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By W. C. Shanks

Metalliferous sediments and/or saline brines have now been discovered in numerous sea floor deeps along 900 km of the axial rift zone of the Red Sea (Fig. 1). The Atlantis II Deep remains the most interesting of the the brine deeps both in terms of intensity of submarine hydrothermal processes and in terms of potential economic importance. In fact, successful completion of a pre-pilot mining test in 1979 indicates that the Atlantis II Deep may become the first deep sea base metal sulfide mine (Mustafa and Amann, 1980). Under the most optimistic scenario, the Atlantis II Mine might be established 4-6 years after completion of a year long pilot mining test in 1983-84. If established, mining would produce about 100,000 tons per day of wet mud with about 200 tons per day contained zinc, and proportional amounts of Cu and Ag. Expected mine life would be about 15-20 years (Anon., 1981). In addition to direct economic importance, the Red Sea brines and metalliferous deposits provide significant insight into exploration for sea floor brine-related ore deposits and for analogous ancient deposits. In this paper I will concentrate on the economic potential of Red Sea deposits and sea floor exploration for brine related deposits, but some ramifications of relevance to ancient deposits are extant.

Jan 1, 1982

By M. L. Harris, J. Organiscak, S. Klima, I. E. Perera

"The National Institute for Occupational Safety and Health (NIOSH) Pittsburgh Mining Research Division (PMRD) conducted underground evaluations in an attempt to quantify respirable rock dust generation when using an untreated rock dust and a rock dust treated with an anti-caking additive. Using PDM 3600s and 3700s, these evaluations measured respirable rock dust levels arising from a flinger-type application of rock dust on rib and roof surfaces. Rock dust with a majority of the respirable component removed was also applied in NIOSH’s Bruceton Experimental Mine (BEM) using a bantam duster. The respirable dust measurements obtained downwind from both of these tests are presented and discussed. This testing did not measure miners’ exposure to respirable coal mine dust under acceptable mining practices, but indicates the need for effective continuous administrative controls to be exercised when rock dusting to minimize the measured amount of rock dust in the sampling device. INTRODUCTION Two factors have combined to renew concern regarding respirable dust measurements and compliance. This renewed recognition is due to two regulations that have modified operators’ rock dust schedules. First, the 80% incombustible requirement in intake entries has increased both on-section and outby-the-section rock dusting. Secondly, the recent respirable dust regulation change that has increased the number of respirable dust samples taken per sampling cycle. Previously, when only 5 samples were needed bi-monthly, rock dusting could be scheduled to not interfere with the respirable dust sampling schedule. The significant number of samples now required under new regulations combined with the need to maintain 80% total incombustible content in active mining areas require the schedules for respirable dust sampling and rock dusting to intermix. This raises concerns about the amount of rock dust that potentially is included in the sampler weight gain. Additionally, the respirable dust maximum was lowered from 2.0 mg/m3 to 1.5 mg/m3 (effective August 1, 2016) increasing concern that rock dust contamination may be an issue. Currently, the acceptable practice is to apply administrative controls such as keeping the miners upwind during rock dust application or to schedule rock dusting during a maintenance shift when the number of miners present is lower. NIOSH’s research examining the quality of rock dust has led to the development of hydrophobic rock dusts."

Jan 1, 2017

By Frank E. Dalton, Bart T. Lynam, Forrest C. Neil

The Tunnel and Reservoir Plan (TARP) is the largest tunneling program in the history of the metropolitan Chicago area (Figure 1). Between September 1976 and September 1977, rock tunnel contracts having an estimated cost of approximately $420 million will be advertised. During the past six months, the Metropolitan Sanitary District of Greater Chicago (MSDGC) has awarded contracts for 16.5 miles of rock tunnels in diam of from 9 ft to 30 ft having a total bid price of $120,261,750. RECENT CONTRACT AWARDS Addison-Wilmette Tunnel $63,140,480 Upper Des Plaines #20 Tunnel 35,749,664 Upper Des Plaines #21 Tunnel 21,371,607 Addison-Wilmette Tunnel The first contract was advertised in September 1975. This contract, described as the Addison-Wilmette Tunnel, runs from Wilmette Harbor at Lake Michigan under the North Shore Channel to Addison Street in the City of Chicago, a distance of 10 miles. The contract calls for constructing approximately 27,750 linear ft of 30 ft ID and 24,020 linear ft of 22 ft ID unlined rock tunnels, excavation of two construction shafts, and the excavation of 32 drop shafts of various diam. Two hundred ninety-three companies were notified of this contract being advertised, 70 companies took out contract specifications and 3 companies ultimately submitted bids. The contract was awarded to a joint venture of:

Jan 1, 1976

By A. D. Davis, G. A. Zabolotney

A ground-water flow model was developed to simulate the postmining recharge rate from precipitation at the Belle Ayr surface coal mine, which is located in the Powder River Basin near Gillette, WY Wyoming state regulations require that reclaimed lands should provide a rate of recharge that approximates the premining rate. Steady-state calibration and transient simulations showed that the postmining recharge rate through spoil is at least 50 mm (2 in.) per year, exceeding the original infiltration rate of 4 mm (0.16 in.) per year to shallow overburden aquifers. Major reasons for the increase include the flat topography of the reclaimed areas and the decreased spoil density. It is expected that the backfill will disaggregate and settle over time, eventually resulting in a nearly constant recharge rate that is less than 50 mm (2 in.) per year but still somewhat greater than the premining rate.

Jan 1, 1997

By M. Measures, W. A. Napier, C. Pomroy, J. M. Jardine

INTRODUCTION In most uranium mines the concentration of uranium in the ore is low enough that inhalation of dust does not present a radiological problem. Conversely in the mills there is a potential inhalation hazard from inhaled uranium in concentrated form, namely "yellowcake" and other compounds intermediate between the ore and the final product of the mill. It was generally assumed that the solubility of yellowcake placed it in the category of Class D materials. This plus the fact that the uranium is non-enriched led health physicists to assume that the potential health hazards were a result of the uranium's chemical properties rather than its radiological properties. Bioassay programs have thus been limited to routine urine sampling at a weekly or monthly frequency. In most cases when a high reading was obtained, subsequent sampling within one or two days showed a rapid return to low levels, therefore tending to confirm the class D nature of the material. Two developments over the last few years have however suggested that modifications to this simple bioassay program may be necessary and practical. Firstly, a number of studies on the solubility of uranium mill products, from ore to yellow cake, in simulated lung fluid (Eidson A.F., Mewhinney J.A., 1980) have shown that some of the material is not class D, but class W or even Y. This raised the possibility that, while the urine levels rapidly decreased, indicating rapid excretion of the class D component, the W and Y components remained in the lung. Thus an unsuspected lung burden could have been accumulated. The second development, the ready availability of [in vivo] counting facilities with which to make direct measurements of uranium lung burdens, offered a means of confirming the hypothesis of unsuspected lung burdens. A study was therefore initiated to correlate lung burdens with urinary excretion. METHOD A total of 79 subjects from 3 uranium mills participated in the study. They were brought to the Ottawa laboratories of the Radiation Protection Bureau in groups of 4 or 6, over a period of several months. In Ottawa, [in vivo] thorax measurements were performed, and a urine sample collected. Each subject also brought with him a urine sample that had been collected the previous day. All samples were split, for analysis by both R.P.B. and the mill laboratory. [Concentrations of Uranium Dust] Results obtained by mills A, B and C for concentrations of uranium in air at various locations are given in table 1. Samples taken at mill B were obtained using continuous sampling with a midget impinger over a six hour period. Samples taken at mills A and C were 5-minute gravimetric samples. [Subject Selection] The subjects were selected to represent a wide range of work history in terms of both length of employment and type and location of job, as shown in table 2.

Jan 1, 1981

By J. W. Stevenson

At this 5th U.S. Mine Ventilation Symposium I would like to focus your attention on safety and the contribution of the ventilation engineer over the past 50 years. This conference affords us the opportunity to express, to listen, and to learn. As ventilation engineers it is indeed an exciting time--a time that not only stimulates our imagination but also pin points our aims in our quest to obtain zero fatalities by the year 2000, the goal expressed by William J. Tattersall, Assistant Secretary of Labor, Mine Safety and Health Administration. In the past, coal mine ventilation has often been unfortunately synonymous with explosions and resulting fires. Those not familiar with us do not recognize the painful evolution of the coal mining industry and particularly the evolution of mine ventilation. To put my presentation into perspective, I recently reviewed a U.S. Bureau of Mines Bulletin No. 586 - "Historical Summary of Coal-Mine Explosions in the United States 1810 - 1958 prepared by H. B. Humphrey Mining Health and Safety Engineer. This bulletin was published in 1960. Mr. Humphrey stated "In 1946, factors contributing to the explosions occurring from 1941 to 1945 were: (1) The rapid increase in the use of electrical equipment at face regions where gas and dust are most likely to be found; (2) the increased demand for coal with the desire to obtain maximum tonnage per machine at minimum cost, which in many mines has resulted in neglect of even rudimentary precautionary measures; (3) multiple-shift operation without provision of time for adequate maintenance of equipment, effective rock dusting, and testing for gas; (4) decrease in supervision owing to demands on supervisors for maximum tonnage and loss of competent supervisors to other employment; (5) lack of effective supervision, owing to numerous new conditions, some of them war-connected; (6) inability to obtain adequate equipment repair and replacement parts; (7) shortage of labor and inadequate training of coal-mine workers; and (8) failure of both miners and mining officials to show regard for safety in changing from hand-loading to mechanized methods". Most of the eight factors listed by Mr. Humphrey continue to affect our work lives as mining engineers. The increased demand for coal with the desire to obtain maximum tonnage per machine at minimum cost has never changed. The local and regionalized coal industry of the 40's is today competing against the world coal market. Our mines work or shut down based on cost. This has always been the case. During the five year time-frame Mr. Humphrey was documenting, 345 miners were killed in major explosions. A major explosion was one in which five or more miners were killed in a single accident. In addition, 105 more miners were killed in minor explosions for a total of 450. What a wanton waste of life. What fears, anxieties, miseries, and sufferings were created by these accidents. However, we must remember and keep in mind that the coal mining company officials, engineers, and foremen, the governmental agencies- federal, state, and local - were using the technical tools and knowledge available to operate the mines as safely as they could. Obviously mine ventilation engineering technology was woefully inadequate.

Jan 1, 1991

By C. A. Fleming, K. W. Sarbutt, P. K. MacPhail

The SART process has been developed for Teck to treat bar­ren solution from the cyanidation/CIP process that has been developed for the Lobo-Marte project in Chile. The objective of the process is to recover cyanide in a form that is suitable for recycling to the leaching part of the process. The Marte and Lobo ore bodies require a dosage rate of about 2 kg/t NaCN, of which only 0.5 kg/t is truly consumed (as thiocyanate, cyanate and ferrocyanide species). The remaining -1.5 kg/t is present as cop­per cyanide (-1 kg/t) and free cyanide (-0.5 kg/t), and is poten­tially recoverable and recyclable. SART is an acronym that describes the unit operations in this cyanide recovery process; namely, .. Sulphidization .. Acidification .. Recycling of precipitate " Thickening of precipitate The basic chemistry has been applied previously, and has also been the subject of two patents (American Cyanamid Company, 1965, Potter et al., 1986). The process involves the conversion of all free and WAD cyanide in solution to hydrocyanic acid by acidification of the tailings solution to pH :s,4, and the conversion of all copper in solution to cuprous sulphide (synthetic chal­cocite) by the addition of a sulphide salt:

Jan 1, 1998

By J. Litsenberger, T. E. Marshall, L. Burke, R. Melville, A. T. Iannacchione

Excessive levels of horizontal stresses cause ground fall hazards in underground mines in the Appalachian Basin. At an underground stone mine in Pennsylvania, a modified stress control mine layout is reducing the hazardous conditions associated with excessive horizontal stresses. A microseismic monitoring system is in place to measure levels of rock stability and provide information on the effectiveness of the design technique. The microseismic data is supplemented with frequent and extensive mapping of roof falls and roof rock damage. Findings to date show that the stress control layout provides more stable conditions and thus a safer environment for the mine workers.

Jan 1, 2001

By Robert C. Freas

Over the past twenty years people throughout the world, but particularly in the United States, have been taking an increasing interest in the environment, sources and causes of pollution, and pollution abatement. Air quality in the United States has been and continues to be an area of primary emphasis. In 1970, the US Congress passed the Clean Air Act and subsequently revised and strengthened it in 1977 with the issuance of the New Source Performance Standards. The Act has been further strengthened with the passage of the 1990 Amendments. By far the largest single source of sulfur dioxide (SO,) pollutants are coal fued electric generating power stations. Hence, a substantial portion of the flue gas desulfurization (FGD) research and equipment installation has been geared to the utility industry. Several competing FGD technologies were touted during the period 1970 to 1990 with varying degrees of success. Subsequently, the more successful technologies have been accepted, revised, and improved to the point that by 1991 third and fourth generation technology is available. A second area of emphasis for FGD has been coal fired industrial boilers. While individually they are much smaller than any electric utility generating station, they do represent a significant source of sulfur pollution. However, due to the size of industrial boilers and their location within industrial and educational complexes, the wet scrubber technology being employed by the utilities is not appropriate for these applications. Hence, fluidized bed combustion (FBC) technology has been developed for the industrial boilers. Since the installation of the first full scale test unit at Georgetown University, several equipment vendors have come forward with an increasing array of more sophisticated FBC units. Currently, FBC technology has matured to the point that the Tennessee Valley Authority (TVA), in cooperation with the Electric Power Research Institute (EPRI), has installed an FBC system on a 250 MW test unit at the Shawnee Power Station in Kentucky.

Jan 1, 1994

By Mahmut Yavuz

to in Old Uighur Turkish as Tolay kopugu, in Ottoman Turkish as Derya kopugu, and in modern Turkish as Luletasi. Scientifically known as sepiolite, it is represented by the formula Si12Mg8O30(OH)4(H2O)4(8H2O). Sepiolite is a natural clay mineral composed of hydrous magnesium silicate, characterized by a structure of alternating blocks and tunnels that develop in the direction of the fiber. Each structural block contains two silica tetrahedral sheets surrounding a central layer of magnesium oxide- hydroxide. Sepiolite exists in two distinct forms in nature: alpha-sepiolite and beta-sepiolite. Alpha- sepiolite is amorphous, compact, and resembles pellets that look like sea froth, hence its Turkish name Luletasi. In contrast, beta-sepiolite appears as amorphous aggregates or small, flat, round particles. These two forms exhibit different physical properties. Meerschaum was also studied by Brauner and Preisinger in 1958, who agreed that it primarily consists of sepiolite (2MgO·3SiO2·nH2O). According to these researchers, water is present in its mineral, zeolitic, crystal, and structural forms. Meerschaum is commonly used for adsorbing various liquids, as an ion exchanger, and for dispersing and insulating paraffin in the chemical industry (Gokdemir and Dogan, 2011). Sepiolite deposits of the meerschaum type are primarily found in Turkey, Somalia, Tanzania, Kenya, and Mexico. Alongside high-quality, low-density, and highly porous meerschaum, Somalia also has sedimentary sepiolite formations. The total reserves of these deposits are estimated at around 50–100 million tons (Sariiz and Isik, 1995). Meerschaum, also known as “white gold” or “sepiolite,” is a unique mineral highly prized for its lightness, heat- resistant properties, and ability to be carved into intricate designs. Found predominantly in Eskisehir, Turkiye, meerschaum mining and processing have become significant contributors to the local economy and cultural heritage. This paper explores the history, economic importance, mining practices, and environmental impacts of meerschaum mining in Eskisehir and discusses the challenges and opportunities faced by the industry.

Feb 1, 2025

By Andrzej M. Wala

A knowledge of the friction factor (K-factor) is necessary in planning and designing mine ventilation systems for least cost and highest effectiveness. The friction factor used must be accurate and reliable for correct estimation of airway resistances. At present, the friction factors are determined by using tables of published values. These values were established principally for metal-mine openings having small effective areas. This presents a problem to the ventilation design engineer whose responsibilities center around the design of effective ventilation systems for large room-and- pillar mechanized underground coal mines. Friction factors selected from these tables could be excessively conservative for large scale room-and- pillar operations. With only limited data on the K-factor (which are considerably site dependent) for Kentucky's coal mines and for all the above mentioned reasons, it was concluded that there is a need for field investigations to determine these values in-situ for existing mining conditions in Kentucky. The author describe the study carried out in ten coal mines located in eastern and western Kentucky and present the results of the investigation in both tabular and graphical form.

Jan 1, 1991

By K. Chandrasekaran

Carbon-in-leach (CIL) is conventionally used in gold processing operations when preg robbing components are present in the ore. The CIL process is becoming attractive even in other situations when preg robbing is not a factor. This article discusses the situations in which the CIL circuit can be applied. The results of leach-carbon-in-pulp (CIP) and CIL circuits at Newmont Gold's Mill No. 2, No. 4 and No. 5 are presented for comparison: Mill No. 4 was initially designed for a throughput of 4.1 kt/d (4500 stpd). It is presently operating at an average throughput of 6.6 kt/d (7300 stpd) without significant changes to the original design. The design considerations of the Mill No. 4 CIL circuit are discussed. The design of any processing facility represents a series of decisions. The final design options should ultimately be resolved based on economic as well as technical considerations. Every design decision represents some sort of compromise. From the designer's view point, these decisions are made with the conviction that the final product represents the most prudent long-term expenditure. The design throughput rate is a decision reflecting a maximum return for the capital and operating costs that the plant will incur over the life of the mineral resource. The result is a plant that is not operator proof or maintenance free because it would be too costly to build. At some point, the incremental increase in the level of capital expenditure would not adequately compensate for the lower

Jan 1, 1991

By R. J. Piniazkiewicz

Talc orebodies are mined in Asia, Australia, Europe, North America, and South America. Operators in North America must be efficient quality suppliers to thrive in a talc market in which many overseas suppliers have advantages such as lower hourly wages, shorter distances to ports, and less government regulation and intervention. Continuous improvement through teamwork and innovation is needed in the following areas to thrive amidst worldwide competition: 1) Maximization of the in-place value of ore. 2) Research and development to yield new processes and products. 3) Safe productivity. 4) Involvement in legislative processes. 5) Wellness programs to lower health care costs.

Jan 1, 1993

By G. J. Joy, J. P. Rider

"Technical advances in longwall mining over the last several years have resulted in much larger and faster shearers that have the capability of mining at speeds over 100 ft./min. This has resulted in a substantial increase in the amount of dust that is generated at the shearer and during shield advances. Meanwhile, increased underground coal production has put a greater demand on longwall dust control systems as some operators have had difficulty maintaining consistent compliance with federal dust standards. To address this need, the National Institute for Occupational Safety and Health [NIOSH] is investigating the effectiveness of a water spray manifold mounted on the tailgate end of the shearer body. The sprays are oriented parallel to the tailgate ranging arm, and these sprays create a water curtain confining the dust plume near the face, preventing the dust from drifting into the walkway. This redirection of dust provides a clean air envelope for the tailgate shearer operator and jacksetters working near the tailgate drum. Three different tailgate spray manifold designs were tested and evaluated at the Office of Mine Safety and Health Research [OMSHR] Pittsburgh Longwall Test Facility. Manifolds were equipped with hollow cone or flat fan water spray nozzles and evaluated at three face air velocities (500, 700, and 900 fpm) and at water pressures of 100, 150, and 200 psi. Reductions in dust concentrations were observed at six face sampling locations for all the tailgate manifold systems at each of the tested spray pressures and face velocities. Based on results from the laboratory tests showing reductions in dust concentrations at sampling locations downwind of the shearer a spray manifold was fabricated and evaluated at an underground operation. INTRODUCTION In 2012, underground coal mining in the United States produced a total of 342 million short tons, with longwall mining accounting for over 53% of this production [EIA 2013]. The longwall mining industry continues to see remarkable and significant improvements in longwall mining equipment and mining practices. Average shift production during compliance dust sampling has increased from 3,500 tons per shift in 1994 to approximately 6,000 tons per shift in 2012. A dramatic decrease in working longwall faces from 80 to 49 has occurred over the same period [Fiscor 2013]. Factors specific to longwall mines that supported productivity improvement include: higher capacity shearers, improved shield hydraulics and control, higher horsepower for equipment, wider face conveyors, and larger faces. Today, the average face width has increased to 1,188 ft., with one longwall operation reporting a face width of 1,650 ft. compared to an average of 750 ft. in 1994 [Fiscor 2009, 2014]. Panel lengths in 2013 averaged 11,307 ft. compared to 7,000 ft. in 1994. Overall production from U.S. longwalls has increased by nearly four percent from 2007 to 2012 with over 182 million tons mined annually [EIA 2013]. Given the volume and rate of coal produced on longwall faces, the potential to generate large quantities of respirable coal mine dust also exists."

Jan 1, 2015

By Éder Domingos Oliveira, Carlos Antonio Morais, Ruberlan Gomes Silva

The objective of this work was to assess the influence of parameters such as residence time, solid content, and temperature in a process route towards purified rare earth liquor which could be adequate to produce marketable rare earth products. The steps of the process route proposed in this study comprised water leaching of a sulfated rare earth ore and removal of impurities through the purification of the rare earth sulfuric liquor, which is the product of the water leaching step. The sample used in the water leaching experiments was a sulfated rare earth ore produced by the reaction of a beneficiated rare earth ore with sulfuric acid followed by roasting at 700 °C for 2 h in a muffle furnace and cooling at 20 °C. The water leaching experiments showed that the best condition comprised a residence time of 30 min, temperature of 20 °C, and solid content of 10% w/w, resulting in a recovery of 71.3% of the rare earth elements (REE) and a lower impurity extraction, e.g., only 15.0% of Fe, representing the main impurity in the rare earth ore. Even when presenting low amounts of impurities, the rare earth sulfuric liquor still needed to be purified before it can be used in the production of marketable rare earth products. In the study, the rare earth sulfuric liquors were purified in two consecutive neutralization steps, first by dosing the limestone pulp to pH up to 3.5 and secondly by adding the lime pulp to raise the pH from 3.5 to 5.0. The application of two consecutive purification steps at 20 °C resulted in the highest overall REE recoveries (65.4%) throughout all the steps of the process, being 71.3% in the water leaching, 99.6% in the partial purification with limestone pulp, and 92.1% in the purification with lime pulp at 20 °C. Under this condition, there was a total elimination of Fe, P, and Th and a significant reduction in the Al, U, and sulfate concentrations in the purified rare earth liquor. The purified rare earth liquor attained through the experiment is expected to be amenable to be used in the production of marketable products due to the lower impurities/REO mass ratios.

By J. Diao

In order to understand the flotation kinetic behavior of coal and associated mineral matter, a detailed investigation was conducted in a 2-liter Denver flotation cell using three high sulfur coals at three feed sizes, minus 28. 100 and 200 mesh. Subsequently. the flotation products were sieved into different size fractions and analyzed for their ash and pyritic sulfur content in order to delineate the effect of particle size on the rate at which each constituent floats. The results were tested with several different flotation kinetic models and the best model for coal was then chosen. The flotation rate constant and the ultimate recovery calculated from a kinetic model were used to delineate the effect of particle size, collector (dodecane) and frother (MIBC) dosages on the flotation kinetics of coal, ash and pyrite.

Jan 1, 1992

We have lived through a period of strong metal pricing during the past eight to 10 years. Copper has been hovering in the range of $3-$4/lb during much of the past eight years, averaging close to $3.30/lb. Gold increased steadily from less than $300/oz levels in the early 2000s, reaching almost $1,890/oz in the third quarter of 2011, before tempering back down to the $1,200-1,400/oz range in 2013. Iron ore prices started to rise in 2005 before the staggering increases experienced in 2008- 2010, reaching almost $190/metric ton in early 2011. During the past four years, iron ore prices have ranged from $100-180/t, averaging approximately $140/t. In the aftermath of such a run, it is perhaps too easy to forget where we have come from: The copper price averaged just 77 cents/lb during the six-year period from 1998 to 2003, and averaged just over $1/lb from 1989 to 1997. The gold price averaged $350/oz during the 10-year period from 1994 to 2003. Even more dramatically, iron ore traded in the range of $10-$15/t between 1985 and 2004. So, we have seen sustained, three- to four-fold increases in copper and gold prices, and 8 to 10-fold increases in iron ore prices.

Mar 1, 2014

By Lawrence Neal Post

Introduction Synthetic soda ash production began to decline in the late 1930s. This was due to high operating costs and waste disposal problems. Since soda ash was used as a major component in manufacturing glass, an alternate method of producing that chemical was necessary. A process to produce soda ash from trona was developed by FMC Wyoming Corp.'s predecessor, as well as an economical way to obtain trona. Existing trona operations are located about 40 km (25 miles) west of Green River, WY. Five operators presently mine trona in this area. FMC's property has 153 km2 (59 sq miles) leased. Its eight shafts reach 488 m (1600 ft) deep to the uniform trona deposit. The seam ranges in thickness from 2.4 to 3.6 m (8 to 12 ft). History Egyptian tombs, dating from 4000 to 3000 BC, contained mummies preserved by an embalming process using soda ash. For many years, soda ash was produced by burning marine plants, then dissolving the soluble material (sodium carbonate) from the ashes. The name soda ash is derived from this method of production. In 1791, Nicola LeBlanc developed black ash, or salt cake, from salt, sulfuric acid, coal, and limestone. This process remained in use until 1863, when Ernest Solvay developed a method for producing soda ash from salt, coke, and limestone with ammonia as a catalyst. This Solvay process was widely used. And, in the 1930s, there were 17 plants in the US alone. High plant operating costs and waste disposal problems, however, necessitated finding an alternative to the Solvay process.

Jan 10, 1985

By Philip Morey

Large scale surface coal mining will be required to fill future energy demand in the United States. In realization of this long-term commitment to coal resources, the U.S. Energy Re-search and Development Administration sponsored a study to determine the costs and problems of large-scale surface coal mining. As a result of this study a series of computer models were developed that can be used by government and industry to: 1) facilitate mine design and equipment selection, 2) provide detailed analyses of mining and processing costs. Two types of models were developed: detailed simulation models, which provide detailed information on mining systems and equipment to assist in mine design and equipment selection1 and broad models, which provide a first-order evaluation of deposits and new ventures for an entire mining complex. The complete model series permits examination of surface coal mining under a variety of terrain conditions. Six different mining methods were analyzed and documented during the overall project: 1) Area stripping with draglines, 2) Area stripping with draglines, extended bench, 3) Contour stripping with draglines, 4) Area stripping with shovels (loaders) and trucks, 5) Mountain top removal with shovels (loaders) and trucks, 6) Multiple dipping seam mining with shovels (loaders) and trucks. Each major operation in the coal mining pro-cess is represented in these models so that the best methods and equipment may be selected and 50 that the impact of such selection will be reflected in estimates of capital investment, operating costs, and coal price. Individual models allow the user to separately analyze overburden removal, overburden drilling and blasting, coal drilling and blasting, coal loading and hauling, coal preparation and handling, land reclamation, and premining, facilities, and administrative expenses. Because it is impossible to cover all the models in this paper, only the detailed simulation models representative of the technique "area stripping with draglines," are described. The Fort Union Strippable Coal Resource Region, one of eight test cases, is used as an example to describe the models.

Jan 1, 1978

By A. J. Getz

Major precious metals projects are underway in Australia, Papua New Guinea, Indonesia, the Philippines and New Zealand. This article reviews the status of those projects. Included in this review are project location, reserves, method of resource exploitation, anticipated production, and other unique conditions. The article is limited to major precious metals projects in the Western Pacific region. This is not meant to negate the attractiveness or importance of the many other projects in that area. The geographical setting of these five countries are shown in [Fig. 1]. The gold output of the countries has increased dramatically during the 1980s. [Table 1] compares 1980 and 1988 production figures. The compound growth rate is 120%. That reflects the attractiveness of the region’s resource base as well as the impact of price increases, new exploration models and processing technologies, and changes in government policies. If the 10 projects discussed here are all brought onstream, their combined output - at about 130 t (4.17 million oz) - would amount to an increase of greater than 50% over 1988 production from this region. Table 2 shows indicative production levels for each of the 10 projects. Dollar amounts are in US currency unless otherwise specified. Australian dollars have been converted to US dollars at the rate of 1.28: 1.00.

Jan 1, 1990