A plant auditing approach can play a key role in assisting the process engineers to achieve the overall objective of enhancing metallurgy and reducing cost with limited resources through evolutionary improvements. FLOTATION SYSTEM The flotation process is the most widely wed concentration process In the world today. The number of variables that can affect the flotation results Is very large, maybe over 100. These variables can be divided Into three main groups: (2) (1) ore and mineral properties; (2) reagent treatment; and (3) flotation machine characteristics. Some of these variables can be controlled while others have to be accepted. Recently, Investigators (3,4) have attempted to describe the flotation process as an engineering system which includes chemistry, equipment and operating variables. Though some of the variables are Independent, most are co-dependent. thus making it difficult to develop simple cause and effect relationships. For example, it is easy to understand the effect of reagent change on recovery and grade In the rougher circuit, but very difficult to quantify Its impact on filtration downstream. Plant audits on a periodic basis are necessary to develop an understanding of the Interactive nature of the system. Auditing provides comparisons of plant operating performance before and after variable changes. OVERVIEW OF PLANT AUDITING An audit Is a formal, thorough and periodic examination of a system. There are various types of audits (i.e., geological, environmental, metallurgical. economic, etc.), and each serves a different purpose. (5) The discussion here is related to plant and/or metallurgical audits. Neale and Flintoff note that two key factors governing changes In an operation are: (1) the opportunity must be recognized; and (2) the justification for change must exist. (6) Both factors emphasize the critical role of successful plant auditing. Without plant audits. many opportunities for improvement must go unrecognized. On the other hand, when a situation ripe for improvement Is recognized. the hard data from a plant audit can reveal the expected impact of an operating change.

Jan 1, 1993

By Thomas O’Neil

Part 2 of this article continues with a discussion of features that distinguish contemporary investment analysis in mining from earlier methods. Part 1 traced the evolution of evaluation methodology from the traditional and discounted cash flow methods to the present. Quantitative investment models of the past 20 years played an important role in mine evaluation, but a new integrated approach to mine investment decision-making is now developing because of growing economic, technical, and political uncertainties. One of those uncertainties -inflation-was discussed in Part 1. Increased Risk At a time when mineral production technology and analytical evaluation tools have never been more sophisticated, it may seem anomalous to claim that mining ventures are riskier. Whereas technical risk has, indeed, been notably reduced, business risk the chance of a major loss due to market factors-has increased. This has been caused by two factors: the rapidly rising capital requirements of new projects and increasingly unpredictable future economic conditions. These difficulties are compounded by the long lead times of large, new mining projects. Rising Capital Requirements A $100 million investment is no longer considered to be a large mining project in most cases. For a large metal mining project, the total cost of the concentrator alone can exceed $100 million. Amoco Minerals' Thompson Creek molybdenum mine being developed in Idaho will cost over $350 million, for a relatively modest sized (23-kt/d or 25,000-stpd) ore operation with no extraordinary features. For a major grass roots copper operation producing blis¬ter or refined copper, the capital requirements are billions of dollars, not millions. Project costs are higher for two reasons. One, the projects themselves are larger; and two, the capital cost per unit of production is rising rapidly. Projects are larger because lower grade ores, yielding smaller margins, must be mined in greater volume-particularly in view of the high fixed costs of modern technology. When analysts contemplate any capital project, an important consideration is the minimum economic size of the facility. This is the smallest size that can be built and still be competitive on a total unit production cost basis. In the conventional copper smelting business, for example, a production rate of at least 91 kt/a (100,000 stpy) of blister copper is about the absolute minimum for reducing unit fixed costs to competitive levels. Clearly, many variables affect the minimum economic size of a mining facility, but with few exceptions that size continues to grow. Unit capital costs of new mining projects also are continuing to rise. The accompanying table shows that the capital costs in primary copper production have risen faster than the cost of living. Of particular note is the capital charge of over $1.70/kg (77411b) required for a new, fully integrated facility in 1981. Thus, the mid-1982 copper price would not even cover the cost of capital for a new copper producing complex without considering the direct operating costs for such a facility. The breath-taking costs of new mining projects are caused by factors such as increasingly complex technology, the rising social costs of creating attractive living and working conditions for employ¬ees, and the expenditures re¬quired for environmental protec¬tion. For whatever reasons, how¬ever, unit capital costs in mining continue to exhibit staggering growth rates. Future Economic Conditions Risk is not only a function of the size of the potential loss-an im¬portant measure of which is the capital intensity of mining dis¬cussed above-but also of the like¬lihood of incurring the loss. It is here that the ever increasing lead times of mining projects and the glaring limitations of economic forecasting are so important. In comparison to nearly all other in¬vestments, mining has more capi¬tal at stake for a longer period of time before commercial uncer¬tainty is resolved by distant eco¬nomic conditions that are subject to high forecasting errors. A sobering example of the pre¬ceding points occurred with the Lakeshore copper mine in south¬ern Arizona that Hecla Mining Co. undertook to finance and manage in 1969. The project was a joint venture of Hecla and the El Paso Natural Gas Co. and included two mines-one producing oxide ore and a second in the sulfide zone¬and a complex processing facility producing both electrowon cop¬per cathodes and copper precipi¬tates. The project originally was

Jan 12, 1982

By Clarice L. Amaral, Ricardo Melamed

Clay-coating has been used as an environmental technology tool for the immobilization of toxic elements. The mechanisms of arsenate retention in clays suggest that not only the system pH is of fundamental importance in the process but also the character of metals and ligands that compose the electrolyte system. The goal of this study was to investigate the interactions between system pH and electrolyte ions on the retention of arsenate in an Oxisol soil. The sorption of arsenate decreases with the increase in system pH, reflecting the competitive sorption of arsenate with OH groups. Solutions containing Ca tend to favour the immobilization of arsenate in a relatively broad pH range.

Jan 1, 1998

By J. C. Yingling, J. W. Leonard

Introduction The absence of coal-face particle management heavily penalizes the transportation of coal from initial loading to final consumption. The penalties include dust problems, significantly reduced mine-loading-cycle productivity, mine-belt spillage, excessively high coal-preparation costs, chute blockages and dangerous pulverizer blockages at the final point of utilization. Fine particles commonly cause environmental and economic problems. It is well known that these fines can cause safety and environmental dust problems. But it is not well understood that these fines can also swell broken coal to a point where 5% to 15% more time and capacity must be used to deliver the same tonnage. In this paper, methods and rewards for reducing and/or managing fines at the mine face are discussed. Computer-based loading-cycle model productivity estimates, viewed from a new perspective, are made on the basis of material volume rather than on the long-established, and frequently misleading, basis of tonnage. It is typically the volume of broken material being transported that defines the capacity of a given transportation system, while the corresponding tonnages are merely a reflection of the specific material densities. Published evidence suggests that the swelling of broken coal can be decreased very significantly using small quantities of certain nonfrothing chemicals, which are added to mine-face spray water, and by employing improved mine-face breakage practices. In a future paper, the effects on transportation productivity beyond the coal mine will be discussed. The precursor to the work presented in this paper, involving the bulk density improvement for broken coal and the subsequent production gains for underground coal mines, was earlier presented in Leonard and Newman (1989). In the past, this topic has been studied and practiced only in byproduct coking in the steel industry. However, a potential exists for an increase in coal-industry productivity by improving the bulk density of coal to yield a subsequent reduction in delivered cost. This can occur with breakage, handling and treatment methods resulting in the loading of greater quantities of coal in fixed volumetric capacity haulage units such as mine cars, shuttle cars and scoops. Laboratory-based experiments to achieve an increase in productivity by increasing coal bulk density were discussed in Leonard, Paradkar and Groppo (1992). Chemical techniques using small quantities of commercially available reagents (surfactants) resulted in about a 13% to 15 % increase in bulk density, which was thought to produce a proportional increase in the productivity of a mine, together with a subsequent reduction in cost. The idea is to mix the reagents with the water that is used to spray coal during mining. In this paper, the impact of bulk density improvements on production rates is presented. Increases in production ranging from 60% to 88% of the bulk density increases are projected. This analysis was performed for atypical continuous-miner section. In the following sections, discussion and results of the analysis are presented. Discussion An analysis was performed to ascertain the impact of bulk density improvements on face-production rates for a typical continuous-miner section. Figure 1 illustrates the section layout and cut sequence. This layout and sequence is identical to the case described in King and Suboleski (1991). As can be seen, the section uses five entries and 12.2-m cuts that are taken by a remotely controlled continuous miner. The seam height is 1.5 m and two shuttle cars (5.7 t nominal capacity) are employed for haulage from the miner to the section feeder, which, throughout the cut sequence, is positioned as illustrated in Fig. 1. The simulation model was coded in the SIMAN simulation language. The major impacts of increased bulk density improvements on such a production system are as follows: •Shuttle-car payloads, in terms of the mass of coal transported per haul cycle, are increased proportionally to the increase in bulk density that results from the application of surfactant. •Shuttle-car discharge times should remain largely unchanged, because they are determined by the volume of material that is discharged, rather than the mass, and this volume does not change.

Jan 1, 1996

By T. M. Plouf, G. A. Lang

INTRODUCTION The finding or acquisition of ore bodies containing 200,000 opt gold and greater is becoming more difficult. There are more investors and small heap leach operators who have or can raise the capital to make the investments in the large ore bodies. Echo Bay Mines' approach has been to regionally develop small ore bodies, two or three in the same geographic location, which when considered together contain over 200,000 opt gold. CONCEPT The "world class" ore body syndrome, which many large gold producing companies maintain, can restrict company growth if the finding or acquisition of these large ore bodies is not forthcoming. Another approach has been taken by Echo Bay Mines. The concept is to regionally develop two or three small ore bodies suitable for processing with portable equipment. The systematic scheduling of mining, processing with a single flexible processing plant using the same management/ operator team can reduce your overall investment. The initial selection of mining and crushing equipment suitable for sun11 open pit operations is essential. A very viable option in the mining sequence is to use contract mining/crushing/screening/agglomerating and stacking. This option also lowers your initial capital investment and increases the flexibility in mining as you move from one mining operation to another. LOCATION AND SEQUENCING One single small ore body of itself is not attractive enough for the large producer. The recovery of the initial capital investment, although very rapid, does not make a viable project. The only alternative is to find two or more small ore bodies close enough geographically to each other so that the same management/operator team can be used to run the operations as each is put into operation. Echo Bay Mines has regionally located several ore bodies ranging in recoverable ounces up to 100,000 oz. These small separate ore bodies, located within a 30-mile radius of each other, for planning purposes have been considered as one single ore body. The first small ore body has the initial burden to recover the capital invested in the office, laboratory, power plant and processing plant. Other capital investments such as water supply, access and haul road, fencing, septic system, pads and ponds are reoccurring at each mine site. The first small ore body must recover the initial capital investment in the movable structures as well as the reoccurring capital investment in the fixed structures very rapidly. A total capital recovery time of 6 to 9 months should be used as an economic decision point for both choosing which small ore body should be processed first, and what the plant design capability should be. The sequencing of the small mines is very critical. All the fixed structures mentioned above must be installed at the second mine site before the first mine has been mined out. The ideal sequencing will allow for only one month between shutdown of the first small mine and production from the second. SPECIFIC OPERATION Echo Bay Mines has completed the construction of the Illipah Mine located in White Pine County, Nevada, with this small mine sequence approach. The Illipah Mine is a 40,000 ounce recoverable gold ore body. This ore body met all the necessary requirements. The Illipah project could support the capital investment in the movable structures which will

Jan 1, 1987

By Dennis Z. Mraz, Maurice B. Dusseault

INTRODUCTION In 1982 and 1983, as part of the mine design program for a New Brunswick potash mine, series of laboratory tests were conducted on potash and salt samples obtained during the mine exploration program. Concurrently, two rock mechanics monitoring programs were initiated in-situ the short term test at site A, and the long term test at site B (figures 1 and 2). SHORT TERM TEST (FIGURE 1) The purpose of this test was to obtain a quick confirmation of laboratory data and comparison of the rock response to excavation with the data available from Saskatchewan potash mines. Instrumentation in the main drift was installed on Day 4 after excavation, and the pressure cells were pressurized on Day 33. The packer type cells in the east crosscut were pressurized on Day 25. Excavation of the west cross-cut was begun on Day 71. This "mine by" test has provided important information required for calibration of laboratory tests, as will be shown later. The test data recorded are shown in a series of diagrams: a. initial room closures, main drift-figure 3; b. convergence graphs, main drift-figure 4; c. stresses recorded, main drift-figures 5 to 8 and, d. stresses recorded, east crosscut - figure15 Three types of instruments were used in this test. In the main drift. Saskatchewan Research Council water type hydraulic inclusions. encapsulated in potash, were used to measure lateral and vertical stresses. In the east crosscut a home-made packer type hydraulic inclusion was used to record radial stresses. The convergence stations consisted of end-grouted bars of reinforcing steel. LONG TERM TEST (FIGURE 2) The purpose of the long term test was to study the steady-state behaviour of a single opening in potash. Site B was placed in a remote area of the mine, well away from any anticipated early mining activities. To assure homogeneity of rock mass, the site was placed within a 30 meter thick vertical limb of a syncline-shaped ore body contained within a 360 metre thick vertical section of salt rocks. It was excavated mechanically as close as possible to the designed square cross-section for the entire length of the room. The square cross- section was chosen in order to achieve a comparable geometry in vertical and horizontal directions; that is, to eliminate any undesired effects of geometry interaction with lithostratigraphy which might have generated non-homogeneous behaviour. The site was extensively instrumented. As in site A, the convergence stations consisted of end-grouted bars of reinforcing steel. However, in order to avoid any measurement inaccuracies which would be associated with the brittle yield zone in the opening walls, bars were placed to the same depth in opposite sides of the opening. This allowed the convergence measurement to be obtained as the absolute convergence between two points, rather than the convergence of a single point with respect to a potentially unstable reference. For stress measurements the Saskatchewan Research Council cells were placed vertically, laterally and radially. Unfortunately, due to cost considerations, it was decided to use a compacted sand, rather than grout, to encapsulate the cells in the boreholes. Due to the extremely small strain rates, only radial ly placed cells could be adequately pressurized. The rest of the cells did not perform adequately, and are currently being replaced.

Jan 1, 1986

By Prasad S. Mulukutla

INTRODUCTION The use of specialty chemicals In mining Is an Important consideration In determining the economic efficiency of a floatation plant. The plant metallurgist would like to maximize the recovery of valuable minerals while maintaining an acceptable grade product. The direct cost of the reagents, while significant, Is only a part of their economic effect. Reagents determine the flotation rates and the selectivity of the mineral separation. In flotation practice, mineral selectivity can be achieved through the addition of various reagents that can control the wettablllty of a particular mineral In the flotation pulp. The reagents which will generally optimize the flotation process are commonly classified Into three distinctive categories: collectors frothers and modifiers. Large tonnages of metal sulphide concentrates which are the source of most of the common base metals are produced economically by fairly standard flotation techniques. The advent of modern technology coupled with the necessity of processing finely disseminated and complex ores of low grade has Increased the need for utilization of non traditional and high selective chemicals reagents for the economic recovery of metals and raw materials. Mining operations in developing countries continue to progress at a rapid growth as many countries mine and process both metallic and nonmetallic ores. The total mine production of base metals, precious metals, Iron ores and coal In developing countries In Africa, Asia and South America has been much higher as compared to the western world production In terms of tonnage and dollar value. Global capital Investments In mining during 1991-95 Is $ 55.1 billion and It Is anticipated that the development of new and specialty reagents may be concentrated In the processing of aluminum , copper, precious metals and Industrial minerals with special reference to potash, phosphates, kaolin and silica sands.(1) The challenge for the design and development of new specialty chemicals Is dependent on Its cost performance and environmental acceptability. Information on the theory and application on the development of mining chemicals has been reviewed In several conference proceedlngs,(2-4) publications(5) and also discussed by several Individual authors.(1,6-10) These publications serve as a comprehensive reference source on the reagents used In several flotation processes. This paper discusses the present trend in the use of specialty mining chemicals In diversified mining operations on several mineral systems with particular reference to sulfide ore systems. The necessity to expose these specialty chemicals through aggressive marketing techniques In mining operations In many developing countries Is also presented. SPECIALTY CHEMICALS Specialty mining chemicals, like other specialty chemicals, share common business characteristics that transcend geographic boundaries. They are manufactured 1n a few selected areas, generally expensive relative to commodity chemicals, shipped and compete globally and usually require specialized knowledge for proper use. Specialty mining chemicals are relative newcomers to chemical markets. The vast majority of mining chemicals are derived from products that are less than two decades old. Traditional collectors and frothers have been available for a considerably longer time. The emphasis on adding value to existing products and byproducts Is responsible for several characteristic features of the mining chemicals business. Most raw materials for these chemicals are related to oil and gas byproducts, chemical refineries, pulp and paper mills and all chemical manufacturers. Until the last several years these operations were located In Industrialized countries, while a significant number of mining chemical users were In developing countries. Therefore International trade In mining chemicals developed through agencies or agents in each particular country.

Jan 1, 1993

By P. Somasundaran, F. F. Aplan

The Formative Years Douglas Winston Fuerstenau was born December 6, 1928 in Hazel, SD to Edwin Arnold and Hazel Pauline (Karterud) Fuerstenau. His father, of German extraction, was a farmer and, later, a cabinet maker, while his mother, of Norwegian lineage, was a school teacher. His parents are still alive today at ages 91 and 88, respectively. Two younger brothers, Richard and Maurice and a sister, Jean, completed the family. The family first lived near Watertown and later in Lemmon and Mobridge, SD. The 1930's were a difficult period for any family in South Dakota. A decade-long drought made this a part of the dust bowl, which together with a serious infestation of grasshoppers (the locust plague described in The Bible), and the world-wide depression, made living difficult for all. In spite of these privations, Doug has many happy memories of his childhood which can be attributed to the resilience of youth and a strong family structure. He has fond memories of summers spent with his cousins on family farms where the children were introduced to farm chores at an early age. When he was old enough, he worked in a grocery store after school. At the age of 14 he spent the summer hearding 2,000 sheep near Lemmon, SD as his contribution to the war effort in that labor-short period. He remembers well the dust storms of the 1930ts, the destruction at his grandmother's farm by tornado, pulling weeds on the farm, threshing grain with a monstrous steam-driven machine and such pleasant childhood activi- ties as fishing for bullheads, hunting for rabbits, searching for arrowheads in newly plowed fields, and ice skating in the winter. Early in life his parents instilled in him a strong work ethic, a love of learning and the need for a good education. In 1943, the family moved to Rapid City, SD. With a population of 13,000, it was

Jan 1, 1989

By Jon P. Thorson, James W. Babcock

The second Belt Symposium in Missoula, MT, Oct. 9-16, 1983, was designed by its conveners as a field conference on progress in the Proterozoic Belt Supergroup rocks. The symposium was sponsored by the University of Montana with corporate support by Anaconda Minerals. Asarco Inc., Exxon Minerals Co.. Hecla Mining Co., Molycorp Inc., Noranda Exploration Co., and Utah International Inc. Introduction The Belt Supergroup is a thick middle Proterozoic sedimentary sequence deposited in a basin along the edge of the North American craton. Subsequent geological events have modified the basin so its remaining rocks outcrop in Montana, Idaho, northeast Washington, southeast British Columbia, and southwest Alberta. The Canadian equivalent of the Belt is called the Purcell Supergroup. Several similar age basins occurred along the western edge of the Proterozoic North American craton. The sedimentary sequence in the Belt basin aggregates is more than 18 km (60,000 ft) in the thickest section, where the bottom is still unexposed and the top removed by erosion. The limits of the Belt basin are poorly known because of later sedimentation and tectonic disruption. Knowledge about the Belt terrane has been difficult and tedious to obtain because the rocks are generally monotonous and fine-grained. During the last decade, new 1:250,000 scale reconnaissance maps of most of the Proterozoic Belt terrane were published, open-filed or otherwise brought close to completion. These maps provided the opportunity and challenge to consider the stratigraphy and tectonics of the Belt Supergroup as a complete basin. This broader view has contributed greatly to the solution of numerous geological problems created by previous, more focused work. Along with the availability of better geological base, there have been new exploration developments. The 1973-1983 decade saw proof of the commercial viability of the Cu-Ag mineralization at the Troy mine. The conclusion was also accepted that the Sullivan mine was a sedimentary massive sulfide. So was the conclusion that the Cu-Co mineralization at the Blackbird mine at Cobalt, ID, was the result of similar submarine exhalative processes. And there were discoveries of another Cu-Ag deposit "larger and higher grade" than the Troy mine and of two new massive sulfide districts. The Belt-Purcell terrane has long intrigued exploration and mining geologists as the host of Pb-Zn-Ag veins of Idaho's Coeur d'Alene district and of the Sullivan mine at Kimberley, British Columbia, one of the largest Pb-Zn-Ag deposits in North America. In the mid-1960s, Bear Creek Mining Co.'s discoveries of the stratabound copper-silver occurrences in the Revett Formation brought new interest in the basin's geology. The recognition of the Sullivan ore body as a sediment-hosted massive sulfide has also generated minerals exploration activity in the US portion of the Belt basin. This activity increased again after the Australian discovery of the world class Olympic Dam deposit at Roxby Downs, South Australia. This occurred because of some suggested similarities in the regional geologic setting of the Belt basin with the Adelaide geosyncline, the host for the Olympic Dam deposit. Technical Sessions Technical sessions topics included Lower Belt and Ravalli Group stratigraphy, Middle Belt carbonate and Missoula Group stratigraphy, Belt-Purcell stratigraphy, structure and tectonics, strata-related mineral resources, geochronology and geophysics, geochemistry and isotope geology, and a summary of significant problems. Evening poster sessions dealt with the topics and papers presented each day. Since the first Belt Symposium, mapping of the stratigraphic units and their tectonic setting has enabled geologists to progress in conducting a basin analysis. Belt Symposium II revealed no concensus on many aspects of belt geology. However, basic stratigraphic studies have reached the point that models can be constructed. The Prichard Formation is equivalent to the Aldridge Formation, but detailed correlations were not previously available. Studies were made by Finch and Baldwin (Cominco American) and by Cressman (US Geological Survey). These studies established important correlations of key markers within the Aldridge Formation to units in the Prichard Formation. This will help eliminate geologic interpretation problems at the US-Canada frontier and make Belt basin reconstruction more accurate. Also, the Sullivan ore deposit was shown to be formed in the deeper distal part of the basin. This contrasts to the more proximal depositional position of the Prichard Formation. Hamilton

Jan 6, 1984

By R. M. Bricka, T. J. Olin

Heavy metals contamination is an environmental problem at Army installations engaged in firearms training and munitions production. At these facilities, weathering and corrosion of expended munitions and leaching from wastewater lagoons, landfills and burn pits have resulted in heavy metals contamination of the soil. The principal metals encountered in firing-range soils are lead, copper and zinc. Cadmium and other metals, such as antimony, that are often incorporated in the munitions are sometimes seen in lesser concentrations. Mercury is associated with various propellants and, while present in much smaller concentrations, is of concern because of its acute toxicity. Chromium is primarily associated with plating operations. The transport of metals into groundwater has been confirmed at some locations, which has required treatment of the soil and groundwater at these sites. Certain treatment processes for contaminated soil produce metals-laden extracts, which also require treatment before reuse or disposal. Ion exchange is generally quite effective for removing metals from aqueous streams. However, resins are expensive and must be regenerated, and activated carbon is generally less effective for most metals and also requires regeneration. Therefore, alternative effective and economic sorbents are needed. Twelve sorbents were screened in initial batch testing. These included activated carbon, bark, chitosan, crown ether, corn cob, xanthate, clay (kaolinite and montmorillonite), peat moss, seaweed and reagent-grade zeolite (aluminosilicate, Sigma Product No. Z3125). Of these, zeolite demonstrated the highest capacity for Pb, Cr and Cd. For this reason, zeolite was selected for further testing in batch, kinetic and column studies. Materials and methods Zeolite. The zeolite used in the second-phase batch and column studies was obtained from a natural deposit of clinoptilolite-rich rock located in South Dakota (Rocky Ford SDH) (Desborough, 1996). Large blocks of the material were crushed and sieved into the following three particle size ranges: 0.5 to 1.0, 1.0 to 4.0 and 2.0 to 4.7 mm. This material demonstrates high structural stability in acidic solutions (pH 2.5) (Desborough 1996) and has a measured surface area of 30 m2/g. The measured total cation-exchange capacity (TCEC) was approximately 10 meq/100 g. This is well below what has been indicated for commercially available zeolite, which has been reported to be about 180 to 220 meq/ 100 g. The TCEC test was repeated (Method 9081, SW 846) using sodium acetate. The test resulted in a TCEC of 54.5 meq/100 g. The difference between values obtained for this material and published values for zeolites may be attributable to the greater heterogeneity in the material used in this study, compared to commercially available materials, or to the effect of the relatively large particle sizes utilized. Batch studies. Seven batch studies were conducted using synthetic metal solutions and soil extracts (Table 1). Extract composition: The P-extract was prepared by sequential surfactant extraction of organics from a burnpit soil followed by acid extraction of metals. The pH of this solution is approximately 1.1. A number of metals and organic compounds were present in the soil. Analysis of the extract was restricted to Pb, Zn and Cu concentrations for this study. The FBH extract was produced from a firing-range soil that was oxidized with a 0.01 M CaO solution and then extracted with 0.1 M acetic acid. This was filtered through 0.5-µm Whatman No. 5 filter paper and stored at room temperature. The pH of the FBH extract was approximately 4.5. pH Control: Calcium carbonate (CaCO3) may be present in the zeolite horizon or bed. Calcium ion (Ca") is released from the exchange sites when in contact with solutions containing ions for which it is more selective, such as lead. This results in a rise in solution pH over time. Acid washing removes most of the carbonates, eliminating the need for a buffer. Batch studies were conducted using both acid-washed zeolite (AW) and unwashed zeolite (UW) for performance comparison. The UW zeolite was rinsed with distilled deionized (DDI) water to remove fine soil particles. Both materials were dried at 105°C (220°F) overnight, so that the dry mass could be determined. Column studies. Ten column studies were conducted. Because it was expected to have the best hydraulic properties, the largest particle

Jan 1, 1999

By J. M. Hans, M. F. O’Connell, G. E. Eadie

INTRODUCTION The U.S. Environmental Protection Agency (EPA) was required, under Section 114(c) of Public Law 95-604, to provide a report to Congress identifying the location, and potential health, safety and environmental hazards of uranium mine wastes together with recommendations, if any, for a program to eliminate the hazards. The approach taken to prepare the report was to develop model active and inactive mines and locate them in a typical mining area to estimate their environmental impact. A list of uranium mines was acquired from the U.S. Department of Energy (DOE). The inactive mines were separated from the list and sorted into surface and underground categories. A literature search was conducted to obtain and consolidate available information concerning the environmental aspects of uranium mining and shortterm field surveys and studies were conducted to augment this information base. Radioactivity emission rates were measured or estimated for each mining category and were entered into computor codes to assess population exposures and subsequent health risks. The general environmental condition of inactive uranium mines was determined by walk-through surveys in several mining areas. INACTIVE SURFACE MINES We assumed that a model inactive surface mine contains a single pit with the wastes (overburden and sub-ore) stacked into a pile adjacent to the pit area. No credit for reclamation is given to the model mine. In lieu of the availability of individual mine production statistics, the model surface mine size was established from the total ore and waste production statistics for all surface mines, divided by the number of inactive surface mines. The number of inactive mines, obtained from the DOE mine listing, are summarized by type and location (Table 1). For modeling, we assumed that there are 1,250 inactive surface mines. The total or cumulative waste and ore production for inactive surface mines from 1950 to 1978 is not fully documented. Uranium mine waste and ore production statistics, on an annual basis, were available for both surface and underground producers from 1959 to 1976 (D0159-76). Annual uranium ore production for each uranium mining type are available for 1948 to 1959 (DOE79) and for combined ore production TABLE 1. Consolidated list of inactive uranium producers by State and type of mining [State Surface Underground AL 0 9 AZ 135 189 CA 13 10 CO 263 902 ID 2 4 MT 9 9 NV 9 12 NJ 0 1 NM 34 142 ND 13 0 OK 3 0 OR 2 1 SD 111 30 TX 38 0 UT 378 698 WA 13 0 WY 223 32 Total 1246 201T] for underground and surface mining from 1932 to 1942 (DO132-42). In order to estimate waste accumulated prior to 1959, the waste-to-ore ratios from the 1959 to 1976 period were plotted vs. time and line-fitted by regression analysis (Figure 1). Unfortunately, the extrapolation of the line to years prior 1959 approached zero in 1954 although surface mining began in 1950. Therefore, a waste-to-ore ratio of 8:1 was used for the period of 1950 to 1959 based on ratios estimated by Clark (C174). The waste to-ore ratios for 1976 to 1978 were estimated using the line established in Figure 1. By using waste-to-ore ratios and ore production data, the cumulative waste and ore production for both surface and underground uranium mining is estimated to 1978 (Table 2). The estimated cumulative waste from uranium surface mining for 1950 to 1978 is 1.73 x 109 MT. A crude estimate of the waste accumulated at the model inactive surface mine can be made by dividing the total waste produced to 1978 by the number of inactive mines. This, however, overestimates the waste tonnage because some of the contemporary wastes are being produced by active mines, and the waste accumulated at newer mines has increased in recent years. To adjust for this overestimate, we assumed that all mines operating in 1970 will be inactive by 1978. This eight year period is approximately one-half the lifetime of a model

Jan 1, 1981