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By Richard E. Thill

The Bureau of Mines' Oil Shale Data- Bank (OSDB) was created in 1976 to serve as a centralized source for the storage, retrieval, and dissemination of technical literature, patents, and information on research projects in oil shale. Thousands of bibliographic references, patents, and research projects relating to the mining of oil shale have been identified, abstracted, and indexed according to selected keyword topics. The resulting data bank listings are contained in three sections, each with a separate keyword breakdown for use in rapidly identifying references pertaining to a specific subject. All references cited in the OSDB were published prior to an established cut-off date of May 1, 1981, when a redirection of Bureau of Mines programs resulted in the loss of project funds and the subsequent phasing-out of data bank operations. ACKNOWLEDGMENTS The oil shale data bank listings are the result of a team effort of the project staff. The contributions of Jimaleen Jansma for keypunching, information acquisition, and data entry, and John Brandenburg and Sandra Brake for keywording and information acquisition are gratefully acknowledged. Appreciation is expressed to the staff members of the Smithsonian Science Information Exchange, Washington D.C. for their cooperation in supplying data on recent research projects in oil shale for inclusion in the OSDB listings. Appreciation is extended also to the staff of the Gallegos Research Group for assisting with computer services to obtain the final printouts and transfer of files onto magnetic tape.

Jan 1, 1982

By C. Richard Tinsley

INTRODUCTION The economic analysis (Chapter 3), the engineering studies (Chapter 10), the credit structure (and the consequential funding sources) - Chapter 11, and the overall feasibility structure for any venture are all directed to determining one major decision element: How much and how certain is the reward for taking these risks? Risk assessment is a surprisingly weak area of analysis and is usually addressed in a checklist fashion. It is much preferable to deal within a risk system so that the inevitable tradeoffs among risks are correctly made. DEFINITION The first step toward risk definition must be to ascertain all of the risks to which a project or a company is exposed; otherwise one will remain ignorant of the possibility of a financially crippling loss until it occurs. (Carter and Crockford, 1974). However, simply stating what "risk" is can lead to debate. For example, risk could be (adapted from Vaughan, 1982): -the chance of a loss - uncertainty of loss -possibility of loss - uncertainty -dispersion of actual from expected results - probability of an outcome different from the one expected - possibility of an occurrence of an undesirable contingency -a condition in which a possibility of a loss exists. For the minerals industry financier, the last definition is probably the most appropriate. However, from a mineral company's point of view the definition might well include the possibility of not achieving the expected financial return. RISK CLASSIFICATION Risk identification will be readily agreed as the all-important first step. There are five major "branches" of risk classification. These are: -Statistical -Modellers -Checklist -Project Financing -Insurance The Statistical set discovered probability theory and any risk discussion soon deteriorates into a mathematical jungle (Megill, 1977; Hartman, 1976; Reutlinger, 1970; and Pouliquen, 1970).Their use of terminology tells its own story: coin toss (binomial distribution), gamblers' ruin, delphi (oracle) techniques, monte-carlo simulations, random walks, etcetera. These techniques inadequately represent the judgements necessary in an increasingly complicated and interrelated world. The risk Modellers have turned to the computer to handle these inter-relationships but soon find barriers because of "simultaneity". Certainly the computer has its uses to conduct various analyses (marketed as "sensitivity" analyses) but very often one finds that this involves changing one or two variables at a time, all other things being equal. Unfortunately, in a modern and complex world all other things do not remain equal. The statisticians will, no doubt, jump to their own defence by citing this

Jan 1, 1985

By Harris J. W

The treatment plant at The Great Boulder Proprietary Gold Mines Limited was designed and erected by Chas. Ruwolt Proprietary Limited in 1934 to mill 12,000 short tons per month. The flow sheet called for fine grinding, straking, flotation, roasting of the flotation concentrates and cyanidation of the resulting calcine.In 1937 considerable alterations and additions were made to the plant. The grindin,g section was duplicated with addition of two 8 ft. x 6 ft. Abco Ball Mills to increase the throughput to 28,000 tons per month, and the flow sheet was changed from "straight flotation" to "precyanidation and flotation" neeessitating the addition of thickeners, agitators, Oliver filters and Edwards roasters.A further change to the flow sheet was made in 1942 with the introduction of sulphur dioxide conditioning ahead of the flotation section, to permit the treatment of 500,000 tons of oxidised ore. Since that date no major alterations were made to the flow sheet or equipment until 1950, when the modifications set out in this paper were commenced with the object of reducing costs and increasing plant efficiency.These modifications are dealt with under the section headings and are therefore not necessarily in chronological order.CRUSHING SECTIONThere has been no material alteration in this section although at the time of writing a 650 ton capacity ore bin is being erected ahead of the primary Jaques crusher. This bin will reduc'e the crushing time from a three shift to a two shift basis and reduce the labour force required by two men, giving an annual saving of £2,000.FINE GRINDING SECTIONAlterations to the grinding section were commenced in 1950. Prior to this date the grinding flow sheet calle...

Jan 1, 1955

By Sam S. Taylor

Subsurface disposal of oil-field brines was practiced in Oklahoma on a limited scale as early as 1925. The technique of subsurface injection has been improved immensely and has been applied with varying success in many oil-producing areas in Oklahoma, and elsewhere. This method of disposal, however, is only one of several methods available to operators of oil-producing properties, but, in contrast with usual methods of handling brines on the surface it eliminates the damaging effect of mineralized water on fresh-water supplies, vegetation, and aquatic and other animal life. Other methods of brine disposal approved by regulatory agencies in some States include (1) impounding in impervious ponds and solar evaporation, (2) controlled diversion into surface waters, and (3) recovery of the mineral salts of the brine. In most, States where oil is produced regulatory bodies authorize the use of one or more of these methods of brine disposal. In Oklahoma, for example, the Oklahoma Planning and Resources Board of the Conservation Commission has been authorized to adopt whatever methods or means may be necessary to prevent contamination by mineralization of all potable waters within the State.

Jan 1, 1942

Jan 1, 1885

By J. E. Campbell, A. W. Wesner, M. Pobereskin

This paper summarizes the results of an investigation sponsored by The Coates Steel Products Company. The objective was to determine the relative effects of certain chemical and physical properties on the wear resistance of grinding balls when grinding cement. In making this study, a number of types of balls were used in both dry and wet grinding in commercial ball mills. The radioactive-tracer technique was used in separating the test balls from the rest of the mill charge and in sorting the different types of halls that were all tested at the same time for each of the grinding operations. PROCESSING OF GRINDING BALLS In order to determine whether the properties of grinding balls, namely, composition, microstructure and hardness, can be related to the rate of wear, the halls used in this study were selected to represent various combinations of these properties. It was recognized that the mechanism of wear was not the same in the wet mill as in the dry mill. Therefore, the types of balls tested in each of the two environments were not necessarily the same. However, all of the types of balls evaluated in the dry mill were tested simultaneously for one time period, while those evaluated in the wet mill were tested simultaneously for three different time periods. All of the test balls were separated from the charge when the mill was dumped. All of the balls were of 0.940-inch diameter when charged, and all but the regular production balls were finish ground to this size to eliminate the decarburized surface layers. The steels and processing methods used, as well as the micro-structures and hardness levels of the balls, are described as follows: Regular production forged SAE 1095 steel balls were used in both the wet and dry mills. They were forged to a nominal diameter of 0.940 inch and quenched in oil from the forging temperature. Each retained a partially decarburized surface layer from the forging bar stock; the maximum depth of this layer was about 0.030 inch and the decarburization was negligible where the ball was sheared from the bar. The hardness at the surface was 40 to 46 Rock-well C in the partially decarburized area and 57 to 62 Rockwell C in the areas not decarburized. The hardened zone extended to a depth of about one-third of the radius. Below the decarburized layer, the hardened zone consisted primarily of hard as-quenched martensite with varying amounts of retained austenite. This structure changed to fine pearlite at a depth of about one-third of the radius. In order to eliminate the effect of the decarburized layer, a set of production balls of SAE 1095 steel was forged to about a 1-inch diameter, quenched in oil from the forging temperature, and ground on ball-bearing grinding equipment to a 0.940-inch diameter to remove the decarburized layer. The balls' hardened zones were similar to those in the previous set. The hardness of the finished surfaces was 52 to 60 Rockwell C. A slight amount of decarburization apparently remained after the removal of about 0.030 inch from the surface. Balls of this type were tested in the dry mill only. Another set of forged SAE 1095 steel balls of 1-inch diameter was heated to 1500 F and quenched

Jan 1, 1961

By Daniel W. H. Su

A nearly 305-m- (1,000-ft-) wide sandstone channel was projected to run through the mid panel area of a 457-m- (1,500-ft-) wide longwall panel in a southwestern Pennsylvania coal mine. This sandstone channel was known to cause serious longwallface roof control problems in the mine?s B-panel area in 1998 and 1999; and most recently, it induced many large face cavities and caused a substantial production delay in the preceding panel. For the two B panels mined in 1998 and 1999, hydraulic fracturing of the massive sandstone was conducted to facilitate better caving, thus relieving cantilever compressive stress at the longwall face. Vastly improved longwall face condition was observed and four weeks of mining delay due to bad roof conditions were eliminated. The implementation of the hydraulic fracturing technique, results of underground observation and measurement, and results of numerical modeling were summarized in a paper presented in the 20th International Conference on Ground Control (Su et al., 2001). This article is a summary of the activities that took place during a four-month period (October 2011 through February 2012) at a Consol Energy longwall mine southwestern Pennsylvania to improve face conditions and safety of underground personnel during mining of the 457-m- (1,500-ft-) wide longwall panel. The elevated concern for safety was driven by adverse face conditions encountered while mining the preceding longwall panel. Many of the problems were attributed to a thick and massive sandstone channel. A joint task force consisting of coal operations engineering, Pennsylvania operations engineering and CNX Gas personnel was formed in August 2011, which recommended drilling four hydraulic fracturing holes along the center axis of the sandstone channel to reduce pressure and improve face conditions at the longwall face. To further provide detailed monitoring of the longwall face and to evaluate the effectiveness of the hydraulic fracturing program, the Longwall Visual Analysis (LVA) software (Hoyer, 2011) was installed in early December 2011 to track the face pressure and c

Jun 1, 2013

By G. P. Halliwell, C. R. Austin

The investigation described in this paper deals with the development of high-temperature alloys of the Konel series over a considerable period of time at t,he Research Laboratories of the Westinghouse Electric and Manufacturing Co. The word ''Konel" has been coined to designate a class of alloys which are essentially nickel-cobalt-iron with other elements added, usually in small quantities. Most of these alloys contain titanium, usually about 2.5 per cent. The main consideration related to strength at high temperatures but the thermionic characteristics of the alloys was also of considerable importance. It is, however, with the mechanical properties at elevated temperatures that we are herein concerned. Much attention has also been devoted to their age-hardening characteristics. There is little need to stress the utilitarian aspect of this subject, as it has been discussed frequently. Alloys exhibiting superior properties at normal or elevated temperatures rapidly find many suitable applications. There may be some question, however, as to what constitutes a "superior alloy," or an alloy that may have good service possibilities at high temperatures. The problem is complex and various means have been employed to predict the useful life that may be expected of a material. In recent years the "creep limitJ' has been used as the best means of evaluation. Such tests necessarily are prolonged and hence collection of data is very slow. In a general survey of many alloys therefore the creep test is not desirable and recourse must be had to much shorter time tests in order more rapidly to select alloys which appear to show promise. For this reason the present survey has been made using the high-temperature tensile test and a bend test developed by Howard Scott.' In the tensile test the conditions have been standardized for all tests, so that values obtained are at least comparable, one alloy with another, although the absolute values recorded merely indicate the mechanical properties at the given temperature under a definite rate of loading.

Jan 1, 1932

By A. Kostin

GIS and multivariate geometric distribution have been used to predict the association of high-level intrusives with hydrothermal alteration related to iron oxide Cu-Au (IOCG) mineralizing systems. Several examples in the Russian northeast are presented in this paper. IOCG ore deposits can have enormous geological resources containing significant reserves of base, precious, and strategic metals, and hence are economically attractive targets for mineral exploration. To date, examples of this type of mineralization have not been reported from the northeastern area of Russia. The Tarinskiy ore cluster, located in Eastern Yakutia, shows brecciated altered rocks with sulphide and iron oxide cement, which is typical for IOCG mineralization of the iron oxide Cu-Au±U type. This cluster was formed near surface and is linked with the Rep- Yuruinskiy pluton. It has potential to host new world-class precious metal deposits in northeastern Russia

Jan 1, 2014

By R. L. Grayson, G. Mishra

Introduction Many old underground mines have been developed extensively into reserves requiring large numbers of workers to maintain and rehabilitate lengthy track haulages and/or belt conveyor systems. Often, large numbers of fans are required to maintain minimal airflows in areas where poor conditions and huge leakages prevail. Problems that beset one such coal mine and projects undertaken to restore it to potential profitability are presented. Brief history The Vesta No. 5 mine was opened in 1908. Drift development began near the Monongahela River, which was the primary avenue for shipping metallurgical coal to coke ovens in Pittsburgh. Over the ensuing 75 years, the main haulage entries were developed a distance of 16.1 km (10 miles) to the present mining area. Track haulage was used to transport coal from working sections to the river loading facility. As haulage distance and areas of high water influx increased, management decided, in 1973, to construct an underground conveyor belt system. The project spanned 36 months and culminated in a system of 22 flights, ranging from 205.7 to 1767.8 m (675 to 5800 ft) long. Six intake and eight return entries were developed as the mains were advanced. These entries provided adequate ventilating capacity over the years, but as extensive gob areas progressed and roof conditions deteriorated, the entries became inaccessible and more resistant to airflow. By 1981, five fans were being used to circulate 943.9 m3/s (555.5 cfm) of air through the mine. An average of 6440 t (7100 st) were produced from 20.8 continuous miner unit shifts and 2.7 longwall unit shifts each day in 1981. The ratio of support to production personnel was 1.94. To maintain the outby 16.1 km (10 miles) of haulage, 110 personnel were used. The average cost per ton was $43.54 ($39.58 per st), resulting in excessive bottom-line losses. The mine possessed a rather promising reserve, mostly in lease, of good quality, medium-sulfur metallurgical coal, but it also had a number of formidable operating problems. Operating problems Contributing factors toward relatively low productivity and high costs were the conditions of the track haulage, conveyor, and ventilation systems. Productivity was affected greatly by downtime associated with the conveyor system, which lacked adequate access for making splices or other repairs. Many areas within the conveyor belt entry itself required more

Jan 3, 1987

By P K. Banerjee, S Kumar

Indian iron ores with less than 62 per cent Fe are currently unexploited, except for a few sectors which export them to foreign countries. It is anticipated that by 2012 the countryÆs domestic steel production is likely to rise to the tune of 60 million tonnes per annum. To meet this requirement, it is necessary to utilise not only the existing high-grade deposits but it is also necessary to look into exploiting the leftover iron ores of lower grades. Hence, reserves with Fe content 62 per cent and below (particularly the ores with In the present paper, an attempt has been made to beneficiate three different types of low-grade iron ores selected from three different parts (states) of the country. In the paper, alternative routes for beneficiation and effective utilisation have been tested using different techniques in the laboratory in an attempt to match their (Al2O3/SiO2) ratio after thorough characterisation and liberation studies. The paper also encompasses the geological formations, occurrences and distribution of Indian iron ores as well as highlighting the problems associated with the fines, possible metallurgical routes, and their market segregation from the perspective of end users.

Jan 1, 2007

By Linda C. M. Elliot, Liangxue Liu

"The deposition of tailings under water is considered to be one of the most environmentally promising and potentially cost effective methods to reduce or eliminate acid mine drainage from tailings. One of the concerns of subaqueous deposition has been that metals may leach into the cover waters, thereby increasing aqueous metal concentrations to higher than acceptable limits. This potential was seen to have an even greater possibility for impact when the tailings are deposited under acidic lake water.The following paper summarizes the results of subaqueous tailings pilot column test work conducted to determine the effects of subaqueously deposited tailings on lake water and tailings pore water. Two types of tailings were tested: high sulphide tailings (pyrrhotite concentrate), and low sulphide tailings (end of pipe slime tailings).Data was collected from five columns over a 13 month test period. The results indicated that a stagnant I 000 mm water cover was capable of maintaining the underlying tailings in a reducing condition. The upwards diffusion of pore water resulted in a pH increase in the waters overlying the thickened pyrrhotite tailings and the pyrrhotite tailings with a shallow cover of substrate. A conductivity gradient between the tailings and the water cover caused a diffusion of salts upwards into the overlying water column. Precipitation noted at the tailings water interface may be related to the precipitation of ferric oxyhydroxides, which resulted in a decrease in Fe concentrations noted in the water cover. Other metal concentrations did not exhibit significant increases in the water cover. A sulphate concentration gradient between the tailings and the water cover may have caused the diffusion of SQ4 from the tailings into the water cover."

Jan 1, 1997

By Donald R. Johnson

Apparatus providing simultaneous differential thermal analysis, thermo-gravimetry, and evolved gas analysis, has been developed by the Bureau of Mines specifically for thermal study of oil shale and other solid, natural fuels. The apparatus is described in detail, and simultaneous analyses on a single sample of an oil shale mineral, nahcolite, and a thermal analysis demonstration compound, calcium oxalate monohydrate, are presented to demonstrate the instrument's capabilities.

Jan 1, 1970

By Arndt C. R

The Ispat Mexicana SA de CV (IMEXSA) pelletising plant is located at Lazaro Cardenas on the west coast of Mexico. It is the primary source of feed stock for the adjacent direct reduction (DR) plants and steelmaking facilities. The initial design of the pellet plant began in 1981 for 3 Mtpa of DR grade pellets using local concentrate received by pipeline.' Several construction delays, an ownership change, and a decision to use imported concentrates delivered in 60 000 to 90 000 tonnes ship cargoes required flowsheet modifications and additions. The design concepts are described for the plant which includes concentrate regrind to final pelletising particle size.

Jan 1, 1999

Mount Read is 'situated in the county of Montagu, on 'the west coast of Tasmania, about 20 miles inland from. Remine in an easterly direction, 15 miles northerly from Mount. Lyell, and 30 miles south-east from Mount Bischoff, and is one of the mountains forming the west coast range; it rises to a height of 3890 ft. above sea level.About eleven years ago the first mineral section was taken up on Mount Read by a Hobart company, known as the Mount Reid Silver Mining Company, consisting of 25,000 shares its holding consisted of one section of 80 acres, held under mineral lease. Prospecting was carried on for some months under great disadvantages, as all supplies had to be carried up the mount from Dundas, a distance of nine miles, on men's backs, until a pack track for horses was made.The track rises 2500 ft. in the first five miles, and then falls 200 ft. from that point to the mines, which are at an altitude of 3000 ft. above sea level, and at a point about two miles westward from the summit of the mountain, approximately midway between Mounts Read and Dundas.

Jan 1, 1901

By J. Cygankiewicz

Efficient prevention of mine fires is impossible without predicting the self heating processes in the given geological and mining conditions. The respective forecasting method is presented in this paper based on the following assumptions: - factor deciding on fire hazard in the underground workings is the process of heating of coal from the primary rock temperature to the critical temperature when the heating rate growth rapidly - another important factor is the fire incubation time defined as time needed to achieve the critical temperature. This time is determined by balance of heat and mass in the heated coal - contact of crushed coal with oxygen in air cause some exothermic processes as : physical and chemical sorption of oxygen , oxidation reaction and evaporation of humidity Heated coal transfers its thermal energy to the environment by conductivity to the rocks and by convection to the flowing gases. The heating is possible if thermal energy generated from coal is greater then the share transferred to the environment. The paper presents the laboratory method of measurement of heat generation in adiabatic conditions. Also the mathematical model with examples of computation enabling to forecast the self heating process of crushed coal is presented.

Jan 1, 2003

By Robert S. Shank

Introduction The Pinto Valley open-pit copper mine is located in southern Arizona, 11 km (7 miles) west of Miami. Production at Pinto Valley began in 1974. About 350 Mt (386 million st) of ore and waste was removed by the end of June 1982. At that time, the mine was using more than 90 major equipment items with a total mobile fleet numbering 320. Some of the major equipment consisted of four Marion M4 rotary drills; six P&H 15, 11.5- and 19.1-m3 (15- and 25-cu yd) shovels; and 24,136- and 154-t (150¬and 170-st) Wabco haulage trucks. This equipment was used to move more than 139 kt/d (154,000 stpd) of ore and waste. The mill was processing more than 50 kt/d (55,000 stpd) of mine ore. Shutdown A two-week vacation shutdown began in July 1982. Maintenance jobs were scheduled, and most employees left for the vacation. During the vacation period, copper prices fell below $1.32/kg ($0.60 per lb). Inventories remained at alarmingly high levels. At this point the mine was losing more money operating than it would by being shut down. The vacation shutdown was first extended to five weeks. But, after further evaluation, the mine was shut down indefinitely. In Dec. 1982, Cities Services Co., owner of Miami Operations (which included the Pinto Valley mine and mill), was acquired by the Occidental Petroleum Corp. On March 1, 1983, Occidental sold the operations to Newmont Mining Corp., and Newmont created Pinto Valley Copper Corp., as a wholly owned subsidiary. Copper continued to be produced by the Pinto Valley solvent extraction-electrowinning (SX- EW) plant because of its low cost per pound of production. To ensure proper equipment care during the shutdown, maintenance management was given three goals; • maintain the SX-EW plant; • maintain the environmental and pollution control systems such as domestic water pumps and mine dewatering; and

Jan 4, 1985

By R. G. Kidel, S. C. Doming

The project has shown that it is feasible for a junior company to reopen a narrow vein-style gold mine and successfully take it into production. An intimate understanding of the mineralisation and a flexible approach to mining are crucial, as are stringent cost control and efforts to keep overheads low. The project has confirmed the geological complexity of the Main Lodes, which consist of highly sheared quartz-sulphide infill in strongly sericitised granodiorite. The existence of discrete high-grade zones within the Hangingwall Lodes has also been confirmed. Broad low-grade zones continue to be recognised and require the local application of bulk longhole mining methods. The initial mine plan called for predominantly longhole stoping, but adverse ground conditions have meant that a switch to air-leg mining has produced higher run-of-mine grades. The excess capacity at the processing plant can best be used by batch processing, with its reduced maintenance requirements

Jun 21, 1905

By Herbert H. Kellogg

The chemistry of sulfide roasting is analyzed to show those aspects of performance which Thecan be predicted from considerations of thermodynamic equilibrium. It is concluded that equilibrium calculations can serve as a yardstick for prediction of roaster-gas composition and sulfate formation in the calcine for fluid-bed roasting practice. A method of calculation which combines the equilibrium and stoichiometric factors is outlined, and illustrated by an example from the roasting of Cu-Fe sulfides. ON first examination the roasting of sulfides would not seem to be a process to which equilibrium calculations could be applied to yield results of practical value. The burning of sulfides is an ex-tremely spontaneous reaction, excess air is always used to obtain rapid reaction, and in most traditional roaster designs (hearth roasting) the temperature and gas composition vary widely from point to point in the roaster. These are all conditions which promote a non-equilibrium state in the roasting process. However, the introduction of the fluid-bed roaster1 has brought to roasting practice new conditions of operation, so that equilibrium calculations can be used to predict certain limited aspects of practical operation. The fluid bed is characterized by uniformity of temperature, gas composition, and calcine composition from point to point throughout the bed. These relatively uniform properties of the bed make the meaningful calculation of the equilibrium state of certain reactions in roasting possible. It is the purpose of this paper to analyze roasting from an equilibrium viewpoint, to show where the method may be expected to yield valuable information, where it fails, and to outline methods of calculation. Main Roasting Reaction The reaction MS (s) + 3/2 0, = MO (s) + SO, [I] might well be called the main roasting reaction. It adequately accounts for the well known facts that metal sulfide and air are the main reactants, and metal oxide and SO, the main products. Deviations from this simple stoichiometry will be discussed under side reactions. For all metal sulfides, Reaction 1 is strongly exothermic and the equilibrium position is far to the right. Since the reaction is exothermic, the equilibrium shifts in the right-to-left direction as the temperature is raised. However, at all practical roaster temperatures (between 500" and 1200°C), the amount of this shift is small, and the equilibrium position remains far to the right. The main roasting reaction is essentially irreversible, and the application of equilibrium considerations to this reaction will not be! a fruitful field of investigation. A knowledge of the rate of this reaction. although not the proper subject of this paper, is of great importance to the roasting process. Analysis of the factors which influence this rate shows that: 1) The rate will increase rapidly with increased temperature. 2) The rate will be proportional to the partial pressure of oxygen at the surface of the reacting particle. Because of diffusion effects, the partial pressure of oxygen at the particle surface may be considerably less than in the bulk of the roaster gas. For this reason, intimate contact between gas and solid and high relative velocity between gas and solid, such as can be obtained in a fluid bed or flash roaster, will markedly increase the rate of reaction. 3) The rate will be proportional to the surface area of metal sulfide. Therefore, smaller particle size will permit a more rapid rate per unit of weight. To achieve practical rates of roasting the partial pressure of oxygen in the roaster gases must be maintained at a finite level at all times. Even where the gas-solid contact is very good, as in the fluid-bed roaster, the oxygen content of the gas must not be allowed to drop below some limiting value (perhaps 1 pct), or the rate of the main roasting reaction will become too small for practical operation. Side Reactions in the Gas Phase The gas phase during roasting of metallic sulfides contains the elements oxygen, sulfur, nitrogen, by-

Jan 1, 1957

By Robert J. Rothwell

Underground mining of coal by longwall mining methods has not been a predominant technique in Ohio. To date, only five mines in the state utilize this method, however, the technique has sparked a certain amount of controversy. As a mining technique, longwalling was introduced into southeastern Ohio in the early 1970's but it did not attract public attention until the mines began moving into privately held surface lands. This occurred coincidentally at about the same time that the state of Ohio began to regulate the environmental impacts of underground mining. Prior to the passage of the Surface Mining Control and Reclamation Act in 1977, (SMCRA), Ohio's reclamation law, Chapter 1513 Ohio Revised Code (O.R.C. 1513) did not regulate underground coal mines. As a result of SMCRA, the Ohio Statute was amended in 1978, to include regulation of the surface effects of subsidence. The inevitable result was that existing mines were immediately subject to regulations that were not considered during the initial mine planning process. (For an excellent survey of Ohio Laws that deal with coal mine subsidence, see Zima, 1985. ) The problem faced at that time by both the coal industry and the Ohio Division of Reclamation (the agency within ODNR charged with implementing O.R.C. Ch. 1513.1, was that of attempting to superimpose permitting requirements, and design performance standards onto mines that' had been operational for a number of years.

Jan 1, 1986