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By D. W. Bridges, Donald R. Peterson, Masami Hayashi

Techniques are described for recovering about 80 pct of the cobalt and nickel contained in superalloy scrap by anodic dissolution of the scrap and simultaneous electrodeposition of a pure cobalt - nickel alloy at the cathodes in an electrolytic cell partitioned by permeable diaphragms into separate anode and cathode compartments . Superalloy scrap was alternatively prepared for subsequent anodic dissolution either by melting and casting massive scrap into suitably sized anodes , by packing moderate - sized discard scrap in inert metal baskets that subsequently served as anodes , or after concentrating the superalloy portion of grinding debris , by magnetic separation to reject corundum or other abrasives , slagging any remaining abrasives from molten grindings , and finally casting the superalloy metal into anodes . In the processing , purified electrolyte -- 1 N in NiSO4 and in NiCl₂ , and 0.1 N in CoSO4 and in CoCl2 --was charged to the cathode compartments and discharged from the anode compartments at a rate sufficient to overcome the backward migration of ionic impurities from the anodes to the cathodes . Discharged electrolyte was purified by reacting it first with SO2 and next with limestone , and then recirculated to the cathode compartments , in which a 50 pct cobalt - 50 pct nickel alloy was deposited on either blanks of stainless steel or starting sheets of cobalt - nickel alloy . The process was made self - sustaining by either adding H2SO4 to the catholyte feed to throttle the rate of cobalt - nickel deposition , thereby compensating for the slower rate of cobalt - nickel dissolution from the anode and losses occurring in electrolyte purification , or alternatively , adding H₂SO4 to

Jan 1, 1964

By H. Kenworthy

The investigations described in this report were undertaken to demonstrate, on a laboratory scale, the possibility of recovering several million pounds of tin plus tungsten contained in stockpiled and currently rejected slags from the Longhorn tin smelter at Texas City, Tex. At the same time, it was anticipated that the treatment procedures developed might be applicable to slags from other tin-smelting operations. Results of the laboratory work show that it is possible, by pyrometallurgical means, to separate the tin and tungsten into products of intermediate grade that may be returned to the smelter circuits for further upgrading. Several methods of treatment were investigated. The methods that proved effective were: Fusion reduction, chloride volatilization, sulfide volatilization, and carbide formation. Over 90 percent of the tin and tungsten could be removed from the slat by any of these methods. Each type of treatment has advantages and disadvantages, and it could not be determined from the small-scale work which method was the more meritorious. Tungsten and (in some types of treatment) tin-bearing products were diluted by iron, the separation of which will require additional studies. Prevention of a portion of the iron from reporting with the tin or tungsten is a distinct advantage and possibility that will require inure laboratory work to develop the needed techniques for its accomplishment. Tin products from single treatments of stag ranged up to 75 percent in tin content but were more generally under 35 percent. Tungsten products were usually under 15 percent; but segregated portions could be selected, at a sacrifice in recovery, that contained up to 50 percent tungstic oxide equivalent.

Jan 1, 1957

By Donald W. Ross

Petroleum oils of widely different grades have long been clarified or decolorized by contact with fuller's earths. More recently, so-called activated clays have also been used for this purpose. Still more recently ouch clays have been used in the form of nodules termed "catalytic pellets." Clays of varied appearance and from different localities possess the desired bleaching qualities, but evaluation of such clays, so useful to the oil industry, has not been simple. Previous methods of identification and evaluation have largely followed procedures of production methods and were not primarily based upon the fundamental properties of the clays. Thus) a false assumption has grown up that) bleaching clays probably have no common properties on which a-method of comparison can be based. Fuller's earths and activated clays have a large absorptive capacity for the color constituents of petroleum and its products decolorizing is accomplished by bringing the oil into intimate contact with the clay and later separating the decolorized oil from the clay with its load of absorbed coloring matter. Former practice in absorbing "color" was to use the clay as granules or In' powdered form. More recent practice is to use clay pellets that are more abrasion-resistant than the granules. The oil is either filtered through a bed of clay, percolation method, or agitated on contact with the clay, contact method. Viscous oils are decolorized at an elevated temperature, so that their viscosity is lowered to the point at which thorough contact with the clay is insured. Nonhumidifying, nonoxidizing gas, such as dry steam or dry nitrogen, is used to seal such hot oil masses from the atmosphere.

Jan 1, 1948

By L. A. Haas

The U.S. Department of the Interior, Bureau of Mines, is conducting laboratory-scale research to determine the leaching conditions required with sulfuric acid for extracting base metal from a matte with a copper-to-nicke1 ratio of 6 to 1. The matte (61.2 pct Cu, 9.7 pct Ni, 3.3 pct Fe, 21 pct S, and about 5 oz of precious metals per ton) was prepared from Duluth Complex (Minnesota) ore by smelting a bulk sulfide flotation concentrate. The matte was pre leached at atmospheric pressure to selectively extract 20 pct of the nickel, and the undissolved solids were referred to as the matte residue. Autoclave leaching of the matte residue extracted over 95 pct of the nickel, copper, and iron. The remaining leach residue in the autoclave contained over 100 oz of precious metals per ton. The leaching conditions required were matte grind of greater than 98 pct minus 325 mesh, pulp concentration of 50 g/1, total pressure of 100 psig with oxygen at 150° C or 400 psig with air at 180° C, 50 g/1 H2S04, 700 rpm agitator speed, and 6 hours of leaching time. The leaching time was decreased to 2 hours by increasing the agitator speed to 1,300 rpm. Sulfuric acid concentration greater than 50 g/1 did not appreciably enhance the base metal extraction, but with less acid, residues with higher iron concentrations were obtained. Supersaturation of the liquor was prevented at 150° C by using a 1ixiviant that contained 30 g/l Ni and 20 g/1 Cu.

Jan 1, 1979

By A. E. Back

A method has been developed in the laboratory for successfully treating titaniferous magnetite ore from the deposit at Iron Mountain, Wyo. The ore is roasted with 15 percent sodium carbonate to convert the vanadium to a water-soluble form and fix the soda for the subsequent electric-smelting step. Approximately 90 percent of the vanadium is recoverable in a product assaying 80 percent V2O5. The leached calcine is briquetted and smelted in a graphite-lined electric-arc furnace. Substantially all of the titanium is recovered in a soda slag, assaying 60 percent TiO2 and 2 percent Fe, which can be upgraded to 75 to 80 percent TiO2 by leaching with dilute sulfuric acid. The upgraded slag could be utilized for the production of titanium dioxide pigments or titanium tetrachloride. About 90 percent of the iron is recovered as pig iron, containing only minor amounts of vanadium, titanium, and sulfur. The high TiO2 slag compares favorably with slag containing about 70 percent TiO2 being produced by the Quebec Iron & Titanium Corp. at Sorel, Quebec. The data obtained from this investigation indicate that a feasible process has been developed for treating Iron Mountain ore, and larger-scale testing is warranted to develop engineering data for the ultimate utilization of this deposit as a source of titanium, iron, and vanadium. There are adequate supplies of basic raw materials, electric power, water, sodium carbonate, coal, and petroleum products within fairly short distances of Cheyenne and Laramie, Wyo., the nearest cities and probable plant- and labor-supply points.

Jan 1, 1952

By James J. Dowd

The investigation to evaluate the reserves of coking coal in the United States suitable for the manufacture of metallurgical coke is being made by the Bureau of Mines in three parts: (1) To estimate known (measured plus indicated) recoverable reserves of all coking coal; (2) to study the carbonizing properties of coals and coal blends not now widely used for metallurgical coke making; and (3) to upgrade marginal coals through effective preparation. This is the second7j of a series of reports giving the results of studies by counties under part (1) of the investigation. This report covers Indiana County, Pa., which comprises parts of the Smicksburg, Punxsutawney, Elders Ridge, Indiana, Barnesboro, Latrobe, New Florence, and Johnstown Quadrangles. (See fig. 1.) A base map for each bed in each Quadrangle was made to the scale 1 inch equals 1,200 feet. Maps of mines, locations of drill holes, bed and total coal thicknesses, and the outcrop of the bed were plotted on the maps. With all available data plotted, isopach lines were drawn to limit areas of known unmined reserves in beds 0 to 14 inches thick, 14 to 28 inches thick, 28 to 42 inches thick, and over 42 inches thick. These areas of coal reserves also were divided into "measured" and "indicated" categories. All areas in each thickness range and in each category, mined-out areas, areas excluded from the estimate, and areas outside of the outcrop were measured by planimeter on the base maps. Estimates of total reserves 14 inches and more thick and maps for individual beds were prepared from these data.

Jan 1, 1950

By James J. Dowd

The investigation to evaluate the reserves of coking coal in the United States suitable for the manufacture of metallurgical coke is being made by the Bureau of Mines in three parts: (1) To estimate known (measured plus indicated) recoverable reserves of all coking coals; (2) to upgrade marginal coals through effective preparation; and (3) to study the carbonizing properties of coals and coal blends not now widely used for metallurgical coke making. This is the fourth of a series of reports giving the results of studies by counties under phase (1) of the investigation. It covers Armstrong County, Pa., which comprises parts of the Foxburg, Clarion, Brookville, Kittanning, Bural Valley, Smicksburg, Freeport, Elders Ridge, and Indiana quadrangles. (See fig. 1.) A base map for each bed in each quadrangle was made to the scale of 1 inch equals 1,200 feet. Maps of mines, locations of drill holes, bed and total coal thicknesses, and the outcrop of the bed were plotted on the maps. With all available data plotted, isopach lines were drawn to limit areas of known unmined reserves in beds 0 to 14 inches thick, 14 to 28 inches thick, 28 to 42 inches thick, and more than 42 inches thick. These areas of coal reserves also were divided into "measured" and "indicated" categories. All areas in each thickness range and in each category, mined-out areas, areas excluded from the estimate, and areas outside the outcrop were measured by planimeter on the base maps, Estimates of total reserves 14 inches and more thick and maps for individual beds were prepared from these data.

Jan 1, 1951

By McElroy. G. E.

"INTRODUCTION The ventilation of development faces in mines for diluting gas or dust concentrations and, in hot mines for providing air-motion cooling effect, by fan pipe units discharging high-velocity streams of air close to the face, is an important phase of mine ventilation. The effectiveness of such units depends largely on the air-flow distribution between the pipe discharge and the face, a subject on which little definite information is available. A study of the air flow from grills into rooms 3/ has provided some basic data, but under conditions no different from mine fan-pipe installations in development ends that the data are of doubtful application to mine installations, owing to differences in air-flow conditions.The purpose of this report is to present and discuss the results of an initial series of tests of a general investigation designed to yield definite laws and design data for the discharge of fan-pipe lines in mines. This series was planned to yield preliminary data on the effects of velocity of discharge, size and shape of pipe, and adjacent walls on the air-flow distribution. Tents were made on 10-inch pine with seven different end conditions and one change of wall condition. Changes in size and shape of discharge wore made by short transformation pieces directly at the discharge end, with the result that effects due to contraction and expansion of the flow, which proved to be of much greater importance than was expected, totally obscured any normal effects due to size and shape. The results of this series of tests, therefore, relate to 10-inch pipe with various discharge designs and are so reported."

Aug 1, 1940

By K. A. Johnson, H. F. Yancey

"INTRODUCTION Coals vary so widely in their burning characteristics that all coals prepared for domestic stoker use cannot be burned efficiently on every type of stoker. The overfeed stoker, a large number of which are in use in the Pacific Northwest, operates satisfactorily when burning low- to high-ash, noncaking bituminous or subbituminous coals, whereas on the clinkering-type, underfeed, bituminous stoker both caking and noncaking bituminous and subbituminous coals are burned, generally with ash contents below 12 percent and ash-softening temperatures between 1,800º and about 2,600º F. Because fly ash continues to be one of the most troublesome factors in domestic underfeed stokers, noncaking coals especially must be chosen carefully with respect to size composition, particularly as regards the exclusion of excessive fines. Their formation as a result of degradation in storage handling or because of crushing in .the stoker worm will increase the fly-ash nuisence.Bituminous and subbituminous coals with higher ash-softening temperatures and, often, higher ash contents have not been able to enter the domestic stoker-coal market in the Pacific northwest because clinkers are not formed readily during clinker-type underfeed-stoker burning, especially in mild weather, and because of the labor and attendant inconvenience required in handling larger volumes of ash. The work described in this report was undertaken to enable mines producing coals of the kind described to enter the stoker-coal markets. The anthracite stoker selected was one that conveys the refuse from the ash pit of the furnace and deposits it into one of two covered, dust-tight disposal cans."

Dec 1, 1945

[This Bureau of Mines report presents results of a survey of underground mining equipment and of two demonstration programs showing the feasibility If quieting; a load-haul-dump (LHD) machine and a personnel vehicle. Typical noise levels are presented for the major machine types used in underground mines, along with estimates of the noise overexposure of miners who operate or work near these machines. General principles of noise control are explained, and the application of these principles is illustrated in the description of modifications made to the LHD machine and the personnel vehicle. The noise control package installed On the persona of vehicle lowered its noise level by 14 dBA after 4 months of operation, and inspection of the LHD machine after 2 1/2 years of operation with the modifications showed its noise level 7 dBA lower than that of the unmodified machine. Noise dosimeter measurements indicated that both machines were in compliance with Federal noise regulations for a typical shift.]

Jan 1, 1980

To minimize the equipment operator?s respirable dust exposure in enclosed cabs of mobile equipment by install-ing cab filtration and pressurization systems and by keep-ing doors and windows closed. Background Enclosed cabs are used on mobile equipment to protect operators from health and safety hazards. The primary health concern is overexposure to respirable dust. Examples of such equipment in mining include drills, dozers, haul trucks, and other equipment. When equipment is new, the operator?s dust exposure is normally at acceptable levels. As the equipment ages and gaskets and seals deteriorate, the air quality inside an enclosed cab can reach a point where it is no longer at acceptable levels for the equipment operators. In an effort to improve the air quality in enclosed cabs of older mining equipment, NIOSH has been performing research to retrofit cabs with new filtration and pressurization systems. Over the past few years, this research has shown that retrofit systems can substantially improve the cab?s air quality back to safe and acceptable levels. How-ever, this is only possible when a positive cab pressure is achieved and maintained. Thus, when a cab door or window is opened, the cab pressure drops to zero and the cab filtration and pressurization system is rendered ineffective. During a recent field study, the magnitude of this impact was noted when a drill operator repeatedly opened the cab door throughout the course of the drill cycle. Upon closer examination, it was determined that the cab door was being opened to manually guide the next drill steel into place each time an additional section of steel was needed to drill deeper holes.

Sep 1, 2008

By M. H. Rinkenbach, C. A. Taylor

The detonating compound cyanuric triazide was discovered*** and patented**** by Erwin Ott, and on account of his claims, the Bureau of Ordnance, U. S. Navy, desired more complete data, so an investigation was made. The results obtained were such that the Bureau of Crdnance recommended publication thereof for the benefit of the explosives industry. Some of the patent claims regarding cyanuric triazide are as follows: 1. It is not poisonous in contradistinction to fulminating mercury and lead azide. 2. In contradistinction to the hitherto employed priming agents it is fusible, and its fusion point (94°C) is so low that the product melted by a single heating in water or steam bath, can be poured without danger in the capsule shell. It solidifies therein after cooling, and shows the density of 1.5. 3. It has been determined that the cyanuric triazide, in spite of its very much higher priming actions, is much less brisent then ful- minating mercury and lead nitride. 4. The chemically pure cyanuric triazide (C3M2) constitutes colorless crystals melting at 94°0., and is not volatile. It is not affected by cold water, and only slowly saponified by hot water. It is relatively insensitive to friction and not too strong concussion, but explodes by impact or when hea ted to 170°C., or when heated suddenly. It is not more sensitive to friction and shocks then fulminating mercury. 5. The preparation of the triazide is preferably carried out by acting with an excess of sodium azide in an aqueous solution; in the cold, on cyamric chloride gradually suspended in the said solution.

Aug 1, 1923

By O. C. Blade

This survey report, the first for this type of fuel, is issued to provide information on the characteristics of the burner-fuel oils now manufactured for sale. Reports of surveys of diesel fuel, conducted under a cooperative agreement between the American Petroleum Institute and the Bureau of Mines, United States Deportment of the Interior, have been published annually since 1950.2/ This survey of burner-fuel oils is similar to the diesel-fuel surveys In pattern and has been conducted under an extension of the same agreement at the request of the Oil-Heat Institute of America and the petroleum industry. It is not intended to conflict with the diesel-fuel survey but rather to supplement it. Burner-fuel oils, also specifically referred to as heating oil, range oil, furnace oil, "Bunker C," and fuel oil, are sold in increasing volumes for use in this country and for export. Sales of all grades of fuel oil in the United States have increased from 340 million barrels in 1926Y to over a billion barrels per year. 4/ Distillate fuel oil sales, in proportion to total fuel-oil volume sold, increased from 42 percent in 1950 to 50 percent in 1954. Sales of distillate and the heavier residual oils during recent years are shown: [ ] It must be realized that the data given are not accurate indications of the use of these materials as burner fuels. Some burner fuel, used as range oil, is kerosine, which is not included above. A large volume of fuel oil--approximately 30 percent of the distillate type and some residual fuel oil--is sold as diesel fuel. Smaller

Jan 1, 1955

By Oliver Bowles

"A New Mica Company in China A report has been received from the American Consulate at Chungking, China, to the affect that a new company has been organized to operate a mica mine north of Ta-chien-lu, in the province of Ssu-Chuan. The company, which is entirely under Chinese management and control, has a working capital of $10,000. While comparatively new, the mine has been worked before, and apparently with some success. The mica is said to be of good quality. It is claimed that the deposit will supply a limited amount of mica that will cut in pieces a foot square or over, and larger quantities of smaller sizes. The chief difficulty is transportation, for it is reported that a ten days' journey, part of it over mountain trails, will be necessary to transport the mica from the mine to Yachow, the nearest practical waterway. It is probable, therefore, that the output of the mine will be limited.Necessity for Careful Trimming and Grading of MicaA recent report to the Department of Trade and Commerce, Ottawa, Canada, by Acting Trade Commissioner C. G. Venus, London, England, contains information which is of vital interest to mica producers in the United States. It is pointed out that mica exports from Canada to England show a gradual decline dur¬ing recent years, the figures being as follows: 1913- 1,383 cwt: 1914 - 1,225 cwt: 1915 - 1,864 cwt: 1S16 - 879 cwt: 1917 444 cwt: 1918 - 480 cwt. Shipments from India to England, on the other hand, show a steady increase from 40,178 cwt. in 1913 to 66,401 cwt, in 1918."

Jan 1, 1920

By Harry R. Nicholls

The objectives of this investigation were to determine the strain propagation laws for a strain pulse in salt and to compare the seismic effects from the detonation of different explosives in salt. The investigation showed that a strain propagation law of the type, e = K(R/W 1/3 ) n, and a strain energy propagation law of the type, Ea/W 1/3 = K(R/WI/3)b , are valid in salt. A change in the regression law coefficient, n , was noted with strain level. A similar change in b was inferred because strain energy is proportional to the square of strain amplitude. Particle velocity data may be used interchangeably with strain data under the test conditions.

Jan 1, 1962

By M. John Brodie

This report reviews the regulatory environment as it pertains to tailings management and disposal from mining operations, excluding coal and placer gold operations, in the state of Alaska. It is based on the EPA's staff position on an effective program to regulate mining wastes as presented in the May, 1990 Strawman II document, The unique conditions which occur in parts of the state, including a fragile environment, high seismicity and permafrost, are identified. The special waste management practices which are necessitated by these conditions are delineated. The report provides a detailed outline of the current technologies that are suitable for designing, constructing, operating and closing a railings disposal facility to meet environmental protection objectives. Cost implications to the mining industry of potential regulatory actions, such as the degree of groundwater protection are examined.

Jan 1, 1992

By A. T. Iannacchione

Synopsis Underground stone mining Is an emerging sector of the U.S. mining industry. As this expansion takes mines under deeper cover, and as more efficient mining methods are utilized, effective stone pillar design methods will become even more Important. Current design practices are examined and a discussion of safe mine layouts is presented as a first approach towards weighing the demands for increased production against increased risk. Risks to underground stone-mine workers include rib Instabilities, pillar failures and roof falls. Seventy-two stone-mine pillar designs were examined. Pillars with width to height ratios of less than 1.5 appear more likely to fail when subjected to excessive stress levels. When width to height ratios fall below 1.0 defects in the pillars, such as through-running discon¬tinuities, can have a significant influence on stability. Discontinuity persistence, dip, material properties and orientation are also important determining factors in pillar strength. During the past three years the number of active underground stone mines in the U.S.A. has ranged between 90 and 100. This number is expected to increase as the crushed stone industry responds to growing demands for its products.1 As more of the industry moves to benefit from the advantages of underground mining production from underground stone mines is expected to increase above its current level of approximately 66 000 000 r/year. Parker2 identified four advantages of underground mining operations: (1) surface developments, zoning laws and environmental concerns are often less of an issue; (2) stripping and restoration requirements are eliminated; (3) additional reserves are often available beneath the quarry floor, under pit slopes or under an adjoining property; and (4) space is created for secondary use. Underground mines enjoy the added benefits of work in a constant underground climate, rather than in the variable surface climate, minimization of community concerns by placement of the crushing, sizing and stowing operations underground and reduction of surface vibration concerns through smaller blasts. The drawbacks of underground mining include added health and safety hazards for the stone miners, whose health can suffer from increased exposure to falls of ground, airborne contaminants and fog in large underground openings. Injuries from falls of ground in stone mines have occasionally exceeded the incidence rates for other mineral resources mined underground. 3 Existing underground stone operations, by comparison with stone operations a decade ago, mine more stone at a faster rate and with larger equipment. Because of high demand mines face increased pressure to yield more stone per production blast. A majority of the miners use the V-cut blasting pattern, which limits the depth or pull for each shot to about 4 m. Therefore, to produce more stone, either the mines must work more faces or existing faces must be enlarged. The enlargement of underground openings poses problems in the maintenance o€ strata stability, First, the widths of the mine entries partially control the amount of sag or deflection that any given roof beam can withstand before failure. This deflection can take pi ace quickly after an opening is excavated or much later as weathering processes form additional, thinner beams. Because increased deflection increases the potential for roof beam failure, there is a limit to how wide a room can safely be made. These limits depend on local geological and stress conditions and, to some extent, the cost effectiveness of roof-bolting. Given these room width limitations, underground stone mines have placed more attention on expansion of production through benching where mining thicknesses permit. The heights of rooms partially control the strength characteristics of adjacent pillars: as the room height increases for a given pillar width the pillar width-to-height ratio (w/h) decreases. In general, stone pillars are very strong, but when the width-to height ratio diminishes the potential for pillar rib instabilities and pillar failures increases (see Fig. 1). [ ] In the underground environment the desire to utilize wider rooms, higher benches and multi-level mining results in the potential for unstable ground conditions. Although the number of failed pillars is currently very low, the potential for additional failures will grow if more slender pillars are developed, especially under deeper cover, with wider mining sections and in multi-level operations. It should also be noted that the passage of time acts to decrease the strength of many existing pillars. An examination of current design practices

By F. E. Cash

"Explosions in bituminous coal mines are preventable. The combination of adequate ventilation, thorough rock-dusting, and water on the cutter bars of mining machines and of empty and loaded cars will certainly reduce the hazard to a minimum, if not remove it entirely. This is the attitude taken by the Woodward Iron Co. in lessening or removing the dust and gas explosion hazard from its Dolomite and Mulga mines.This report was made possible through the willingness of Frank H. Crockard, president of the Woodward Iron Co. to make records available and to authorize their publication for the benefit of the coal-mining industry.The compilation of these data and preparation of maps was facilitated by the cooperation of M. Heard, general superintendent of coal mines, J. B. Whatley, assistant general super-intendent of coal, and R. P. Whetmore, chief mining engineer.The Dolomite and Mulga mines are about 16 miles southwest of Birmingham, Jefferson County, Ala., on a company owned and operated railroad which connects ore mines, coal mines, coke ovens, and furnaces.In the Dolomite property there is a 30°, 860-foot, double-tracked slope for hoisting, three manway and air slopes and one air shaft, all in rock.At Mulga there is a 220-foot hoisting shaft, an air shaft and two manways, and material and air slopes, all in rock to the coal bed.The coal in both mines comes from the Pratt bed, which is the next coal above the Mary Lee bed in the Warrior Basin. The coal is horizontal or slightly undulating, with thickness varying from 36 to 72 inches and averaging 50 inches.The method of mining is essentially room-and-pillar with modifications for mechanized sections. The mines produced approximately 1,100,000 tons with 2,328,309 man-hours exposure n 1929 and 900,000 tons with 1,697,436 man-hours exposure in 1930.Through the work of the safety department the entire personnel in the coal mines, as well as in other departments, makes safety an essential part of efficient operation.Both mines are gassy; Dolomite produces about 334,000 cubic feet and Mulga about 919,000 cubic feet of methane every 24 hours.The ventilation maps, two of Dolomite mines (figs. 1 and 2) and one of Mulga (fig. 3) give as much of the details of ventilation as is practicable to put on maps of such scale. As will be seen from these maps there are two main ventilating fans in Dolomite mine and two in Mulga mine and the general direction of the air travel is indicated by long arrows to fan 1 and short arrows to fan 2 in each mine. Each fan has a gas or steam auxiliary drive and each mine has a complete fan and power unit ready for operation in emergency.In Dolomite, recently taken data show that fan 1 exhausts 82,000 cubic feet of air per minute at 3.2-inch water-gage pressure, with an analyzed methane content of 0.15 per cent, and fan 2 exhausts 115,500 cubic feet of air at 3.8-inch pressure and 0.10 per cent of methane. In addition, a Capell fan at No. 2 air shaft is a standby or emergency unit.In Mulga recently taken data show that fan 1 exhausts 83,000 cubic feet of air per minute at 3.0-inch pressure and 0.18 per cent of methane, and fan 3 exhaust 119,000 cubic feet of air per minute at 5.4 inch pressure and 0.35 per cent of methane. Fan 2 at No. 2 shaft is a standby or emergency unit for the Mulga mine.The ventilation is on a split system in both mines. All part split and total returns are measured, sampled, and analyzed weekly, requiring one man two days each week. During December, 1930, 77 samples were collected and analyzed in Mulga and 78 in Dolomite. Two weekly reports covering part of December, 1930, are given here so that they can be studied to show control of ventilation in the Dolomite and Mulga mines. In the event that a split sample shows more than 0.50 per cent of methane, immediate steps are taken to lower the explosive-gas content. The reports follow:"

Mar 1, 1932

By Maria I. De Rosa

Table 10 and figure 4 show the number of fires and fire injuries occurring at the surface of underground coal mines by state during 199&1999. Table 10 also shows by state the risk rate, employees' working hours, and lost workdays. A total of 65 fires occurred in 10 states. Thirteen of those fires caused 12 injuries and 1 fatality (the yearly average was 6.5 fires and 1.2 injuries). Four fues and one fire injury involved con- tractors. The Ewhr value was 97 x l O6 hr (Irr = 0.025); the LWD value was 6,206. Pennsylvania had the most fues (20 fues and 5 injuries), followed by West Virginia (1 6 fires and 1 fatality) and Kentucky (1 5 fires and 3 injuries). Among these states, Pennsylvania had the highest injury risk rate value (Irr - 0.095). Table 1 1, partly illustrated in figure 5, shows by time period the number of fires, fire injuries, and fire fatalities; risk rates; employees' working hours; and lost workdays. The number of fires and fire injuries show a decrease followed by an increase during the five time periods, accompanied by a decline in employees' working hours throughout the periods (see table 11 and figure 5). The Irr values follow patterns similar to those shown by the injury values. Tables 12- 17 show the number of fires by ignition source, method of detection and suppression, equipment involved, location, and burning material by time period. Figure 6 shows the major variables during 199&1999. Table 18 shows the number of fire injuries per number of fires causing injuries and total fires by year, ignition source, equipment involved, and location.

Jan 1, 2004

By J. E. Pahlman

As part of its goal to minimize the requirements for mineral commodities through conservation and substitution, the Bureau of Mines is conducting research to determine the resistance of alumina refractories to corrosion by alkali-containing slags that could result from burning of the abundant Western lignites in metallurgical operations. Static corrosion tests were performed at 1,000° to 1,350° C and dynamic tests were performed at 1,400° to 1,600° C to investigate the stability of refractory specimens in slag environments. In general, the degree of refractory attack increased with decreasing alumina content in the refractory and with increasing alkali content of the lignite ash. The parabolic kinetics observed in static tests suggested that a reaction product layer is formed through which ionic diffusion must occur to sustain the reaction with the refractory material. The linear kinetics observed in dynamic tests in high-viscosity slags indicated a shearing of the reaction product layer by viscous slags. The observed lesser attack of refractories by low-viscosity slags can be explained on this basis. Two 99-percent-alumina refractories and two 90-percent-alumina refractories formed an expanded protective layer with the more viscous slags but not with the less viscous slags. Corrosion of 90-percent-alumina refractories increased exponentially with temperature. In refractories with the same nominal alumina content, those that contain a combination of a-Al203 and mullite (3A1203?2Si02) have better corrosion resistance than those containing a combination of a-Al203, cristobalite (Si02 mineral), and mullite. In alkali environments (1,300 to 1,400° C), 90-to 99-percent-alumina refractories are recommended for use as furnace linings.

Jan 1, 1980