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By Robert G. Beverly

INTRODUCTION Standards limiting the annual exposure of United States uranium miners to radon daughters were established in 1967 at 12 Working-Level-Months (WLM). The standard was reduced by a factor of three, to 4 WLM, in 1971. Currently, the standard is again being examined to determine if it should be changed. Since 1967, Union Carbide has calculated individual monthly exposures in company and contract-operated mines located on the Colorado Plateau. Although it has been possible, by extensive ventilation control measures and accurate routine sampling, to meet the current exposure standard, there are many miners whose exposures closely approach the 4 WLM standard for any given year. However, it was noted that for miners who work for any extended period of years the [average] exposure was much less than the standard. The primary purpose of this paper is to show that, in effect, any annual exposure standard to radon daughters results in a long-term exposure considerably below that standard. Further, most miners, due to their job assignments and/or employment habits, only receive a small fraction of the standard. HISTORY OF EXPOSURE STANDARDS Prior to 1967, radiation protection in uranium mines was fundamentally based on a radon daughter concentration guide. In 1960, the American Standards Association published mine and mill radiation protection standards (ASA-1960). The Colorado Department of Mines, in 1961, adoped a standard which followed the ASA Standard and provided that if concentrations exceeded 10 Working Levels (WL), the area was to be shut down until corrective action was taken; if between 3 and 10 WL, corrective action was to be initiated; between 1 and 3, additional samples were to be taken and individual exposures evaluated; and if below 1 WL, conditions were considered to be controlled. In 1967, the U.S. Department of Labor issued the first exposure standard which called originally for limiting annual exposures to 3.6 WLM but which was later changed to 12 WLM. The complicated regulatory developments leading to this standard have been described elsewhere (Beverly-1969, Rock & Walker-1970). Effective July 1, 1971, this exposure standard was lowered to 4 WLM per year, which is the current standard. Over the past year, there has been speculation about the potential risk to uranium miners working at the present standard. A recent NIOSH Study Group Report (NIOSH-1980) concluded: "There is also strong evidence that a substantial risk extends to and below 120 WLM of exposure." The 120 WLM corresponds to a miner working in uranium mines for 30 years, a rare occurrence, at an exposure rate of 4 WLM per year, an even rarer occurrence. On the other hand, the General Accounting Office, in a recent Report to the Congress (GAO-1981), was very critical of reports by NIOSH on general low-level radiation risks. The GAO recognized that”...important questions remain unanswered about the cancer risks of low-level ionizing radiation exposure;" and recommended that Congress enact legislation giving statutory authority to an interagency committee to coordinate Federal research on health effects of ionizing radiation exposure. The International Commission on Radiation Protection at its March, 1980 meeting recommended limiting the inhalation of radon daughters to 0.02 J per year, equivalent to 0.4 WL, which on an annual basis would be 4.8 WLM and noted it is common to reduce this figure by 20% for allowance in the case of uranium miners for external and/or dust exposure(Sowby-1980). This is essentially equal to the present standard of 4 WLM. As earlier uranium miner exposure studies are reevaluated, and as new studies are conducted, it is important that the relationship between regulatory standards and the resulting actual exposures be recognized. UNION CARBIDE URANIUM MINING EXPERIENCE Union Carbide started mining Colorado Plateau uranium-vanadium ores in the late 1920s for the contained vanadium values. In the early 1950s, the Atomic Energy Commission contracted Union Carbide to produce uranium at mills located in Uravan and Rifle, Colorado. The company now has over fifty years of mining experience in the area. Some mines are operated as company mines and others are operated by private mining companies under a contractual arrangement. Ventilation, sampling, and exposure calculations are carried out the same in contract mines as in company-operated mines. Data presented in this report do not differentiate between company or contract employees and include all employees who worked underground any portion of a year in Union Carbide mines from 1967 through 1980. At the peak of uranium mining activities in 1970, there were 577 miners employed at year end (285 company employees and 292 contract) and 52 mines in operation (8 company-operated and 44 contract mines). Contract mines varied from two-man operations up to 15 employees. Company mines were generally the larger operations and employed from 20 to 100 miners.

Jan 1, 1981

By R. Svensson

Mill linings play an important role in all comminution processes with tumbling mills. The selection of the lining profile and the lining material has a great impact on over-all grinding cost and mill throughput. The cost of the lining itself is minor in comparison with media cost, energy cost and the value of the material produced in the mill. The total scope must be considered when mill linings are selected. This paper presents case stories and explains the technology behind: -different lining profiles in rubber and steel -OREBED magnetic lining -Steel-cap rubber lining

Jan 1, 1994

By Ayman A. El‑Midany, Hassan E. El‑Shall, Kerri‑Ann A. Hue

Modification of clays is necessary for the zearalenone (ZEN), as long-chain mycotoxins, removal. The organic-cations were usually used to modify the clay hydrophobicity to accommodate the large mycotoxin molecules. In this study, tri capryl methylammonium (TCMA) chloride and common corn oil were used in clay modification. Several characterization techniques such as XRD, FTIR, and TGA were used to identify the d-spacing changes after modification either by TCMA or corn oil. The results of applying the TCMA-corn oil-modified clays in ZEN removal indicated a noticeable improvement where more than 90% of ZEN could be removed in vitro at acidic and neutral pH values. The success of eliminating a high percentage of ZEN molecules is due to the expansion of intergallery spacing as a result of TCMA surfactant intercalation within the clayforming layers. The removal percentage does not depend only on pH but also on the amount of clay used. In addition, the ZEN removal in the presence of corn oil is better at acidic pH and small clay amount where the ZEN can be fully removed.

Sep 27, 2021

By Matthew Holden, Jason Emery, Matthew Benjamin, Peter Craig

Over the last 25 years, through adversity and a growing level of risk mitigation, Australian Longwall mines have been using higher and higher capacity cable bolts for secondary support. A recent survey revealed 85% of Australian longwall cable bolts are a 70 tonne capacity pre-tensioned hollow post groutable system. Mines avoiding pumping cementitious grouts are sometimes using pre-tensioned resin encapsulated solid cable bolts of over 90 tonne capacity. Results from various industry and University research has led to manufacturer's fine tuning the construction of their cable bolts. A review of this research explains how smooth wire construction is now dominant. Aggressive shear has also led to some mines using 8 tonnes pre-tension instead of the previous common 20 tonnes pre-tension. Lower pre-tension acceptance is now leading to a greater use of threaded cable bolt end fittings, these are tensioned to 8 tonne using the drill motor torque. A recent development is the TEX cable bolt, a slim line threaded end fitting that fits with the drill hole like a standard rock bolt. This cable is being installed at the coal face as primary support off bolter-miners. The laboratory and in-situ testing of the TEX cable is presented. In a bid to improve gateroad development rates and ground support in the last several years, a 100 tonne tensile capacity Goliath cable system was developed. The latest development in resin anchored solid cables is a 110 tonne capacity cable bolt, with laboratory and in-situ testing results presented.

Jun 25, 2024

By Hans W. Schreiber, Rudi P. Fronk

IINTRODUCTION Widespread activities in precious metal projects in the western U.S., Australia, Canada and other parts of the world have opened a flood-gate of consul- ting expertise to aid and accelerate the development process. Utilization of outside expertise has taken on many shapes and forms that have in some instances actually "decelerated" or "retarded" project progress. Small mines are especially vulnerable to apparent false-starts and misdirections, be- cause of their generally lower capital commitments, shorter lead times and limited mineral in-house expertise. Mine development for small mines requires fundamental approaches by knowledgeable consultants that are timely and cost-efficient and in line with accelerated development, if they are to be successful. The development process is obviously that process which brings a given deposit into production. Or, put another way, the process comprises all activities, decisions and commitments which take place from the moment an authorizing entity (usually a board of directors) com- mits to development and production (usually on the basis of a final feasibility study) until the first ton of ore is processed to a saleable product. The activities, decisions and commitments over a development period--such a period can range from as short as a year or less to as long as 5 years or more--are many, highly varied, and often interdependent. The development period is also particularly risky, not only because of the varied and interdependent decisions and commitments, but also because of capital demand amounts. Delays in the development process are thus particularly costly - sometimes so costly that small and often undercapitalized mining companies, failing to reach production and a cash flow within a planned period, face, if not bankruptcy, at least severe equity dilution or particularly onerous and restrictive covenants on additional loan financing. It therefor behooves especially small mining companies to examine ways to accelerate development. By doing so, the development process may actually meet planned scheduling. The development process generally comprises environmental impact studies completion, final engineering, permit- ting, financing, equipment and machinery procurement, contractor selection and construction -- such activities largely being overlapping. The paper addresses what the small mines operator or the small mining company can do with consulting firms to accelerate development. By "small mines" is meant open pit operations of 2000 tpd or less and underground operations of 500 tpd or less. By "consulting firms" is meant entities who do not possess an engineering and construction capability, such as Derry Michener Booth & Wahl, Pincock Allen & Holt, Behre Dolbear-Riverside, etc. The paper concentrates on gold projects. PRIORITY CONSIDERATIONS Of the items listed in the previous paragraph which comprise the development process, the two which most consistently appear to cause delays are permitting and financing. The former's delays can be due to political overtones and thus the delays may not be reasonably foreseeable. The latter's delays are most often due to inadequate preparation, and thus are avoidable. For the reason stated above, and considering that financing is fundamental to all of the activities which constitute the development process, the paper focuses on financing as the priority item which can most assist in accelerating or retarding the development process.

Jan 1, 1987

By K. O. Groves

Although chlorine oxidation has been found to be effective in destroying the CN ion, an excess of chlorine is needed to drive the reaction to completion. This often results in the formation of residual chloramines, which are toxic to aquatic life and often regulated by various states. The interhalogen, bromine chloride (BrCl), has been found to be just as effective but without the formation of persistent, residual chloramines. The hydrolysis and subsequent disproportionation of the highly-polar BrCl molecule essentially yields a solution of hypobromous acid (HOBr) and nonoxidizing hydrochloric acid (HC1). Studies have been carried out to directly compare the mechanism of CN destruction with chlorine and bromine chloride.

Jan 1, 1982

By Thomas M. Lynberg

WIDCO is a surface coal mine located near Centralia, Washington. The WIDCO preparation plant processes over 7 million tons of run- of-mine coal each year with a product yield averaging 72%. In the jig type preparation plant the fine refuse amounts to approximately 130 dry tons per hour of 28mesh by 0. This fine refuse material is thickened in a 160ft. diameter static thickener with the use of lime as a coagulant and organic polymer as a flocculant. Once thickened the fine refuse, which is predominately bentonitic clays, has been dewatered and disposed of using several different methods. These methods will be discussed and include vacuum disc filters, ponds, solid bowl centrifuges, belt filter presses and experiments with electrokinetic dewatering.

Jan 1, 1981

The structure of a mine ventilation network (M.V.N.) can be represented graphically using either the mine plan (map) itself or a number of other types of diagrams. In the U.S., the two most common approaches are the line and the schematic diagrams (see Figures 1 and 2). The purpose of this paper is to introduce another type of diagram, called the canonical diagram (6). The method of M.V.N. representation which uses canonical diagrams is used extensively in Europe, and especially in Poland, where it was developed. The canonical diagram is a noncalibrated topological image of a ventilation system. It gives a simple representation of the ventilation network, allows for the classification of currents within it and enables rapid computer analysis and computation, as well as the representation of the results. It is especially useful in studying and controlling the air flow, and it may be a life saving tool in conducting rescue operations during and after mine fires.

Jan 1, 1986

By Dieter Loh, Mathias Knoll, Florian Werres

When a new tunnelling project starts, it has been a vital tradition for decades to cast and erect master rings horizontally on a flat surface to demonstrate the compliance with dimensional tolerances. Some tenders ask for the erection of two or even three rings upon each other or repeated master ring erections at intervals throughout the project. Especially for large diameters, this is a challenging and risky task for the workforce in the segment factory. A truck-mounted crane is often necessary, and in some cases it can take up to a week to complete a master ring. The concept of virtual rings has the potential to replace this tradition by assembling master rings digitally instead of physically. This is based on a best-fit approach, considering the 3D coordinates from laser tracker measurements of segments or segment moulds. Building a virtual ring, a sub-millimeter accuracy can be achieved, the alignment of neighboring circumferential and ring joints observed, and the fit of bolt holes checked. The traditional method, on the other hand, is discussed controversially despite the significant amount of work and risks—particularly as the horizontal position of the ring and absence of gaskets do not simulate the reality in the excavated tunnel. This paper will highlight the technical aspects of virtual rings, as well as their pros and cons, illustrated by a case study from London Thames Tideway where a total of 30 physical master rings could be replaced by virtual rings.

Jun 13, 2023

By George Hufford

An investigation is made into the applicability of queueing theory to simulating the operation of a large open pit copper mine in the southwestern United States. Actual data from a working mine was used to construct curves for truck loading time, haulage and return time, and dumping times. Known results of investigations of queueing models (Koenigsberg - 1958, Gordon and Newell - 1967, Cabera and Maher - 1973) were used to simulate the actual operation. These did not give acceptable results so commonly used probability distributions were fitted to the actual data. A large computer program known as Q-GERT was used with the various parameters obtained from fitting the curves. Kolmogorov- Smirnov tests were run that showed that this was not satisfactory either. Finally, Q-GERT was used with the actual data from the mine. An appropriate correction factor is then found to apply to the exponential probability distributions tested previously and satisfactory results are obtained.

Jan 1, 1981

By Garry G. Lowder

The regional geologic setting is essentially a volcanic island arc built upon an oceanic basement, consisting of basalt and minor pelagic sediments (the Tinombo Formation of Eocene - lower Miocene age. Trail et. al. 1974). -- An east-west chain of intermediate and acid calc-alkaline volcanoes (the Bilungala Volcanics) was built on the oceanic basement during the Miocene. The Bilungala Volcanics are flanked to the north largely by contemporaneous andesites and volcanoclastic sediments of the Dolokapa Formation. During middle to upper Miocene time the Bilungala Volcanics were invaded by large masses of predominantly dioritic intrusive rock and it is within one of these dioritic intrusive complexes, the Bone Diorite, that the cluster of porphyry copper systems occurs. Renewed volcanic activity in the Plio-Pleistocene produced acid pyroclastic rocks (the Pinogu Volcanicsi whose distribution is controlled by present-day topography. Structural trends, largely interpreted from SLR imagery, are pre- dominantly NW or NIWJ. The right-lateral Gorontalo Fault, which trends hW through the town of Gorontalo (Katili, 19691 is the most prominent structural feature. Many of the major structures in the Tapadaa district have similar orientations and may be sympathetically related to the Gorontalo Fault.

Jan 1, 1976

By Sheldon P. Wimpfen

SHELDON P. WIMPFEN The measure of the success of any mining operation is the size of the difference between sales realization and cost of product, including operating, sales, administra¬tive, capital and others. Evaluation of a mining project necessitates projecting costs and sales realization as accurately as possible. Possible variations in specific cost elements and sales prices must be considered. To accomplish this, project studies should include giving effect to possible fluctuations to determine the break¬even points with varying costs and prices. Development of cost data is never an easy task. Mining is characterized by a great variety of conditions which result in still more variables. Achieving realistic figures as of any specific time involves consideration of changes, which may be frequent, in labor-contract terms and equipment used, variations in working condi¬tions and mining practices, and changes in mineralization and metallurgy. The purchasing power of the dollar, which has been declining persistently, also affects the situation at any given time. The only thing that can be said of costs, with certainty, is that they will change. With this in mind, the "guideline" data in this section are offered for the use of the mining engineer. It was provided for the most part by representative companies, whose contributions are gratefully acknowledged. In reviewing the tabu¬lations and other material it is of course necessary to remember that the data presented are valid only for the time periods specified, and will change with changing economic conditions in any other. 31.1-CAPITAL COST OF PLANTS Plant capital costs are set forth in Tables 31-6 and 31-7 by year of construction. To estimate capital cost of plants built in future years, various cost bases and indexes are utilized. Design and Construction Manpower: Monthly Salary Ranges, January, 1970 Project manager $1,400-82,600 Project engineer1,000- 1,500 Project design engineer1,050- 1,800 Senior engineer1,000- 1,700 Senior design engineer1,100- 1,700 Senior engineering designer1,000- 1,300 Layout draftsman750- 1,000 Detail draftsman650- 1,000 Material engineer900- 1,400 Construction-project manager1,500- 2,500 Project engineer1,500- 2,500 Construction superintendent 1,500- 2,500 A rule of thumb in design work is one man-month per drawing. The number will vary for each specific plant. For a recently constructed 30,000-tpd (tons per day) copper concentrator, 1,800 drawings were made. Design should provide for the addition of milling units when these may become justified in relation to investment markets and life of the deposit.

Jan 1, 1973

By Richard D. Nichols

A new method of slaking lime has been developed by DENVER EQUIPMENT DIVISION of Joy Industrial Equipment Company and patented under U. S. Patent 4,261,953. The new method utilizes an attrition machine, commonly used in mineral processing circuits as a sand scrubbing device, to slake lime. The scrubbing effect of the attrition slaker continuously produces a fresh reactive surface so that further slaking may occur. This process allows for utilization of slaking all grades of burnt lime and most grades of dolomite lime. Testing indicates that poor grades of water usually not considered for this service can now be tolerated using the attrition slaker. The slaked lime produced is of higher quality, faster reacting and much less apt to produce scaling problems than previously experienced. The attrition slaker can be combined with a spiral classifier and ball mill to eliminate any need for grit handling or grit removal. The attrition slaker can be designed for all burnt lime tonnages up to the limit of 175 tph and 1" pebble lime size. Control can be maintained easily with minimal operator assistance. The equipment, the principals of process metallurgy, and the above topics will be presented.

Jan 1, 1982

By Gene E. McClelland, Dirk J. A. van Zyl

5.1 INTRODUCTION Heap leach cyanidation techniques possess considerable potential for exploitation of low-grade ores, small ore bodies, mine strip waste and tailing materials where fine grinding is not necessary for good extraction. For heap leaching to be successfu1, ore must exhibit certain characteristics, as previously discussed in Chapter 1. Primary among the necessary ore characteristics is that the ore must have good permeability after being crushed or treated and stacked into heaps if uniform distribution of the cyanide leach solution is to be achieved. Gold and silver ores containing excessive amounts of clays or fines generated by crushing are some of the most difficult to treat successfully by heap leaching. The presence of excessive amounts of slimes (generally classified as minus 50-micron or 270 Tyler mesh sieve size particles) in the heap leach feed will slow the percolation flow of the leach solution, cause channeling, or produce dormant or unleached areas within the heap. This may result in unreasonably long leaching periods and poor extractions. In extreme cases, the clays or slimes can completely seal the ore heap, causing the leach solution to run off the sides of the heap rather than to penetrate the ore heap. The problem of heap leaching ores containing fines can be aggravated during stacking and preparation of the ore heaps, because natural sorting of coarse and fine material occurs during these steps (Johanson, 1978). This phenomenon results in a concentration of ore fines at the center of individual ore piles and a concomitant concentration of larger rock fragments on the lower slopes and base of the piles. When the individual piles within the heap are leveled off prior to installation of the sprinkling system which delivers the leach solution, additional segregation occurs as the fines sift through the coarser ore particles. The segregation results in localized areas or zones with marked differences in permeability. As a consequence, the leach solutions follow the path of least resistance, percolating downward through the coarse ore regions and bypassing or barely wetting areas that contain large amounts of fines or slimes. Effective utilization of marginal gold and silver resources through heap leach processing requires the development of new methods in order to achieve more uniform size distribution during ore heap preparation and better slime control during leaching.

Jan 1, 1988

By K. -A. Björkstedt

INTRODUCTION Strassa lies in the central part of Bergslagen, a tradi¬tional mining district, on the eastern side of the Stora Valley at an elevation of about 200 m above sea level. A railway siding runs between the mine and the Stora railway station from which there are railway connections to the shipping port and iron and steel works in Oxelosund, about 224 km away. The distance to the provin¬cial capital Orebro is about 60 km. The climate is typi¬cal for this part of central Sweden and is illustrated by the diagram of monthly precipitation and temperatures for the years 1968-1975 (Fig. 1). HISTORY There is no certain information as to when the Strassa mine was first worked, but it is known from sur¬viving accounts of mine inspectors that there were smelt¬ing works in operation in nearby villages in the 12th century. An example is the Gusselhytta ore smelting works, 10 km south of Strassa, which dates from this period. Around the year 1540 there were two smelting works in Strassa, the Upper Karberg and Lower Karberg works. Ore for these smelters was probably taken from Strassa and from the adjacent Blanka mine. In the year 1624 Strassa is mentioned by the painter Jons Nils Krook in an account of the iron mines in the Linde mining district (Linde Bergslags Jarngruvor). Several mines were listed in the area, the deepest being about 30 m. An impressive power installation is mentioned in 1639, including a piston system of lashed poles for transmit¬ting power from the Stora River to the Strassa fields. Its length was 2670 m. Common ground comprising about 20.2 km2 (5000 acres) of forest was allocated in 1689 for the furtherance of mining operations. Until the beginning of this century only the rich cen¬tral parts of the ore body were mined and these yielded, after handpicking, lump ore suitable for smelter feed. An example of the ore grades from these early times is an analysis of ore from the "Big Mine" (Storgruvan) from the year 1873: 48.5% Fe, 0.008% P, and 0.06% S. This same year a total of about 18 000 t was ex¬tracted from the Strassa mine. OWNERSHIP The mine was owned and run until 1874 by a min¬ing association made up of 119 so-called "bergsman," who were homesteaders often engaged in agriculture and timber-cutting as well. In that year the Strossa Grufvebolag (Mining Co.) was founded. In 1906 it was con¬verted into a joint stock company, the Strossa Gruveaktiebolag. This was acquired in 1907 by Metallurgiska AB for the implementation of Gustav Grondal's beneficiating and briquetting methods, for which the Strassa ore was well suited. The same year saw the completion of a new ore dressing plant with an annual production of 46 000 t of ore concentrate. In 1911 the mine passed to new hands, and in 1913 it was purchased by an Austrian company. Extensive new installations were made and in 1915 a new dressing and briquetting plant was completed with twice the capacity of the old one. In 1917 the Strassa mine was acquired by Granges. Be¬cause of unfavorable business trends and technical diffi¬culties, mining operations were brought to a close in 1923. Pumping kept the mine free of water until 1933 but it was completely filled ten years later. Up to 1950 the surface buildings and installations remained intact but the large dressing and briquetting plant burned to the ground in that year. Today only the machine shop re¬mains from this earlier period of operation, now housing parts of the Mineral Processing Laboratory. The decision to take up mining operations again was made in 1955 and construction work began the follow¬ing year. Of the old installation, only the "southern shaft" could be used for some development drifting after it had been completed with a new headframe. Other¬wise, all the buildings and installations required for the operations had to be rebuilt. New installations ready by 1960 were office and personnel facilities, a new shaft and headframe, a sorting and concentrating plant, a macadam plant, settling basins, pump stations, and a railway and yard with transport equipment. The instal¬lation was completed with two plants

Jan 1, 1982

By None None

This month, Mining, Metallurgy & Exploration (MME) is launching with a special issue on mineral and metallurgical processing in honor of the 90th birthday of Prof. Emeritus Douglas W. Fuerstenau. Headed by a special review article by Prof. Fuerstenau, this special issue contains 22 papers by the top experts in the field.

By J. B. Zachwieja, P. E. Richardson, G. W. Walker

The electrochemical nature of mineral-collector interactions and their relationship to flotation have been investigated to develop a better understanding of the factors controlling sulfide mineral beneficiation. Spectrophotometric and electrochemical techniques have been used to examine the reactions occurring on electrodes composed of a bed of bornite particles in the presence and absence of potassium ethylxanthate (KEtX) at a pH of 9.2. The combined techniques showed that KEtX sequentially reacts with bornite by (1) a chemisorption reaction forming ca hydrophobic species between --0.4 and -0.2 V (SCE), (2) an electrochemical oxidation reaction forming cuprous xanthate between -- 0.2 V and the xanthate/ dixanthogen redox potential, and (3) oxidation to dixanthogen when the potential exceeds the reversible value for the xanthate/dixanthogen couple. The flotation response of bornite closely parallels spectrophotometric and electrochemical evidence for adsorption of KEtX by the reactions listed above. The degree of surface oxidation resulting from the formation of iron oxides on the surface dictates the flotation response as evidenced by the fact that oxidized bornite floats at more anodic potentials than unoxidized bornite. The strong correlations between the flotation response, degree of oxidation, and potential indicate that chemisorbed xanthate produces a hydrophobic layer sufficient to float unoxidized bornite, but cuprous xanthate and dixanthogen arc required to float oxidized bornite.

Jan 1, 1988

By H. G. Fleshman

?Wyoming bentonite? has become almost a world-wide trade mark for high quality sodium bentonite. Although bentonite is produced in more than a dozen U.S. states the largest producing area covers Wyoming and adjacent parts of Montana and South Dakota. Bentonite was originally known as Taylorite and later renamed Bentonite by Dr. Sam Knight, the State Geologist of Wyoming. The new name came from the Fort Benton shale member of the Mowry Formation. Bentonite is a clay consisting essentially of the montmorillonite group of minerals. Since bentonite has good ion exchange capabilities, various cat-ions are easily replaced. These cations are usually calcium and sodium; however, magnesium, lithium, potassium, and others can be exchanged. One way of classifying the clay is based on its swelling capacities when wet or added to water. Bentonite having sodium as the dominant exchangeable ion typically has very high swelling capacities and forms gel-like masses when added to water. Bentonite in which exchangeable calcium is more abundant than other ions has much lower swelling properties than sodium varieties.

Jan 1, 1987

By Fred Morelko, Michael Connelly, Brad L. Ricks

INTRODUCTION The $1.1 billion Bajo la Alumbrera Project, started up in 1997 by Minera Alumbrera Limited in the Catamarca Province of Northwestern Argentina, is the first large scale modern mining project to be developed in Argentina. Like many modern Andean mining projects, existing infrastructure was not available to economically transport the 875,000 tonnes of product per year. A variety of alternatives were investigated, including upgrading of existing roads and railroads, slurry pipelines, and combinations of these systems. The final system selected was to combine a 312 km slurry pipeline with an 820 km rail haul to a ship loading facility near Rosario. The concentrate pipeline crosses divergent terrain between the minesite at 2400 m elevation and the filter plant at 400 m elevation. The system selected represented the lowest net present cost alternative by a significant margin. BACKGROUND The Bajo Alumbrera deposit was discovered in 1947 by Professor Cetioni from the University of Tucuman. It is a large (>750 million tonnes) medium-grade (0.6%) copper deposit with high gold values (0.5 g/t*). The deposit outcrops, making it very suitable to modern, large scale, open pit mining techniques using 218 t trucks and 43 m3 shovels. Advances in gravity concentration of the gold helped to demonstrate the feasibility of the project in a study prepared by Fluor Daniel for Minera Alumbrera in 1994. The initial concentrator is the largest green-field concentrator ever constructed, rated at 85,000 t/d. Production will peak at 190,000 tonnes of copper metal and 710,000 troy ounces of gold per year, one of the top 10 copper and top 15 gold producers in the world. * Throughout this paper, the word tonne and abbreviation refer to the metric ton of 1000 kg.

Jan 1, 1998

By R. Reeves

Coal producers and utilities depend on automated coal blending systems to reduce costs and meet stringent environmental regulations. Over 30 systems installed in the United States since 1987 use on-line PGNAA (Prompt Gamma Neutron Activation Analysis) analyzers to accurately control the blending process. Experience and detailed engineering are required to integrate real-time data produced by the coal analyzer with a control system to meet demanding performance specifications. This article reviews successful installations and lists important features including interfacing analyzers with plant control systems, specifying feeders and weigh scales, and managing data.

Jan 1, 1995