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By F. Calizaya

This study summarizes the results of various numerical and laboratory experiments per-formed to study the effect of airflow recirculation on gas concentration. The numerical exercises were carried out on a simulated mine ventilation net-work. The laboratory experiments were conducted at the Mackay School of Mines' airflow monitoring system in Reno, Nevada. In both cases, the airflow was recirculated through the system by means of a main fan and a set of booster fans. The booster fans were positioned to produce two types of recirculation: In line and Cross-cut. For each type, various recirculation conditions were generated by changing the booster fan pressure. During each test, the gas emission rate at the simulated working face was kept constant. The study illustrates the conditions under which controlled airflow recirculation may be used to solve a district ventilation problem.

Jan 1, 1991

By Paul D. Bybee

The Freeport Sulphur Company operates four sulphur mines in the coastal marsh area-03 South Louisiana. Employing the Frasch process of mining, approximately 13-million gallons of water per day is heated to a mining temperature of 325°F and is injected into the sulphur bearing formations of the four mines. A volume of cold waste water, roughly equal to the Input, is removed from the formations and presents a disposal problem. Two additional mines are currently being developed and will increase the disposal volume by an additional 5-million gallons per day. Although these properties are in a relatively small area, geographically speaking, as shown in Figure 1, they each represent a separate disposal problem with the treatment methods designed for the individual mine.

Jan 1, 1959

By F. J. Hurst

Phosphate rock, which is a prime source of phosphate used in fertilizer manufacture in the United States, contains uranium. In general, the uranium content of the rock increases with the phosphate content and most high-grade phosphate deposits (>30% P2O5) contain 0.01- 0.02% U3 8. The large tonnages of rock being mined represent a significant potential source of uranium that lends itself to recovery as a by-product of the wet-process phosphate industry. For example, in 1953, the U. S. Geological Survey estimated that recoverable phosphate reserves in the U. S. totaled about 5 billion tons and contained 600,000 tons of uranium. I This estimate did not include any rock from North Carolina which has since been estimated at about 2 billion tons. For the most part, the known, economically mineable U. S. reserves - estimated at about 7 billion tons - are concentrated in a few areas: 43% in the western states of Idaho, Montana, Utah and Wyoming; 28% in each of Florida and North Carolina and less than 1% in Tennessee. In the past, Florida has supplied about 80% of the total U. S. production, practically all of which is land pebble phosphate. All forecasts indicate increased production rates from North Carolina and the western states.

Jan 1, 1976

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

By Alfred L. Anderson

Thorite and rare earth deposits in the Lemhi Pass area of southeastern Lemhi County, Idaho, are contained along complex shear and fracture zones and in reopened gold-quartz veins and copper lodes in quartzitic and phyllitic rocks of the Pre-cambrian belt series. Some of the shear and fracture zones are several tons of feet across and comprise broad zones of irregularly mineralized rock. These lodes contain a notable concentration of thorium and rare earth elements along with considerable amounts of barium, .potassium, calcium, iron, aluminum, phosphorus, sulfur, and silicon. The principal minerals are thorite, locally allanite, monazite, xenotime (?), apatite, specular hematite, barite, potash feldspar, occasionally calcite and quartz. Exept for some of 'the specular hematite, barite, reddish feldspar, calcite and quartz, the minerals are not distinguishable without the microscope. The thorite and locally allanite were deposited first, then the monazite and other phosphates, which were in turn followed successively by barite, potash, feldspar, calcite, and quartz. In exceptional cases, deposition of thorite and specular hematite was repeated after the feldspar. Some of the deposits contain several percent thoria, but the average os generally under one percent. The rare earth content is comparable with that of the thoria. The deposits appear to be of hydrothermal origin and are probably related genetically to late Cretaceous or early Tertiary igneous activity, manifested locally by a few dioritie and lamprophyric dikes.

Jan 1, 1959

By T. Tamagawa, I. Iwasaki, N. X. Fu, M. Kobayashi

A three-stage fluidizing-bed process consisting of chlorination, selective oxidation and sulfur chloride conversion was tested for processing chalcopyrite concentrate. It was found that the use of excess oxygen was essential for the stable operation of the oxidation reactor. The excess-oxygen gas inevitably causes the oxidation of the sulfide concentrate in the chlorination and sulfur chloride conversion reactors. A new flowsheet is proposed to separate the chlorination and selective oxidation steps with separate flow paths of the fluidizing gases, thereby, preventing the passage of excess-oxygen gas in the chlorination and sulfur chloride conversion reactors. The feasibility of the proposed flowsheet was demonstrated in a laboratoryscale fiuidizing bed setup.

Jan 1, 2002

By Rafael Moure

On April 21, 1980 the Oil, Chemical and Atomic Workers Union (OCAW) submitted a petition urging the revision of the Mine Safety and Health Administration (MSHA) permissible exposure limit to radon daughters in underground mines. This petition asks for: 1. Annual worker exposure to be lowered from the current 4 WLM/year to 0.7 WLM/year; 2. Full economic protection for workers whose exposure exceeds the proposed 0.7 WLM/ year; and 3. The primary means of lowering workers' radiation exposures shall be environmental controls. I will present the scientific rationale for the lower standard requested in our petition to the Department of Labor. [1. Radiation Exposures to Uranium Miners] Two types of radiation exposures are present in the environment of uranium mines: A. [External Exposure] to gamma sources which until 1978 was considered as negligible compared with internal alpha radiation. R. E. Utting (1), in an investigation of three Canadian uranium mines, found levels of exposures adding to an annual dose equivalent of 3 rems (30 mSv)/year. He concluded that gamma dosages may be a major fraction of the maximum permissible dose for many individual miners. The U.S. experience has shown that there is significant gamma exposure in U.S. uranium mines and, therefore, it could contribute to the total radiation exposure to the lungs in combination with internal alpha radiation. Before ventilation improvements in uranium mines, the contribution of external radiation to total lung dosage was relatively small. After improvements in ventilation, the reduction in internal radiation exposure appears to be in 1980 as low as 1/10 of the 1950 dosages. However, ventilation improvements do not affect external gamma radiation. Gamma radiation remains at the same 1950's level, and it represents today a higher proportion of the total radiation burden to the miner's lungs. MSHA regulations (57.5-47) state that if gamma exposure is below 2.0 milliroentgens per hour, then workers are not required to wear gamma dosimeters and cummulative gamma radiation records need not be kept. Assuming an exposure of 1.9 milliroentgens per hour per 173 hours/month for 12 months,a miner would accumulate a dose from external radiation of 3.9 rems/ year. This dosage would, of course, go unrecorded. MSHA, therefore, should consider a downward revision of the external radiation permissible exposure limit of 2.0 milliroentgens/hour. MSHA should make clear that this exposure level, when added to the internal exposure, does not make for sound health physics -even though the measured current levels do not approach 2.0 milliroentgens/hour. B. [Internal Exposure] to alpha radiation is still the larger contributor to lung irradiation in uranium miners. The target organ is still the lung. The parameters to be considered in determining the biological radiation load appear in Table I for radon daughters present in the upper respiratory tract (URT). TABLE I RADIATION DOSE PARAMETERS IN THE URT OF URANIUM MINERS Parameter Uranium Underground Mining [1. Quantity of the Absorbed Radon daughters (Po-218, Nuclide Pb-214, Bi-214) deposit in bronchial and lung tissue. (Measured by air concentration, Working Levels) 2. Type of Radiation Emitted Alpha Particles 3. Energy of Radiation High-Linear Energy Transfer 4. Half-life of the Nuclide Ra 222, 3.8 days in the Organ Po-218 (RaA), T1/2 = 3.1 minutes Pb-214 (RaB), T1/2 = 26.8 minutes Bi-264 (RaC), Vi = 19.7 minutes 5. Selection Factor of the Direct inhalation with mine Organ dust to be deposited in the tipper Respiratory Tract 6. Mass and Form of the Organ Especially cancer sensitive, tracheobronchial region and bronchi* 7. Distribution of the Activity Dependent on deposited dust in the Organ distribution in the Upper Respiratory Tract 8. Chemical Nature of All compounds (RaA, RaB, RaC) Incorporated Substance biologically insoluble. (Only localized, irradiation) *See Reference (2)] The parameters of Table I depend, in turn, on the demography of the exposed population, especially: 1. Age 2. Sex 3. Health Status 4. Genetic Constitution 5. Lifestyle (Smoking, Diet) Some of these demographic characteristics of uran-

Jan 1, 1981

By Kyeong Woo Chung, Rina Kim, Ho‑Seok Jeon, Yeongjin Ahn, Byunghun Go, Min‑seuk Kim, Hong‑In Kim, Sujeong Lee, Yuik Eom

The study focused on the recovery of vanadium from a Korean vanadium-bearing titaniferous magnetite (VTM) ore, examining the behaviors of magnetic separation and Na2CO3 roasting-water leaching. Two concentrates were prepared using single-stage dry magnetic separation and three-step wet magnetic separation, resulting in V2O5 grades of 0.8% and 1.0% and major gangue mineral SiO2 contents of 6.5% and 1.7%, respectively. These concentrates were subsequently roasted at 1050 °C with the addition of Na2CO3 and then water-leached at 25 °C. Various stoichiometric ratios of V2O5 to Na2CO3 were tested, ranging from 1:1 to 1:40 of the theoretically required amount. As the amount of Na2CO3 increased, the vanadium leaching efficiency showed fluctuations between 4 and 35% for ratios of 1:1 to 1:10, while it linearly increased from 4 to 82% for ratios of 1:10 to 1:40 in the leaching of the low-grade concentrate. In contrast, the leaching efficiency increased from 47 to 74% for ratios of 1:1 to 1:8 and then sharply rose to 84% at a ratio of 1:10 in the leaching of the high-grade concentrate. Subsequently, it reached a plateau of 92 to 94% for ratios of 1:20 to 1:40. The difference in leaching efficiency was primarily due to the reaction of gangue minerals with Na2CO3, which hindered vanadium leaching from the VTM concentrates. This hindrance was notably more pronounced in the low-grade concentrate with the lower salt ratio, as the gangue minerals exhibited a preference for consuming Na2CO3 over vanadium. The mineral formation was analyzed in detail using XRD and MLA to shed light on the mechanism of different vanadium leaching behavior depending on the concentrate grade and vanadium to salt ratio. Based on the results of this study, it is evident that applying a proper magnetic separation procedure is significant to achieving satisfactory vanadium recovery with less salt amount by removing gangue minerals ahead.

Mar 15, 2024

By J. W. Leonard

By using physical and chemical method to increase the bulk density of newly mined coal, the productivity of a mine can be significantly increased. Greater amounts of coal can be placed in fixed volume shuttle car by increasing bulk density. This can increase mine productivity by about 15%. Small, easy to achieve changes in mining operations are required to obtain the increased productivity. This article deals with chemical methods used to increase coal bulk density, the results and discussion are based on experiments performed in the laboratory.

Jan 1, 1992

By J. L. Vanderbeek

Investigations have been conducted at Chino Mines Company into the effect that SAG mill recycle material (critical size) exerts on SAG mill throughput and specific energy requirements. Tests were conducted on the laboratory, pilot, and plant scales, to evaluate the merits of alternative treatment options. Removing critical size material from the SAG mill circuit has been shown to improve new feed rates by 34% with a reduction in specific energy (kWh/ton) of 33%. Test results obtained to date are presented.

Jan 1, 1985

By Jean-Michel Rendu

In most mining operations, some degree of selectivity is applied. Whether one deals with surface or underground operation, ore zones, and waste zones are defined not only on the basis of the value of the ore contained, but also taking mining constraints into account. A number of computer techniques are presented which can be used to determine the optimum location of the footwall and hanging wall of ore zones intersected by drill holes. Minimum mining width and minimum internal waste width constraints are taken into account, as well as other economic constraints. The system is presented as an integral part of a larger manumatic ore reserve and mine planning system.

Jan 1, 1982

By Paul Piercy

Bougainville Copper Limited operates a porphyry copper mine in Papua New Guinea. The waste dumps will contain 400 to 500 million tomes of rock at the end of nine life. The paper describes investigations carried out over a period of en years and highlights the problems associated with waste lump leaching in a tropical and seismic active region. Aspects discussed include mineralogy, chemical reaction, leaching rates, microbiological activity, dump stability, dump structures, effects of tropical conditions, seismic activity and topography.

Jan 1, 1981

By J. B. Hiskey

Thiourea is a reagent capable of complexing gold and silver. With the proper oxidant, thiourea can effectively dissolve metallic gold and silver. A thiourea leaching system possesses potential advantages over conventional cyanidation including acidic solutions instead of alkaline, cationic complexes in contrast to metal-cyanide anionic complexes, and faster leaching rates with acidic solution of thiourea. Generally, the rate of metal dissolution in thiourea solutions is fastest when ferric ion is employed as an oxidant. Acidic thiourea solutions for certain applications could offer a nontoxic and noncorrosive option for processing gold and silver. This paper provides information regarding the fundamental chemistry of the acidic thiourea system, and reviews the dissolution of both silver and gold in acidic solutions of thiourea. The recovery of gold and silver from acidic thiourea solution is discussed.

Jan 1, 1981

By E. H. Boswell

The Mississippi embayment east of the Mississippi River is part of a vast geologic and hydrologic province. The abundance of water available is largely responsible for the agricultural history of the area and is perhaps the key factor in the rapidly accelerating trend toward industrialization. The multiple artesian aquifer systems that underlie the region, and the alluvial aquifer along the Mississippi River are virtually undeveloped sources of fresh water in that they are yielding only a fraction of the water available. The potential yield of the 12 aquifer systems is estimated to be about 25,000 mgd (million gallons per day), of which only about 600 mgd is presently being withdrawn. The amount of water in storage in these systems is about 1,600 trillion gallons. Ground water having a dissolved-solids content of less than 500 parts per million is generally available at depths of less than 1,000 feet, and water having a dissolved-solids content of less than 1,000 parts per million is available in some places to depths of more than 3,000 feet. Iron is the most common troublesome chemical constituent in the ground water. Water in varying amounts is also available from streams within the region. The amount of water which originates within the region and which leaves it as streamflow during a year averages about 50 million acre-feet (about 45,000 mgd). An additional 375 million acre-feet (about 340,000 mgd) leaves the region as streamflow during an average year, this amount having originated outside of the region. Some streams have considerable flow during periods of no precipitation, and the water during these periods generally has a low dissolved-solids content.

Jan 1, 1966

By Brad Hewitt

Since Campbell Mine began processing ore, there has been a gravity gold recovery circuit. From 1947 to 1982, gravity gold was concentrated through the combination of jigs and mercury amalgamation. In 1982, jigs with concentrating table upgrading commenced and was utilized until the end of 1994. As of January of 1995 to the present, Knelson Concentrators in com­bination with concentrating tables have been employed to achieve a refinable product. An evaluation of the justification used for the last change is the focus of this paper.

Jan 1, 1996

By N. A. Sareyed-Dim

Nickel ores may be classified into two major categories : sulphide ores, and oxidized lateritic ores, these latter containing nearly 80% of the known reserves of the metal (1). Two types of lateritic ores are recognized; ores high in iron but low in magnesia and silica are known as "limonitic", while ores low in iron but containing high percentages of magnesia and silica are called "silicate" (see Table I). Sulphide ores can be beneficiated readily by conventional ore dressing techniques and the technology for nickel recovery from such ores has been well tested in a number of commercial processes (1-4). Unfortunately, no economically viable process for the concentration of nickel from low-grade deposits has yet been developed, and so only the high grade ores are able to be worked profitably. Silicate ores have been treated pyrometallurgically for the production of ferronickel (2) (3) (5); limonitic ores have been processed hydrometallurgically to produce nickel metal (2) (3) (6) (7).

Jan 1, 1975

Geostatistics is well known as a tool for ore reserves calculations at the exploration stage. The goal of this paper is to show geoscatistics can also contribute to productivity gains, allow better selection of the mining methods and better control of the grades during production. The first two examples show how geostatistics can provide sophisticated models, using conditional simulations and the use of them for simulations or optimizations of mining processes. A third example will show how to use geostatistics for short term planning in a large polymetallic vein type mine.

Jan 1, 1983

By R. O. Burt

The paper examines the role of gravity concentration in both an historical contest and within the modern mineral processing industry. It discusses the principles of the process on both a theoretical and practical level. The importance of ore mineralogy and its effect on the choice of the wide range of equipment available is discussed. The paper concludes with a discussion on the relative merits of gravity concentration to other forms of processing and seeks to answer the classical question - "Quo Vadis?"

Jan 1, 1982

By W. Gibson Jaworek

It has become apparent in recent years to everyone intimately concerned with the future of the American fuels industries that the growth in demand for energy has become a topic receiving considerable attention, and that investigators are leaning more and more toward a mathematical approach to deduce what pattern that growth will follow. Though coal has been included by necessity In all-studies of future energy markets, it has not been the subject of mathematical analysis or description as frequently as has been total energy or some of the other fuels, particularly petroleum, It is the intent in this particular paper, to subject bituminous coal consumption to a preliminary mathematical scrutiny which can lead to a more logical estimation of the general level of bituminous coal consumption for the next decade, as well as formulating an approach toward describing how coal consumption may vary as the result of the annual variations in activity in the United States, economy, The approach utilized is more sophisticated than the mere projection of the historical trend of bituminous coal consumption. Such a projection would lead to a continuation of the recent decline, indicating that coal consumption would reach zero some thirty years from now. On the other hand, the mathematical analysis described here is a relatively simple, straightforward approach and- is not intended to represent the-ultimate in mathematical economic analysis as now practiced, It is hoped that this will contribute to our knowledge of bituminous coal markets. Even more, it is hoped that it will help arouse a greater interestin utilizing mathematical and statistical analysis when studying the bituminous coal industry. - Since this is a preliminary study, comments and criticisms are anticipated.

Jan 1, 1960

By M. L. Smith

SAG mill throughput is often a bottleneck in increasing metal productivity. Improved size distribution control in the mill feed can be used to increase mill production. This can be accomplished by controlling blast fragmentation as a function of energy input and rock characteristics. Such an approach requires the integration of rock fragmentation and comminution modeling. The optimal mill feed size distribution is determined by adapting a population balance comminution model to a measure of rock hardness or fragility which is also relevant to the rock fragmentation that is achieved by blasting. The relation between the size distribution and fragility of the feed and of the mill's product and the critical size in a SAG mill is examined.

Jan 1, 1993