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By H. Dho

The role of sodium silicate in the anionic flotation of a Florida phosphate ore was investigated using laboratory continuous flotation tests, frothability tests and SEM observations. In the absence of sodium silicate, Ca++ promoted the unwanted activation of quartz by forming calcium bearing precipitates. The addition of sodium silicate enhanced the flotation efficiency through: (1) the removal of impurity minerals and calcium-bearing. precipitates from quartz surfaces by dispersion, (2) drier and more persistent froths stabilized by oily droplets containing calcium silicate precipitates, and (3) higher specific flotation rates of phosphate relative to quartz, leading to faster flotation rates and increased selectivity of separation.

Jan 1, 1990

By W. J. Miles

At the beginning of the 20th century, barite was a mineral of minor importance in the USA. Its primary use was to extend lead oxide in white paint. Barite also added the special qualities of hiding power and flatness to paint. Because white lead sometimes turns black or yellow and barite does not, ground barite became a pigment on its own merit. Barium chloride was added to bricks before firing to stop efflorescence when exposed to weather. Barium carbonate or nitrate added luster and brilliance to glass. Barium chemicals were added to refining sugar, enameling iron, making oilcloths and paper collars, and in the manufacture of paper, artificial ivory, rubber and lithopone. Lithopone became a major paint ingredient in the 1920s; its use lasted until the 1950s, peaking at 150,000 metric tons of barite in both 1937 and 1947 for the manufacture of lithopone. Later in the 20th century, titanium dioxide replaced lead oxide, barite and zinc oxide in the coatings industry. In the 21st century, barite is used as a filler, extender, or weighting agent in products such as paints, plastics, and rubber. Some specific uses include its use in brake and clutch pads for automobiles, automobile paint primer for metal protection and gloss, add weight to rubber mud flaps on trucks, and add weight to the cement jacket around petroleum pipelines under water. In the metals casting industry, barite is part of mold release compounds. Because barite significantly blocks X-ray and gamma ray emissions, it is used as aggregate in high density concrete for radiation shielding around x-ray units in hospitals, nuclear power plants, and university nuclear research facilities. Ultra-pure barite consumed as liquid is used as a contrast medium in medical X-ray examinations. Barite is the raw material for barium chemicals, such as barium carbonate, which is an ingredient in faceplate glass in the cathode-ray tubes of televisions and computer monitors. In the 21st century, barite industrial uses account for an estimated 78,000 tons per year in the United States. Another barite use was patented in 1926 by the oil industry. In this patent, barite is primarily used as a non-abrasive weighting agent for drilling for oil and gas exploration and production. As a weighting agent, it provides the downward force in a water based or oil based drilling fluid for cutting through rock with a tri-cone diamond drill bit. In order to minimize the wear on the drill bit and drill stem, minimum abrasion of the weighting agent is preferred. In recent years in the USA, 95% of the barite produced and about 80% on a world basis are used as a fluid weighting material during the search for oil and gas. During the latter part of the 20th century, barite was used in the drilling mud of oil and gas wells, particularly for wells deeper than 2000 meters, and petroleum well drilling demand for barite became the mineral?s major end use during economic expansions after World War II and subsequent increased use of oil and gas in the USA. The profitability of oil and gas production had a strong effect on the number of drill rigs and consumption of barite. The United States was the leading petroleum well drilling country starting in 1970, reaching 4,530 drill rigs in December 1981, and was the largest consumer of barite. In the USA, record highs of about 2.6 million tons of barite produced and 4.3 million tons consumed were reported in 1981. From 1997 to 2006, barite production declined from 662,000 to 589,000 metric tons. However, USA imports increased from 1,470,000 to 2,530,000 metric tons for lump barite. World production of barite increased from 6,780,000 to 7,960,000 metric tons during the same period. Table 1 (see Appendix) lists USA production, imports and exports and world production from 1997 to 2006. Major sources of imported barite include China, India and Morocco. Table 2 (see Appendix) lists the USA imports and price of barite by country. China, India and Morocco are the three largest exporters of barite to the USA and the world. Vietnam is a new producer, beginning in 2006. IMPACT OF CHINA ON DRILLING GRADE BARITE In the 1970s or before, China began supplying barite to the world drilling market in order to help develop their economy. When China first entered the barite export market, their high production volume, low pricing and low ocean transport expenses to the Gulf Coast forced barite producers in the USA to supply only regional markets, such as Battle Mountain Nevada barite for the western and Rocky Mountain states and Canada. Until recently, it was China?s government policy for domestic consumers to use barite below the API 4.20 specific gravity specification, leaving higher grades for export, but this is no longer the case. In recent years the internal demand for barite in China increased tremendously as their economy rapidly developed. China reduced their former 15% VAT tax rebates to 5% for Chinese barite producers, and then eliminated all VAT tax rebates on December 15, 2006. On the supply side, some high-grade barite deposits in China have been depleted. In particular, the low heavy metals sources are less plentiful. Additional mine capacity is being developed, but the new mines are underground and therefore costs will be higher. China?s appetite for other raw materials has tied up a large portion of the world?s bulk vessel fleet, including Panamax vessels (preferred for barite ore transport). Ocean transport freight rates have increased in recent years, and receiving ports in the USA are congested with imports of higher-added-value products. The Panamax Index, which tracks maritime shipping rates, more than doubled in 2006. China has also devalued its currency from RMB8.30 to RMB7.80 per dollar. These factors are driving up the import price of crude and ground barite for drilling. In addition, in the USA, rail congestion and high fuel prices led to higher trucking rates, increasing the delivered cost of ground barite to oil customers. India is the second largest supplier of crude barite to the USA and world drilling market. It produces all of its barite from a single deposit in the Cuddapah region of Andhra Pradesh State. More than 90% of production comes from the State-owned Andhra Pradesh Mineral Development Corp., and the government continues to closely control production, export volumes, and prices. India saw a dramatic increase in barite prices as a result of its last 3-year tender in 2004. However, in 2004 and 2005 their shipments of crude barite to the USA were not significant. In 2005, India began supplying barite pulverized to meet API specifications to the Gulf Coast. In 2006, Indian production was reduced owing to severe flooding in November 2005 that curtailed mining until mid-April 2006. Morocco has been the third largest supplier of crude barite to the USA and world drilling market. In 2003 and 2004 their shipments of crude barite to the USA were not significant. In 2005, Morocco resumed shipment of crude barite to the Gulf Coast and began supplying barite ground to meet API specifications to the Gulf Coast.

Jan 1, 2009

By A. S. Haugh

OHSAS 18001 is a series of international standards for occupational health and safety management. It is comprised of two parts, 18001 and 180029. OHSAS 18001 was formulated by international certifying bodies and is based on BS8800 and was first published in 19996,11. OHSAS 18001 was developed to be compatible with ISO 9001 and ISO 14001 in order to facilitate integration of quality, environment, and occupational health and safety management systems5. It is a voluntary standard and the overall aim of OHSAS 18001 is to enable organizations to control their occupational health and safety (OH&S) risks and improve overall health and safety performance5. OHSAS 18001 is widely accepted in the UK, Europe, and Asia, is gaining momentum in North and South America, and it was used as the basis and conforms to the requirements of ANSI/AIHA Z.10-2005: American National Standard for Occupational Health and Safety Management Systems. It defines a set of requirements for an occupational health and safety management system suitable for any kind and size of organization4. Businesses of all sizes are developing and implementing occupational health and safety management systems (OHSMS) that are in compliance with the OHSAS 18001 standard.

Jan 1, 2012

By D. J. Boron, R. Kollrack

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

Jan 1, 1987

By J. Hanna, I. J. Anazia

The Mineral Resources Institute (MRI) has developed a unique process for selective fatty-acid flotation of phosphate minerals from siliceous ores. The process takes advantage of preferential adsorption (over silica surfaces) and high reactivity of fatty-acid collectors toward apatite under slightly acidic conditions (pH 5 - 6.5). The process was successful in upgrading various Florida and North Carolina phosphate samples containing 7.7% to 15.3% P205 and 62.3% to 77.2% siliceous gangue (Insol). The resulting concentrates assayed 28.1 % to 33.1 % P205 and 4.1 % to 9.9% Insol, with high P205 recoveries of 81.8% to 92.2%. Depending on the degree of surface contamination of the silica particles in the flotation feed, one or more cleaning steps, with or without sodium silicate, would be required to achieve higher-grade phosphate concentrates.

Jan 1, 1994

By R. Mensah-Biney

The adsorption of gold-thiourea complex onto a cation exchange resin, Amberlite 252, has been evaluated in a batch stirred reactor. The observed non equilibrium saturation capacity was 3.20 meg Au/g dry resin (620 kg/t) whereas the equilibrium ion exchange capacity was only 0.53 meq/g (103 kg/t). The enthalpy of adsorption was 4 -25 kJ/mole under equilibrium conditions and increased to 20-200 kJ/mole under non equilibrium conditions. Under non equilibrium conditions, the adsorption capacity of the resin decreased with increasing temperature and increasing thiourea concentration and increased with increasing gold concentration. This dependence of adsorption capacity on thiourea concentration and temperature may be a useful tool for elution of gold from the loaded resin. Scanning electron microscopy revealed evidence of deposition of elemental gold on the loaded resin under non equilibrium conditions. The proposed mechanism for "adsorption" of gold-thiourea complex onto this cation exchange resin consists of an ion exchange reaction followed by a reduction of the exchanged gold complex to elemental gold on the surface of the resin.

Jan 1, 1997

By Ian Mitchell, Peter Faganl, Paul Haddad, Rob Dunne

During the gold cyanidation process, cyanide can be lost from the leachate by various chemical routes. The products of these cyanide losses include base metal cyanide complexes, thio­cyanate and cyanate. These losses are particularly important when treating pyritic or cupriferous ores since the two most sig­nificant cyanic ides, in order of importance, are sulfides and cop­per bearing minerals (I). Cyanidation of pyritic minerals results in the formation of thiocyanate. The problems caused by cyanide soluble copper minerals are well known to the gold processing industry and have been reviewed by several authors (e.g. 2, 3). An important consider­ation, especially with a number of more recent leaching tech­nologies, is the effect of the CN :Cu mole ratio, R, on the gold leaching kinetics. In order to efficiently leach cupriferous ores, it is important that R is maintained at a level that allows suffi­ciently rapid leaching without excessive use of cyanide. The optimum value for R in a fixed leaching system would appear to lie between 3 and 4. This is based on the early observations that the leaching rate is very slow when R is below 3, with a rapid increase in the leaching rate above 3 (2, 3,4). The leaching rate increases gradually when R is greater than 4 (4). The major Cu(I)-cyanide complex in this range of R values is the [Cu(CN)J]2· species. It has been postulated that leaching occurs due to either uncomplexed cyanide or the fourth ligand in [Cu(CN)4]3·, since the stepwise formation constant for this complex is very small (2, 3). An additional consideration when leaching cupriferous ores is the oxidation of cyanide by Cu(II) minerals, resulting in the formation of cyanate. It has also been reported that both sulfide minerals (5) and activated carbon (6, 7) catalyse the oxidation of cyanide to cyanate. From the above discussion, it is important for the leaching con­ditions to be carefully controlled when treating pyritic and cuprifer­ous ores in order to achieve efficient and economically viable leach­ing. For example, large concentrations of thiocyanate in the leachate would reveal large cyanide losses and thus indicate the need for a pretreatment step to oxidize the sulfides. Alternatively, a low R value would indicate slow leaching kinetics and consequent low gold recoveries, indicating the need for increased cyanide dosage.

Jan 1, 1996

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada's Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989

By Christopher J. Hall

It has been known for a long time that summing pressure drops and rises due to friction, shock, fans and natural ventilation is incorrect. In fact, lost works and works should be summed. In spite of this, the traditional method of finding a fan/system operating position has been the H/Q Curve. This paper presents (lost) work/m curves, progressing from a single fan moved around the system to include natural ventilation work, and fans in a series, parallel and series/parallel. The solutions obtained graphically are substantiated by dry air thermodynamic solutions, and the traditional solutions are shown to be in error. This method of solving problems is particularly useful when moving fans around the system, and essential in series/parallel and series situations with substantially different fan densities. The dry air analysis is employed because of its simplicity. Real air, with evaporation and condensation, will react somewhat differently, and one should use real densities (i.e.[l + w +w]/va ) when using this method. One source of error when applying this method to real air, although probably not major, will be investigated in a subsequent paper. Friction and shock losses are proportional to the square of the mass rate of flow, leading to the typical parabolic system curve. However, the unbalanced water stuff potential energy term, on a unit mass basis, is substantially independent of the mass rate of flow. It is possible that this could be handled as a negative natural ventilation work.

Jan 1, 1987

By G. Deschenes, M. Fulton

The New Britannia Mine processes a qumtz-carbonate gold ore containing 4.63 git Au at a rate of 2,200 tpd. The ore contains approximately 3 .0% arsenopyrite, 2. 7% pyrite and the rest is gangue minerals. Eight to ten percent of the gold is refractory to leaching. Two years ago, New Britannia Mine experienced cyanidation problems. Gold extraction was not stable and occasionally decreased significantly. Preliminary test work, at laboratory scale and in the plant, indicated that lead nitrate addition was beneficial. A laboratory study was initiated to evaluate the level of control of the parmneters (grinding size, dissolved oxygen, pH, free cyanide and lead nitrate) to optimize the efficiency of the process. It was found that lead nitrate addition is required to alleviate the effect of metallic sulphides. This addition causes a significant increase in the gold leaching kinetics and a reduction in the cyanide consumption. The optimum system appeared to be: leaching for 24-30 hours at pH 11.8, 50 git lead nitrate, 250 ppm NaCN and 8 ppm 02. The gold extraction was 89.8% with grinding at 89%-74Fm (leached tail at 0.48 git). Cyanide and lime consumption are 0.40 kglt NaCN and 1.9 kglt respectively. Following the incorporation of lead nitrate, overall gold extraction was increased from 89.2% to 91.1 % but no change was noticed in cyanide consumption. The gross profit is $CDN 988,000 per year.

Jan 1, 1999

Recent expansion programs have made the Yunkang underground coal mine the number one producer in Shenxi Province. The mine is the newest in the Datong Coal Administration, currently rated at around 3 Mt/a of coal. Located in the Wa Chushi mountain range about 23 km north of the city of Datong, the mine has undergone a series of expansion phases since startup in 1973. Initial design capacity was 1.5 Mt/a and ultimately, production will be brought to an estimated 7.0 Mt/a by 1985, according to mine superintendent Li Jia-zhi. How Coal Occurs The coal, Jurassic in age, occurs in three economically identifiable seams designated No. 2, 3, and 7. Seams are 100-250 m below the surface. Reserves from this economic zone were not quantified, but Superintendent Li indicated that there was enough mapped out to sustain the expanded production rate for over 20 years. Below the Jurassic coal is another possible economic zone, a Permocarboniferous coal lying 400-600 m below the surface. Little development drilling has been conducted to ascertain its potential. The Jurassic beds are relatively flat and coal quality is reportedly one of the best in China. Heat content of the coal runs around 33.5 MJ/kg and the ash content is less than 10%. In some areas, it is less than 7%. Coal preparation is simple and involves a washing and screening separation into four principal size fractions. Washed coal is stored in one of six storage silos, each with 4000-t capacity. Mining the No. 2 Seam Since startup, primary production has been derived from the No. 2 seam. The coal is relatively hard and is topped by a hard cement-like conglomerate. These factors, coupled with an average thickness between 4 to 6 m, precluded the use of conventional longwall techniques, according to mine engineers. The solution was development of a room-and-pillar tech-

Jan 3, 1980

By L. Yan, D. Yantek

Federal regulations require refuge alternatives in underground coal mines to sustain life for 96 h while maintaining an apparent temperature below 35 °C (95 °F). Research by the U.S. National Institute for Occupational Safety and Health (NIOSH) has shown that heat and humidity buildup is a major concern with refuge alternatives because they have limited ability to dissipate heat, and high internal air temperature and relative humidity (RH) may expose occupants to heat stress. The heat transfer process within and surrounding a refuge alternative is complex and not easily defined, analytically or experimentally. To investigate heat and humidity buildup in refuge alternatives, NIOSH conducted multiple in-mine, 96-h tests on a 10-person tent-type refuge alternative, a 23-person tent-type refuge alternative and a six-person metal-type refuge alternative. The results show that when moisture was introduced to represent perspiration and respiration from miners (wet tests), the average temperature at midheight increased by 10.5 °C (18.9 °F) and the RH approached 88 percent for the 10-person tent-type refuge alternative; the average temperature at midheight increased by 9.4 °C (16.9 °F) and the RH approached 94 percent for the 23-person tent-type refuge alternative; and the average temperature at midheight increased by 7.7 °C (13.9 °F) and the RH approached 95 percent for the sixperson metal-type refuge alternative. For the dry tests, where no moisture was introduced, the average internal temperature increased by 12.6 °C (22.7 °F) for the 10-person tent-type refuge alternative, by 10.3 °C (18.5 °F) for the 23-person tent-type refuge alternative and by 8.4 °C (15.1 °F) for the six-person metal-type refuge alternative. These results may provide refuge alternative manufacturers and mine operators with guidelines and considerations for evaluating temperature profiles for portable refuge alternatives. The information may then be used to make decisions on occupancy ratings and heat mitigation strategies based on the thermal environment in which the refuge alternatives will be installed.

Jan 10, 2018

By Kegomoditswe Koitsiwe, Tsuyoshi Adachi

"This paper characterizes quarterly LME copper price fundamentals from 1995 to 2011 by analyzing the transformation of the market mechanism based on structural change perspective. Using chow test for structural change based on the least square multiple regression, we divide the price fluctuation into: Stable price period (1995Q3-1999Q2), Low stable price period (1999Q3 – 2003Q3) and Price fluctuation period (2003Q4 – 2011Q2). The results show the existence of structural breaks disproves the investigation of the full sample period as a whole. In different structural breakpoints the main drivers of copper prices changes and their impact are significantly different. Moreover, the paper examines the role of speculation in copper price fluctuations.Introduction The rise in commodities price volatility and financial and economic crisis have led to questions over the organisation of futures and physical commodities markets. Spikes in commodity prices and volatility have coincided with a surge in financial speculators in the commodity markets inspiring a polarised and hotly debate among academics and practitioners about the role of investors in the development of physical markets. In his seminal work Kaldor (1939) focused primarily on whether speculative activity in the futures markets reduces the variance of commodity spot prices. Yet, part of the literature finds that the introduction of futures contracts destabilizes the spot market. IMF (2006) studied commodities boom and its causes concluding that there is little support for the hypothesis that speculative activity affects either price levels over the long run or price swings in the short-run. The impact of speculative activity on commodity prices remains a highly contentious topic. Other market analysts take the view that the supply and demand for physical commodities remain the major determinants of market trends (Krugman 2008). It is the aim of this study to contribute to the literature by examining the drivers of copper price changes from a structural change perspective. In this context we are attempting to answer to: What are the main drivers leading to copper price fluctuations? Are these factors having the same effect in different periods or not? It is meaningful to understand drivers of copper prices and thereby, their fluctuations because it will help developing countries depending on copper export to anticipate market situation and avoid bad economic development challenges that can result from price fluctuations. Moreover, this can ensure sustainable mineral supply."

Jan 1, 2017

By Shane Yanagisawa

The Mineral Creek Diversion Project is at the Asarco Ray Mine Complex, about128 km (80 miles) east of Phoenix, Arizona. The new tunnel will intercept the waters of Mineral Creek about 2,134 m (7,000 ft) upstream from an existing tunnel. The 5.74 m(18.83 ft) hard rock TBM will break into the old tunnel about 2,162 m (7,090 ft) in from the old portal after mining 3,888 m (12,753 ft). The TBM will mine through a series of igneous and metamorphic formations including volcanic tuff, soft volcanic ash, massive diabase, fractured quartzite, and schist. The TBM will hole through into the existing 4.88 m (16 ft) dia. horseshoe tunnel where the TBM will be disassembled and removed from the tunnel.

Jan 1, 2001

By L. Alagha, J. L. Corchado-Albelo

To address the U.S. Department of Energy’s critical minerals strategies regarding rare earth elements (REE) in the U.S.A. This study evaluated apatite-rich mine tailing from the historic Pea Ridge Iron Mine (PR) for possible enrichment of REE by the froth flotation process. Comprehensive characterization studies were performed on the tailing samples to quantify the REE content, identify the REEbearing minerals and their liberation behavior. The aforementioned characterization studies were: x-ray fluorescence analysis (XRF), inductively coupled plasma mass spectrometry (ICP-MS), and TESCAN’s Integrated Mineral Analysis (TIMA). The ICP-MS analysis showed that the concentrations of heavy-REE and light-REE were 8000 ppm and 3000 ppm, respectively. TIMA showed that monazite was the major REE-bearing mineral. Additionally, it was shown that monazite and other REE-bearing minerals were strongly locked within larger apatite grains. Findings from the characterization studies suggested that REE could be enriched using phosphate minerals’ direct froth flotation process. Flotation feasibility tests were performed using selected organic and inorganic iron oxides’ depressants, including polyacrylamide (PAM), in-house synthesized hybrid polyacrylamide (HyPAM), chitosan, sodium silicates, and corn starch. Results showed that the depressant’s efficiency at pH 9 was in the following order: HyPAM > chitosan > PAM > sodium metasilicate > corn starch. Zeta potential studies on the selected depressants indicated that chitosan had the highest adsorption on magnetite’s surface at pH 7. Therefore, additional batch flotation experiments using chitosan as an iron oxides depressant were conducted as a function of chitosan dosage. Results indicated that the highest P2O5 recovery of 55% and grade of 28 wt.% were obtained at 150g/t of chitosan. However, the concentrate’s highest total REE enrichment of 1.07% was achieved at a 200 g/t dosage of chitosan. The low enrichment of REE highlighted the complex mineralogy of the PR tailings and the need to implement additional beneficiation strategies before and after the froth flotation process.

Mar 2, 2022

By G. M. Wilemon

The U.S. Bureau of Mines has conducted preliminary investigations on the leaching of phosphate ores obtained from deposits in the western United States using sulfuric acid (H2SO4) in the presence of methanol. Phosphate extractions in excess of 80 pct were obtained on an ore sample from Pocatello, ID. Initial leaching studies using the ore as received yielded extraction levels of approximately 35 pct. However, when the sample was ground for as little as 5 min in a rod mill, in methanol, then treated with HzS04, 82 pct extraction was achieved. The R2O3-P2O5 (R2O3 = Al2O3 + Fe2O3) ratio of the starting material was 0.11. Solubilized products after leaching exhibited an R2O3-P2O5 ratio of 0.02, which is well within desired limits. These results illustrate the potential of acid-alcohol leaching for recovering phosphate values from marginal-grade ores and as an alternative to methods involving current beneficiation techniques that yield lower phosphate recoveries.

Jan 1, 1992

By O. M. Johnson

Barite and bentonite are the two principal components of drilling fluid systems used in the petroleum industry for oil and gas well drilling with rotary-type rigs. Approximately 87 percent of the world's annual production of barite is used in drilling applications. Of the worldwide production of bentonite, an estimated 30 percent is used. The drilling fluid, or drilling mud, to which barite and bentonite are added, performs five distinct functions: Lubrication-The liquid phase of the drilling mud lubricates the drill bit, preventing premature failure of the cutting edges and bearings. it also helps to soften the formation being drilled. Cuttings Removal-A mud system must be able to carry rock cuttings broken loose by the bit up to the surface where they can be removed by mechanical or chemical methods. The fluid's ability to perform this function depends upon maintaining a proper balance between the rate of circulation of the fluid and the colloidal properties required to entrain the cuttings in the mud to prevent their settling to the bottom of the hole. Hole Stability-The pressure exerted by the mud against the wall of the bore hole prevents the collapse of the uncased portions of the hole. Downhole Information-The fluid characteristics of the mud should be such that good cuttings samples, cores and electric logs may be obtained to help evaluate a well's productive zones and potential. Control Oil, Gas and Water Flows-It is essential that a drilling fluid offset any pressure from naturally occurring formation fluids in order to prevent a potentially wasteful and dangerous blowout.

Jan 1, 1980

By C. Bryce

Rhenium is primarily recovered as a byproduct from molybdenum roasting circuits. However, rhenium is also present in copper heap leach solutions at very low concentrations, and this represents a potential new rhenium source. Freeport-McMoRan (FCX) has developed a new process using carbon adsorption followed by ion exchange or solvent extraction to concentrate rhenium from less than 1 ppm to 15 g/L. The product is compatible as a feed to an existing rhenium processing plant. A pilot plant operated at Freeport-McMoRan?s Sierrita Mine has demonstrated the technical and financial feasibility of this new process. Copper heap leaching dissolves more than just copper from the copper ore. Several metals can be identified in leach solutions at concentrations of financial interest, rhenium in particular. Rhenium spot prices ranged from a yearly average of $1,070/kg ($483/lb) in 2005 (Magyar, 2008) to $10,400/kg ($4,706/lb) in 2008 (Polyak, 2013). The spot price in September 2013 was approximately $3,000/kg ($1,357/lb) of rhenium (Metal Bulletin, 2013). Several facilities at Freeport-McMoRan have very low rhenium concentrations in their copper leach solutions, ranging from 0.1 to 0.7 mg/L. The large quantity of solution available, if coupled with an attractive rhenium price, means that the rhenium contained in leach solutions may represent a significant value if a method of extraction could be developed. Testing has confirmed that rhenium adsorbs onto activated carbon much more selectively than other metals in the pregnant leach solution (PLS). A 100-gpm pilot plant was constructed to verify the lab-scale results on a larger basis and establish operating parameters for a full-scale operation. This report summarizes the initial technology development and pilot plant operation.

By Sukumar Bandopadhyay

In recent years, research in the field of artificial intelligence (AI) has had many important successes. Among the most significant of these has been the development of powerful new computer software known as the "expert systems". Expert system programs designed to provide expert-level consultative advice in mineral exploration (Duda et al, 1981), scientific (Feigenbaum, Buchanan and Lederburg, 1971) and medical (Shortliffe, 1976) problem solving are generally acknowledged to be among the forerunners of this research. As the decision makers virtually in all fields face a more complex and involved world within which to operate, the need for some decision support is also becoming urgent. Thus applications of expert systems continue to spread out reaching problems that, because of their dimensions or particular aspects, set more demand on the decision methodologies. To meet these new requirements, many activities which were performed in the past by the domain expert or engineer must become automated. Furthermore, previous research (Dawes and Corrigan, 1974; Dawes, 1979) has shown that automating an expert's decision rules often provide better decision than the expert does. For large systems it would be very useful (or even necessary) to have formal tools, allowing one for example, to automatically discover inconsistencies, contradictions or redundancies, or to identify the possibilities of wrong lines of reasoning in the decision process.

Jan 1, 1990

By H. Köse, C. Tatar, C. Helvaci, F. Simsir

The Çayirhan coal basin is located 124 km (77 miles) west of Ankara,Turkey. With approximately 520 Mt (570 million st) of lignite reserves, the Neogene Çayirhan basin is one of the important coal districts of the central Anatolia region. The Çoraklar formation consists of crossbedded sandstone, channel fill conglomerates, mudstones, carbonaceous shales and volcanoclastic pebblestones, which are fluvial in origin. The lower and upper lignite seams occur, respectively, in the lower and the uppermost parts of the Çoraklar formation. The lignite deposits are in the Çayirhan and Koyunagili portions of the Beypazari Neogene basin. The Çayirhan coalfield consists of lower and upper lignite seams. The lower lignite seam does not outcrop at the surface. Its lateral discontinuity and relatively low sulfur content, plus sedimentologic and stratigraphic features of the associated sediments, indicate a fluvial origin. Thus, the lateral spreading and geometry of the lower lignite deposits were controlled by the channel, flood plain and well drained marshes of a fluvial system. The mining method employed at Çayirhan is the fullymechanized, retreat, working longwall with caving system. Due to ventilation, haulage and tectonic conditions, panel lengths from 0.6 to 1.7km (0.4 to 1.1 miles) are chosen, and 220-m (720ft) face lengths are used. The annual coal production is 2.2 Mt (2.4 million st).

Jan 1, 1994