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The Bureau has published much data on sulfur (4, 15, 23)3, most of it dealing with national or international aspects of the sulfur industry. Little information, however, is available on sulfur marketing in specific States or areas, Therefore, the Bureau has conducted a survey of the resources, production, consumption, and economic future of the sulfur and sulfuric acid industry in States of the Central and Southern Rocky Mountain area - Wyoming, Colorado, Utah, Arizona, and New Mexico - and in the Northern Plains States - North Dakota, South Dakota, and Nebraska, even though these States do not represent an economic or marketing unit. The purpose of this survey is to review the demand for sulfur and primary sulfur compounds that draws upon the natural resources of the region, to discover opportunities for better use of these resources, and also to analyze and evaluate the patterns of consumption, sources of supply, and other factors that affect the demand for these mineral products. While this survey stresses the economic and marketing phases of the industry, a complementary report (32) emphasizes the resources of sulfur in Wyoming, Colorado, Utah, New Mexico, and Arizona.

Jan 1, 1962

By Robert P. Apa

Energy Incorporated has prepared this final report presenting the results obtained from the program for extracting alumina from anthracite culm along with the recovery of energy. This work has been accomplished under the auspices of the U.S. Bureau of Mines with technical coordination through the bureau's Reno, Nevada facility. The objective of this program was to develop a process to recover alumina and energy from Northeastern Pennsylvania anthracite mining waste. The program included Identification of sufficient resources to support a commercial facility, identification of the operating parameters which maximize alumina and energy recovery, the development of a conceptual process design, preparation of material and energy balance(s), and the performance of a system economical analysis. The program was conducted in three major phases. In the first phase the program concentrated on selecting several anthracite culm banks capable of supporting a commercial alumina producing facility. The chosen banks were then sampled and analyzed leading to the selection of two banks which were chosen far testing. The Harry E. and Eckley culm banks were the ones selected. Material from these banks was tested in Energy Incorporated?s Idaho Falls, Idaho testing facility. Parametric testing was accomplished in an eight-inch-diameter fluidized bed test facility. The test program consisted of twelve separate tests. After critical operating parameters were determined, the second phase started: process development in EI's 18-inch fluidized bed test facility. Three material and energy balance tests were completed in this facility. Following these tests a 72-hour demonstration test was performed. The data acquired from these tests were used in the last phase: the process conceptual design and economic analysis. The design basis was a commercial scale facility capable of producing 500,000 tons of alumina per year for 15 years. Based on the test program, It is technically feasible to build and operate such a plant today. The facility would produce, along with the alumina, 81.5 MWe of electricity for outside consumption. A preliminary rate of return indicates that the plant is not feasible with today's economic conditions. An increase in the market price of either the alumina or the electricity is required to provide an attractive rate of return for the facility.

Jan 1, 1982

By Arnis Mangolds

A study of the technical and economic feasibility of using a unique undercut and fill system developed by USBM engineers to mine steeply pitching coal is discussed. A brief resource assessment has indicated that significant reserves exist in most coal regions to justify the development of a safe and economical method of extracting steeply pitching coal. The undercut and fill system combines hydraulic coal cutting with unassisted gravity fluming of the coal slurry. Mining occurs across strike and the extended entry is backfilled to provide uniform strata control and an impermeable water and air barrier. Subsequent mining occurs beneath the previously filled stope such that the roof consists of a solid, controlled back. In addition to providing positive roof and hanging-wall support, the backfill system eliminates subsidence potential, post-mining drainage, and the danger of spontaneous combustion from naturally drafting air currents. A unique through-flowing ventilation system permits passage of return air through the production face via ducting located at the backfill-coal junction. This provides a constant supply of fresh air at the face, eliminates the need for brattice cloths, and provides an additional escapeway for the miners. The preliminary estimates indicate a production rate in the neighborhood of 31.9 tons/ manshift. Capital costs are approximately $3.17/raw ton over a 20-year life with an initial investment per annual ton of $48.29. Total cash-costs for labor, supplies, power and benefits range in the vicinity of $19.59 for a 704,460 ton/year mine, or a total cost of $22.76/ton. These costs do not include federal and local taxes or post-of-money; however, the cost of coal is well within the area selling price.

Jan 1, 1981

The Federal Coal Mine Health and Safety Act of 1969 limits to 2 milligrams per cubic meter (for dust containing a maximum of 5% quartz) the average respirable dust concentration to which individual miners may be exposed. Section 205 of the Act directs that an appropriate exposure formula for dusts with quartz content in excess of 5% be established. This formula, entered in F. R. Document 70-12351 filed September 16, 1970, is 10 milligrams per cubic meter divided by % quartz. Thus, the total respirable dust exposure permitted limits the quartz to a maximum of 100 micrograms. For example, if % quartz is 5%, 10 mg/5 = 2 milligrams, and (.05) (2 mg) = 100 micrograms. Similarly, if % quartz is 10%, 10/10 = 1, and (.10) (1 mg) = 100 micrograms, etc. If the % quartz is less than 5%, the 2 mg/M3 limit prevails.

Jan 1, 1974

By William A. Hustrulid

An overall summary and the individual reports of eight workshops to determine research and development needs in western mining ore presented. Problem areas are identified, and recommendations regarding priorities, funding sources, and prime movers are made. The objective of this contract was to identify research and development needs for advancing mining technology as well as health and safety measures for western U.S. Surface and underground mining. This was accomplished by convening a series of workshops where problems could be thrashed out by producers of trona, uranium, open pit copper, underground salt (from the Midwest and New York), underground metal and nonmetal, etc. In addition to problems, future priorities, funding, and the desicion as to who should be the prime mover on R & D were discussed.

Jan 1, 1976

By Arno C. Fieldner

Forty years have elapsed since the Bureau of Mines was established on July 1, 1910, by Act of Congress. The convocation of this Sixth International Conference of Safety in Mines Research Directors, 24 days after our 40th birthday, seems to be an appropriate time for reviewing our contributions to mine safety research and the problems yet to be solved from the American point of view. Actually research on safety in coal mines was started by the United States Government in the Technologic Branch of the U. S. Geological Survey of the Department of the Interior in 1908, as a result of a succession of disastrous coal mine explosions in which many lives were lost. In 1910 this work was transferred to the newly created Bureau of Mines. 394 lives were lost annually per 100,000 men employed in the coal mines of the United States in the 5-year period ended in 1910, as compared to 203 lives lost annually in the 5-year period ended with 1949. In 1908 no safe, permissible types of explosives and no certified permissible equipment of any kind were used in American coal mines. Black powder was in universal use and open oil or acetylene lights were common in all but mines known to be gassy. Rock dusting was not practiced to limit propagation of explosions, nor was ventilation in many mines adequate, Our safety practice lagged behind that of foreign coal- mining countries and therefore one of the first steps in carrying out the Congressional Authorization of 1908 to investigate the causes of mine explosions was to obtain assistance from the leaders of mine safety research in foreign countries. At the request of Dr. Joseph H. Holmes, director of the Technologic Branch of the Geological Survey, James R. Garfield, Secretary of the Interior, obtained the cooperation of the Governments of England, Belgium and Germany in sending Captain Arthur Desborough, Inspector of Explosives in England, Mr. Victor Watteyne, Inspector-General of Mines in Belgium, and Mr. Carl Meissner, Counselor for Mines in Germany, to visit American mines and make recommendations for increasing safety and efficiency in mining coal in the United States.

Jun 1, 1950

By Miller, ALEX U, Edward Thomas

In March 1937 the entire country was shocked by news of an explosion of gas in a New London (Tex.) schoolhouse, which resulted in the death of 280 pupils and 14 t eachers. The Government investigators of this disaster believed that the cause was the ignition and explosion of a large volume of explosive natural gas-air mixture that had accumulated in a poorly ventilated space beneath the first floor of the main school building and that this accumulation was due to leakage from gas lin suspended underneath the floor to carry the gas supply to the school.!i/ For some years there has been a potential explosion hazard in school houses and other buildings that are supplied with water from deep wells in certain parts of Illinois. Notwi thstand.ing the disaster at New London, Tex., and the fact that there had been gas ignitions at four water wells in different parts of the State, the explosion hazard in schools and other buildings was not fully recognized until a few months after the disaster in Texas, when an explosion occurred in the settling basin of the municipal water works of Clinton, Ill., and another in a pump house of the municipal water works in Normal, Ill, The explosion in the settling tank of the municipal water works in Clinton, Ill., occurred on August 12, 1937, and a.id considerable damage to the settling tank, but fortunately no one was injured. At the time of this explosion rapairs'and aJ.terations to the water plant were in progress, and 10 men were employed in the immediate vicinity of the settling tank. Two of these men were enlarging a hole in the concrete side with an air hammer for the installation of a larger discharge line.

Aug 1, 1938

By Harold J. Bennett

This report deals with the processing and marketing of iron and steal scrap in the intermountain States of Arizona, Colorado, New Mexico, Utah, and Wyoming and the northwestern Plains States of Nebraska, North Dakota, and South Dakota. The quantity and quality of iron and steel scrap consumed in both the steel mills and iron foundries in these States in 1964 were decided by (1) the type of facilities used to produce an end product; (2) the availability of a dependable source of iron and steel scrap at competitive prices; and (3) the, ratio of the cost of iron in iron ore to iron in scrap. In recent years the quantity of scrap used per ton of steel ingot produced has decreased. The major individual. consumers of iron and steel scrap were the steel mills in Arizona, Colorado, and Utah, foundries in Utah and Arizona, and copper leaching, operations in Arizona, New Mexico, and Utah. About 2 million tons of iron and steel scrap was used by the iron and steel industry in 1964. Although there has been a decline in the amount of scrap used per ton of steel ingot produced because of changes in steelmaking methods, there has been a slight increase in the overall quantity of scrap used because of the increase in ingot production. It is estimated that the scrap consumption in the steel industry will increase by 19 percent by 1980. In addition to the scrap consumed by the steel industry, 273,000 tons of ferrous scrap were used by copper-leaching operations. Thirty-one percent of the ferrous scrap used by the steel industry in 1964 was purchased from dealers and other sources. The CF&I Steel Corp. (formerly The Colorado Fuel and Iron Corp.) at Pueblo, Colo., complemented its blast- and open-hearth-furnace operations with basic-oxygen converters in 1962. The basic-oxygen and open-hearth furnaces were charged with 30 and 90 percent scrap, respectively. Future use of the open-hearths would probably be used to supplement ingot production requirements. Because it was more economical to use less scrap and a high molten-metal charge, the United States Steel Corp. at Geneva, Utah, which uses the open-hearth process to make steel, charges the furnaces with 25 to 35 percent scrap. The Allison Steel Manufacturing Co. at Tempe, Ariz., uses a 100 percent scrap charge in its electric furnace. Although it was estimated that there were 327,000 tons of various grades of motor vehicle scrap generated in 1963, no estimates were made concerning the total quantity and grade of scrap available in the area. Most of the scrap used in the steelmaking industry originated within the area. It was advantageous to sell scrap generated in eastern Nebraska, North Dakota, and South Dakota to markets in the Midwest because of freight costs.

Jan 1, 1967

By Irving Hartmann

The record of safety in the use of permissible explosives in United States coal mines has been excellent, due in great measure to the persistence with which research has been carried on in the Bureau of Mines, to the development of permissible explosives, and to the cooperation of mine operators and workers in using permissibles in a recommended manner and in the effective though gradual replacement of black powder and other nonpermissibles. The demand for increased rate of coal production, changes in mining conditions and minim methods, and introduction of new mining machinery have resulted in modified or new blasting practices. To keep abreast of these developments the Bureau of Mines conducts investigations under actual mining conditions at the Experimental coal mine near Pittsburgh, Pa. Several years ago an extensive study was complete to determine conditions under which the allowable charge limit for permissible explosives per shot hole might be raised from 1.5 to 3 pounds without undue risk of gas or coal-dust ignitions. As a result of the study, this increase in the charge limit was authorized under certain specified conditions in 1948,5/ greatly facilitating production schedules.

Jan 1, 1954

By Harry R. Nicholls

A simple economical method has been developed for determining the in situ longitudinal and shear velocities in rock from which its elastic constants can be calculated. By detonating small charges of high explosives in shallow drill holes near a free surface, both longitudinal and shear waves were generated simultaneously. Accelerometers and velocity gages mounted to the rock surface were used to record the arrival time for both waves over travel path lengths from 50 to 450 feet, Measurements were made in two rock types. From the arrival time measurements, the longitudinal and shear wave velocities were calculated. From these velocities and the density of the rock, the elastic constants were calculated. The dynamic elastic constants of drill core samples of the two rock types were determined in the laboratory by the resonant-frequency method, The lab- oratory and in situ elastic constants are compared and significant differences noted. In general, there is less spread in the in situ data than in the laboratory data, and Poisson's ratio does not fall outside the theoretical limits for the in situ data as is often the case for laboratory-determined elastic constants.

Jan 1, 1961

By Jianmin Qin, Raja V. Ramani, Wilson Miola

"The number of underground work entities operating under reduced dust standards in U.S. coal mines has been increasing since the start of the decade of the 80s and is currently approaching 10,000. The concern for controlling the quartz levels in airborne respirable dust in coal mine has manifested itself in several improvements in equipment selection and operation, face ventilation practices, and dust control procedures. At the same time, fundamental studies have been initiated on determining the source-sink relationships of the airborne quartz dust. The P-7 method is the prescribed method for the determination of quartz in the respirable airborne dust samples from coal mines. Questions have been raised on the reliability of the P-7 infrared spectrometry method for quartz analysis in powders. No specific method has been defined for quartz determination in bulk samples from the mine.In a recent study on the relationship between quartz in the airborne dust and that in the materials being mined, both infrared (IR) and x-ray diffraction (XRD) methods were used on a suite of samples from several coal mines. These samples encompassed airborne respirable dust samples, airborne total dust samples, the oversize fraction from the respirable samplers, ROM samples, channel samples, and chip samples from the face, roof, and floor materials. In this paper, the sample preparation and analysis procedures for the samples are discussed. The data, analysis results, and comparisons of the P-7 and XRD results are presented. The quartz percent determined for the same sample by the two methods are different. In 85% of the samples analyzed, the quartz percents determined by the XRD method are higher than those obtained by the P-7 method. However, there is a strong linear relationship between the results from the two methods. Further, the XRD method indicates an average of 20 percent more quartz than that from the P-7 method. The reasons for the differences in the results and the need for additional studies are explored.Since the enactment of the Federal Coal Mine Health and Safety Act of 1969 and the Federal Occupational Safety and Health Act of 1970, the prevalence of free silica as an occupational hazard in dusty work environments has been a matter of increasing concern. In underground coal mining, sometimes, roof, floor, clay veins and other non-coaly material must also be mined along with the coal. Mining these materials may increase the amount of quartz in the airborne dust (Pimentel and Adam, 1985; Taylor et al., 1984; Colinet et al., 1985). The problem of high quartz levels in the respirable airborne dust is widespread in continuous miner sections. In addition to the need for controlling the silica amount in mine atmosphere, a need for a sensitive, easily performed method for the determination of free silica in airborne dust samples was also recognized."

Mar 1, 1992

By M. E. Hawkins

Mineral analyses and electrical resistivity measurements of 324 representative samples of oilfield brines from 24 oil-productive counties in Mississippi and Alabama are presented in tabular form. Results of the mineral analyses, which include calcium, magnesium, carbonates, bicarbonates, sulfates, chlorides, sodium, and barium (with strontium if present), are reported in milligrams per liter. Electrical resistivity measurements in ohm-meter2 per meter at 75° F are reported in ohm-meters in the tables. Chemical analyses of formation waters are helpful to operators having water problems associated with petroleum recovery; electrical resistivity values are helpful in making interpretations from electric logs.

Jan 1, 1963

By R. V. Ramani, R. A. Jankowski, R. Bhaskar

"The spatial distribution of airborne dusts, deposition of dust in the airway, as well as the effect of airflow velocity and particle density on dust behavior in mine airways were studied in a set of experiments performed in a mine airway under controlled conditions. Considerable deposition of dust near the source was indicated. Pronounced effects due to velocity were observed in ambient dust concentration and deposition along the mine airway, both in the respirable and the total size ranges. Cross-sectional concentration measurements indicate an increase in concentration from the roof to the floor. The average cross-sectional concentration is about 75% of the concentration in the center of the airway.IntroductionA number of studies have been performed since the early 1950s on the behavior of dust in mine air (Bradshaw and Godbert, 1954; Dawes and Slack, 1954; Bradshaw et al., 1954; Hall, 1955-56; Reinhardt, 1971; Ford, 1971; Courtney et al., 1982). While the experimental data available from these studies are indeed useful in understanding dust cloud behavior in a specific situation, no generalized conclusions can be drawn because of the lack of data on parameters such as flow velocity, size distribution of the source dust, and particle density. These change from mine to mine or even within the same mine. The present study is aimed at providing a better understanding of the temporal and spatial behavior of dust particles in underground' mine airways (National Academy of Sciences, 1980). The characteristics studied encompassed airborne concentrations and floor depositions over the entire size range as well as the respirable size range. The experimental results in this paper are intended to complement a mathematical model developed to predict ambient concentration and deposition of airborne dust in mine airways (Bhaskar and Ramani, 1986)."

Aug 31, 1989

By K. Robert Dorman

Spray-applied polyurethane foam was used to insulate the walls of heated rooms excavated in permafrost. The experiment was designed to determine whether polyurethane foam can be sprayed successfully on frozen ground under freezing conditions and to measure the effectiveness of the insulation in maintaining--at elevated temperatures for an extended period--a room excavated in frozen ground. Two manufacturers' foams were investigated in the laboratory. Subsequently, one of the foams was employed to insulate two small rooms, one in frozen gravel and one in frozen silt, in an experimental tunnel belonging to the U.S. Army Cold Regions Research and Engineering Laboratory (USACRREL) near Fairbanks, Alaska. Following insulation, heat was introduced into the two rooms. Heat migration across the insulating barrier into the walls, in addition to wall closures, was monitored over an 8-month period. The results show that polyurethane foam can be successfully sprayed on frozen ground in below-freezing temperatures. Given this type of insulation, together with winter cold to supercool the surrounding ground, it should be possible to maintain a shaft or heated room in permafrost for several years.

Jan 1, 1970

By L. C. IlsLey, R. A. Kearns

As the use of electricity becomes more and more general in mining opera- tions, interest in regulations governing its application becomes greater. During the past year the electrical section of the Bureau of Mines has made a study of State regulations now in force relating to the use of electricity in coal mines. Technical Paper 271, entitled "State Mining Laws on the Use of Electricity in and about Coal Mines", has been published as a result of this study. The present paper represents a compilation of all laws relating to accident prevention with respect to electrical circuits. This group of regulations is perhaps the most important of any that relate to safety in coal mines. Its importance is shown by reference to accident statistics in Bulletin 115, (Fay, A.H., Coal-Mine Fatalities in United States, 1870-1914, Bureau of Mines, 1916), which show that about 77 per cent of the fatalities due to electricity in and about coal mines were caused by direct or indirect contact with trolley or other power circuits. In actual number this means that of 710 men killed by electricity, 546 were killed by contact with electric wires. Undoubtedly, a large percentage of these accidents might have been prevented had the wires been properly guarded. The compilation follows:

Mar 1, 1921

By C. A. Allen

"The problem of signaling to the hoisting engineer from a moving cage in the shaft has been considered by many mine operators for a number of years. There are several methods in use in different mining districts, and so far as the writer is aware they have been further developed in Utah than in any other state. The latest installations are of interest because they use a very low-voltage circuit for the shaft wires and also because the wiring can be so arranged that the hoist can be stopped from the cage, if so desired.The first installation in Utah was in the shaft of the Centennial-Eureka mine under the management of C. E. Allen. It was placed in operation in the summer of 1899, end has been in continuous and satisfactory operation for 21 years. The arrangement consists or two bare wires down the shaft carrying a current at 60 volts potential, with a device on the cage for making connection between the two wires. This device is simply a short piece of wire which is thoroughly insulated from the cage and is operated by a short lever. Whenever the cage rider wishes to signal the engineer he pulls down the lever, forcing the short wire out against the two bare wires, thus making a circuit which rings the bell in the hoist-room. A similar device was planned in the Grand Central shaft at Mammoth, Utah, but was later abandoned. The cause of failure was reported to be that the shaft was very wet."

Jun 1, 1920

By Fun-Den Wang

Finite-element structural analyses were performed to determine the stress distribution and displacements around a single rectangular coal mine opening in a multilayered rock system. The effects of changes in mechanical properties of roof and coal layers, roof layer and mine opening geometry, horizontal-to-vertical load ratio, and structural geologic features on the stress distribution about the opening were studied. Variability in the stiffness and relative thickness of the immediate roof layers can produce a considerable difference in the stress distribution about the opening as compared with that of a homogeneous material. Changes in opening width and horizontal-to-vertical load ratio have different effects on the stress distribution about the opening depending on the composition of the roof layers. In general, where a stiff layer exists in the roof at the surface of the opening, the stresses are higher in the roof and ribs around the opening. Horizontal changes in facies in the roof layers produced by structural geologic features can also greatly alter the stress distribution about the opening compared with the simple layered cases.

Jan 1, 1974

By G. M. Potter, F. S. Wartman, M. D. Schmid

"In matte smelting of copper it has been established that the sulphpr present will combine with copper as long as the latter is present, and that thereafter it will combine with iron. The Cu-Fe ratio in the matte is determined by the Cu-S ratio, presumably owing to the much greater affinity of copper for sulphur or, expressed thermodynamically, the greater stability of copper sulphide in contact with silicate slag. Now, the work of Fournet and of Schütz, which is quoted in many elementary metallurgical texts, indicates that manganese ranks ahead of copper in its affinity for sulphur. Hence, we might expect that manganese could be concentrated from its admixture with iron. Compounds by matte smelting, and Rotts / has patented such a process. This greater affinity of manganese or sulphur is used to advantage in iron metallurgy, also, in which it is well-known that sulphur will combine with manganese and separate from metallic iron as manganese sulphide. Manganese sulphide,however, is quite soluble in silicate slags. 5/ The loss of manganese in slag in any attempted matte smelting is therefore great.If this solubility of manganese sulphide is slag is a physical one and the reaction MnO + FeS MnS + FeOIs substantially complete, then it might be expected that the addition of copper sulphide, which is almost entirety insoluble in slag, would serve to extract the manganese sulphide with the formation of a manganese-copper matte on the other hand, if an equilibrium exists in the above reaction, a copper sulphide layer would displace that equilibrium in one direction or the other, according to whether iron sulphide or manganese sulphide were more soluble in the matter layer."

May 1, 1937

By W. W. Adams

Reports received by the Bureau of Mines, Department of Commerce, from State mine inspectors, for November, 1930, showed that 219 men were killed in the coal mines of the United States during the month, an increase of 25 over the previous month and 33 over November, 1929. On the other hand the production of coal decreased, the output of 43,329,000 tons in November being 8,397,000 tons less than in October and 9,005,000 tons less than in November a year ago. The larger number of fatalities accompanied by a smaller output of coal resulted in a higher death rate4 per million tons in November than in the same month last year or in October, 1930. The November rate was 5.05, as compared with 3.55 for November, 1929 and 3.75 for October, 1930. The increase in the rate for the industry as a whole was chargeable entirely to bituminous mines,, as the anthracite mines had fewer fatalities in November than in either of the previous months to which reference has been made. Reports for bituminous mines alone 'allowed that accidents at class of mines resulted in a death rate of 5.12, based on 195 fatalities and a production of 38,122,000 tons. This rate indicated an increase of 55 per cent over the rate for October, of the same year, and 72 per cent over the rate for November, 1929. The October record was 146 fatalities and 44,150,000 tons, and that for November a year ago was 138 deaths and a production of 46,514,000 tons. Twenty-four deaths occurred at anthracite mines in November, 1970: the output of coal was 5,207,000 tons, hence the rate was 4.61 per million tons. During the corresponding month a year ago there were 48 deaths and 5,820,000 tons of coal produced, which indicated a death rate of 8.25. In October, 1930, the death rate was o.34; based on 48 deaths and 7,576,000 tons.

Jan 1, 1930

By Albert L. Toenges

This circular deals primarily with the various methods of longwall mining used in some mines in certain districts of the Middle Western States. In some instances underground methods were considered comprehensively, and in others only the actual methods of mining and face advance were noted. Results obtained by the mines in output per man were compared only on the basis of an operating day, since idle time and number of men employed would not, be comparable due to the great difference between mines in different districts. Man-hours per ton per operating day are noted in a general way. Material and power costs are not included. ACKNOWLEDGMENT The information used in preparing this paper was obtained through the courtesy and cooperation of the presidents and other officials of the comparnies operating the mines described. The approval of the executive of each property was obtained for publishing this manuscript, and the accuracy was checked.

Jan 1, 1936