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By Michael S. Wieland

The work-principle method was developed for theoretically resolving toxic fumes from non-ideal charge formulations. The whole description of a wanted process refers to macroscopic changes resulting from removing and imposing constraints that work one at a time to chronologically render trajectory terminal (key) states. The trajectory resolution utilizes a thermodynamic quasi-potential or negative work function, and numeric results require some appropriate standard thermodynamic (equilibrium) reaction code TDRC. The underlying method is unable to restrict reactions, tailor the work output or resolve states within the trajectory transitions, so the trajectory terminology is used somewhat figuratively. The work-principle model has the working fluid (reaction material) interacting with a tremendous reservoir (chamber air and walls), while forecasting the work performed upon a mechanical agent. Within the underground test chamber, the widening fume zone undergoes a decline in temperature, a reduction in reaction rates, resulting in out-of- equilibrium circumstances at the Q*-point, representing the measurement state. Rather than coping with the hard issues of non-equilibrium, the toxic fumes are resolved numerically for the Z-point, rendered by a zero-net- interaction energy constraint. Though other techniques require complex equation-of-state EOS for the whole trajectory, the work-principle formulation eludes the issue, requiring an EOS that works just for the terminal-state. Under the circumstances, the Z-state is reconcilable with the ideal-gas domain, and relevant TDRC results yield the restricted state in a numerical way for different charge formulations. The work principle makes it hard to remedy or to tamper with the resulting Z-point fumes: non-macroscopic trajectory readjustment is forsaken, terminal states are results rather than stipulations, and resetting the EOS formula constants would be ineffectual. Though the non-equilibrium portion remains unresolved, it is noteworthy that the work principle rigidly fixes the Z-state, thereby reducing the uncertainty for rendering reaction-rate models characterizing the non-equilibrium Z-Q* transition. Though more refinements are needed, some Z-point fume results are included to illustrate the threshold (minimum) resolving capacity of the work- principle formulation.

By Donald P. Tuchman, Robert P. Vinson, Steven J. Page, Linda J. McWilliams, Gerald J. Joy, Steven E. Mischler, Jon C. Volkwein

The United States National Institute for Occupational Safety and Health, through an informal partnership with industry, labor, and the United States Mine Safety and Health Administration, has developed and tested a new instrument known as the Personal Dust Monitor (PDM). The new dust monitor is an integral part of the cap lamp that coal miners normally carry to work and provides continuous information about the concentration of respirable coal mine dust within the breathing zone of that individual. Previous laboratory testing demonstrated that there is a 95% con?dence that greater than 95% of individual PDM measurements fall within ±25% of reference measurements. The work presented in this paper focuses on the relationship between the PDM and respirable dust concentrations currently measured by a coal mine dust personal sampler unit utilizing a 10 mm Dorr Oliver nylon cyclone. The United Kingdom Mining Research Establishment instrument, used as the basis for coal mine respirable dust standards, had been designed speci?cally to match the United Kingdom British Medical Research Council (BMRC) criterion. The personal sampler is used with a 1.38 multiplier to convert readings to the BMRC criterion. A strati?ed random sampling design incorporating a proportionate allocation strategy was used to select a sample of mechanized mining units representative of all US underground coal mines. A sample of 180 mechanized mining units was chosen, representing approximately 20% of the mechanized mining units in production at the time the sample was selected. A total of 129 valid PDM/personal sampler dust sample sets were obtained. A weighted linear regression analysis of this data base shows that, in comparison with the personal sampler, the PDM requires a mass equivalency conversion multiplier of 1.05 [95% C.I. (1.03, 1.08)] when the small intercept term is removed from the analysis. Removal of the intercept term results in a personal sampler equivalent concentration increase of 2.9% at a PDM measurement of 2.0 mg m-3.

By Maria I. De Rosa

Table 46 and figure 16 show the number of fires and fire injuries for metal/nonmetal mills by state during 1990-2001. Table 46 also shows the injury risk rates, employees' working hours, and lost workdays. In all, 77 fires occurred in 26 states. Thirty-seven of the fires caused 41 injuries and 1 fatality (including 5 fires and 3 injuries involving contractors). The yearly average was 6.4 fires and 3.4 injuries. Fifty-four fires with 26 injuries and 1 fatality occurred at metal mills; 23 fires with 15 injuries occurred at nonmetal mills. The Ewhr value was 845 x 106 hr (Irr = 0.01), and the LWD value was 6,681. Minnesota had the most fires (16 fires and 5 injuries), followed by Arizona (9 fires and 3 injuries), Wyoming (8 fires and 3 injuries), Nevada (7 fires, 4 injuries, and 1 fatality), and Texas (4 fires and 7 injuries). Of these states, Texas had the highest injury risk rate value (Irr = 0.033). Table 47, partly illustrated in figure 17, shows the number of fires, fire injuries, fire fatalities, risk rates, employees' working hours, and lost workdays by time period. The number of fires increased slightly during the second period, then decreased during the remaining periods. The number of fire injuries show a decrease during most of the periods (an increase is seen during the third period), accompanied by a decline in employees' working hours during most of the periods. The Irr values follow patterns similar to those shown by the injury values. Tables 48-53 show the number of fires by ignition source, method of detection and suppression, equipment involved, location, and burning material by time period. Figure 18 shows the major variables related to fires for 1990-2001. Table 54 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 L. Y. Sadler

The Bureau of Mines conducted a study Lo determine the loss in weight of silica (Si02) volatilization from low-alumina (A1203) refractories used as hot face liner materials in such process equipment as coal gasifiers, synthesis gas transfer lines, and ammonia plant secondary re-formers. Small specimens of 80-and 40-pct-A1203 brick and 95-pct-SiO2 brick were exposed to steam in high-pressure furnaces at pressures of up to 1,000 psig and temperatures up to 1,400° C for periods ranging up to 120 h. Weight losses from the 40-pct-A1203 brick after 40 h of exposure were highest, 2.75 pet at 1,400° C and 1,000 psig. The rate of weight loss had dropped to near zero after 80 h of exposure at all temperatures. No relationship between total silica content of the specimen and weight loss could be established nor did scanning electron microscopic examination of the specimens before and after exposure identify specific regions (e.g., surface) suffering a disproportionate share of the silica loss. It was concluded that minor silica volatilization is likely to occur, if at all, at temperatures of less than 1,000° C and pressures below 1,000 psig, and that low-alumina refractories should be used as dry-ash coal gasifier hot face linings.

Jan 1, 1986

By E. J. Lintner

This paper is one of a series being prepared by the Bureau of Mines describing; gypsum mining and crushing methods and costs at various operations throughout the, United States. These papers are designed to dissemminate technical information regarding methods used. The cost tabulations represent local operating expenditures only and not total production costs. It is recognized that publication of total production costs may in many instances cause embarrassment to the individual operator as well as to the industry as a whole; on the other hand, operating costs are essential to the technical discussion and. study of the methods employed. The attention of the reader is specifically called to this differentiation in order that no misunderstanding of the scope of the cost tabulation shall ensue. The following report deals specifically with the mining and primary crushing methods employed at the plant of the Ebsary Gypsum Co., Inc., Wheatland, Monroe County, N. Y. This company is capable of maintaining a maximum yearly production of 220,000 tons of crushed gypsum, thereby ranking as one of the largest gypsum operators of the State. The methods and practices employed typify the standard methods followed by gypsum producers in western Now York State, The bed of gypsum mined lies only 35 feet beneath the surface and is above average thickness for this vicinity.

Jan 1, 1937

By W. T. Abel

The difficulties associated with wet-processing fine size coal led the Bureau of Mines to investigate a dry process for removing pyrite from coal pulverized to powerplant fineness. The dry process that was investigated consists of two cleaning steps: First, centrifugal separation is used to remove a fine, pyrite-depleted portion as a product; second, the remaining coarser, pyrite-enriched portion is taken to electrostatic separation in which pyrite is concentrated and rejected. Ball milling and Hardgrove milling were used to pulverize the coal. Three coals from the Pittsburgh seam were treated by the dry-removal process. Over 55 pct of the pyrite, which amounted to 70 pct of the available pyrite, was removed in a reject of 10 to 20 pct from one of the coals. Similar treatment of the other two coals removed about 35 to 50 pct of the pyritic sulfur, which amounted to about 70 pct of the available pyritie, sulfur from these coals. Both centrifugal and electrostatic separations were more efficient when pyrite particles were as large as possible. Separation efficiency was improved by confining the particle size of the feed to a relatively narrow range and by avoiding excessive amounts (more than 10 pct) of dust in the electrostatic range. The removal efficiency was increased with the use of stage grinding.

Jan 1, 1973

By H. S. Caldwell

Molten zinc, condensed from zinc vapor, was used by the Federal Bureau of Mines to leach aluminum from an aluminum-silicon alloy made by carbothermic reduction of silicious aluminum ores. Zinc requirements were minimized by re-fluxing condensed zinc in a low-pressure, Soxhlet-type distillation unit. The minimum ratio of zinc used to aluminum in the crude alloy was approximately 2:1. When the system pressure was reduced from atmospheric pressure to 100 microns or less, the rate of aluminum extraction increased by a factor of 20. Leaching rates were much higher when the zinc vapor was condensed directly on the crude alloy than when molten zinc was passed over the alloy in the absence of appreciable zinc vapors. Observations indicated higher extraction efficiencies from crude alloys containing greater amounts of iron. Complete extraction was obtained with negligible loss of zinc. Assuming complete removal of zinc, the aluminum products will contain 1 to 6 percent silicon and about 0.5 percent iron plus titanium. The leached crude-alloy residue is a highly friable spongelike material containing about 80 percent silicon, several percent iron and titanium, with the remainder aluminum. This extraction method has been patented (Patent 2,810,637) and assigned to the Government. Research and development investigations are continuing at the College Park Metallurgy Research Center.

Jan 1, 1961

By James P. McGee

Use of nuclear reactors to furnish high-temperature heat is a field that has not received as much attention as propulsion and electric power production. Investigations are being conducted under a joint Atomic Energy Commission (AEC)--Bureau of Mines research program to determine the feasibility of using nuclear heat for chemical processing. A part of this process-heat-reactor program is the study of a system in [hich] helium transfers heat from 8 nuclear reactor to a chemical processing [essel]. This requires construction of (1) a nuclear reactor to operate with helium exit temperature of 2,500° F. and with pressures in the range of 250 to 500 p.s.i.g. and (2) a processing vessel (heat exchanger) to convert the enthalpy of the helium into chemical process reaction heat. Neither component exists at present, and development of each promises to be a formidable task. To aid in the study of such a system, the Bureau of Mines has designed and constructed a vessel heated by electrical induction to simulate a source of nuclear fission heat. This vessel is undergoing preliminary testing. Gasification of coal and reforming of hydrocarbons are promising uses for nuclear heat. It appears that a higher price per unit of nuclear heat can be tolerated for chemical processing than that allowable for power generation.

Jan 1, 1960

By A. W. Ambrose

A few years ago, (March 3, 1915) Congress passed a law providing for the establishments of ten mining experiment stations under the supervision of the Bureau of Mines. In 1917 the Secretary of the Interior decided that one of these stations should be devoted to Petroleum, and in the early part of 1918, approved the establishment of the Petroleum Experiment Station at Bartlesville, Oklahoma, upon the condition that the city furnish the site for the buildings and guarantee $50,000. for the erection of the buildings. A building site of 5 acres and the necessary building fund was furnished by the Bartlesville Chamber of Commerce.The buildings consist of an administration and laboratory building, both of brick construction. The former is a two-story structure of eight rooms. A large room occupying over half of the main floor is devoted to the library, reception room, and clerical force. The remaining rooms comprise a drafting room and six offices occupied by the staff assigned to this station. The laboratory is a one-story brick building and contains a well-equipped chemical and physical laboratory, also a machine shop and mechanical laboratory. In addition to the main buildings there now are on the property an experimental refinery, a blue-print house, and a store house. These latter buildings are all of sheet-iron construction.

Oct 1, 1920

By Robert L. Davidoff

Growing domestic and global demands for mineral resources have established an imperative need for developing comprehensive, versatile mineral supply analysis evaluation tools. The Minerals Availability System (MAS) of the Bureau of Mines is meeting this need with its computerized Supply Analysis Model (SAM). This paper describes the SAM system and the quantitative methods used by the MAS for mineral supply analyses. The MAS evaluating process consists of deposit identification, tonnage and grade determinations, engineering and cost evaluations, financial analyses, and supply curve generation. MAS has developed this computerized system primarily to facilitate the maintenance of data information for the constructing of commodity supply curves. The SAM system presently has the analytical capabilities of simultaneously (I) evaluating many mineral deposits, (2) updating mineral supply information to reflect inflationary changes in costs and prices, and (3) conducting sensitivity analysis to determine impacts on mineral supply under various conditions. Many potential users in both government and industry will find the SAM system a valuable tool for policy analysis, forecasting, and planning. The SAM system has been constructed so that any types of information that are similar to the data elements in the MAS data base may be used to perform supply studies. The storage space needed for the SAM system?s programs, files, and output data is approximately 14 million characters.

Jan 1, 1980

By W. H. Kerns

The Taylor Creek lead-zinc deposit is on Taylor Creek, approximately 1-1/2 miles from tidewater, at the head of Duncan Canal, on the east side of Kupreanof Island, Petersburg district, southeastern Alaska. The property was located in 1903, and a 30-foot discovery cut was put to the west tank of the creek. Very little prospecting end developing has been done since that time. The property was examined by J. C. Roehm, associate mining engineer, Territorial Department of Mines, Juneau, Alaska, in June 1946. From exposure n the creek, a mineralized zone with possible dimensions of 1,700 feet by 1,000 feet was indicated. The general strike was northwest, and the dip was low and variable to the northwest. During September through November 1948, the Bureau of Mines conducted a project on the property. Four diamond-drill holes, totaling 770.5 linear feet, were drilled from the surface and 280 lineal feet of hand trenching done. The mineral deposit was found to occur as patches that appear to be a replacement in dolomitic limestone. The patches of mineralization are small, irregularly shaped, and irregularly distributed. Maps showing the location of the drill holes and trenches, analyses of the cores from the drill holes, and anal lee of the samples from the trenches accompany this report (see, also, appendixes A, B, and C).

Jan 1, 1950

By S. M. Shelton, M. M. Fine, T. L. Johnston

"INTRODUCTION The Bynum property is about 15 miles west of Gadsden and 1 mile north of Walnut Grove, Etowah County, Ala. The ore lies in layers and as irregular masses and pockets in the clay residue from decomposed quartzite and limestone formations to the district.The property has been developed by means of numerous shallow shafts sunk to water level - about 25 foot. Operations are confined to small-scale, selective mining of pockets of high-grade ore. The ore is marketed directly, without beneficiation. The property does not yield a large amount of manganese ores. No estimates of reserve ore tonnages are available.A 206-pound sample was furnished for test. It was taken at random from a stock rile estimated to contain 130 tons of ore. The sample con¬sisted of a mixture of manganese oxides, iron oxides, and siliceous fragments in clay. The manganese occurred largely as soft, unconsolidated pyrolusite and wad. Some hard, crystalline pieces of psilomelane were distributed through the soft material. Small amounts of iron oxides were associated with the manganese oxides. Chert, quartzite, and clay comprised the gangue minerals.Dry analysis of the sample was as follows: 36.0 percent manganese, 6.1 percent iron, 2: .0 percent silica, 1.5 percent calcium carbonate, 0.36 percent phosphorus, 3.0 percent alumina, 2.2 percent barium, and 0.31 percent zinc."

May 1, 1942

By Robert C. Hickman

"INTRODUCTION The Big Four fault and the Sandy Creek fault, with several cross faults, traverse the Stalion-Curnel fluorspar prospect in a northeasterly direction. The Big Four mine, about three-quarters of a mile northeast of the prospect, is the only active mine in the immediate area.A resistivity survey was made of a portion of the prospect by the Division of Geophysical Exploration, Bureau of Mines, in May and June 1946. As a result of this survey, five holes were drilled by the Bureau of Mines during the period from October 8 to December 16, 1946. Two Longyear No. 12 Junior Straitline diamond drills were used for the work.ACKNOWLEDGMENTSIn its program of investigation of mineral deposits, the Bureau of Mines has as its primary objective more effective utilization of our mineral resources to the end that they make the greatest possible contribution to national security and economy. It is the policy of the Bureau to publish the facts developed by each project as soon as practicable after its conclusion. The Mining Division, Lowell B. Moon, chief, conducts preliminary examinations, does the actual investigative work, and prepares the final report. The Metallurgical Division, Oliver C. Ralston, chief, analyzes samples and makes beneficiation tests.The project was under the supervision of Mitchell H. Kline, chief, Raleigh Branch, Mining Division."

Mar 1, 1948

By John R. Duda

The Bureau of Mines is developing estimates of capital investment, operating cost, and selling price for underground mines producing bituminous coal by various mining methods. This study estimates the required capital investment, operating costs, and selling prices for two hypothetical mines designed to produce 1.5 and 3 million tons per year (MM tpy)3 by using a longwall system in conjunction with a continuous-mining system. The coal properties being mined are assumed capable of sustaining a 20-year product Lon period. Wages and union welfare payments used in this study are those in effect as of December 6, 1975, as set forth under the National Bituminous Coal Wage Agreement of 1974. Costs of materials and equipment are based on January 1976 Indexes. Initial capital investment is $33.8 million for the 1.5-MM-tpy mine and $61 million for the 3-MM-tpy mine. Capital requirements, including deferred investments, were estimated to be $52.5 and $94.2 million, respectively. These figures translate into total capital investments of $34.87 and $31.31 per ton of annual production. Assuming a desired 15-percent discounted cash flow rate of return, with Federal income taxes at 50 percent, the required selling prices would be $12.06 and $10.94 per ton for the 1.5- and 3-MM-tpy mines.

Jan 1, 1976

By John R. Duda

The Bureau of Mines is developing estimates of capital investment, operating cost, and selling price for underground mines producing bituminous coal by various mining methods. This study estimates the required capital investment, operating costs, and selling prices for two hypothetical mines designed to produce 1.3 and 2.6 million tons per year (MM tpy)3 by using a longwall system in conjunction with a continuous-mining system. The coal properties being mined are assumed capable of sustaining a 20-year production period. Wages and union welfare payments used in this study are those in effect as of December 6, 1975, as set forth under the National Bituminous Coal Wage Agreement of 1974. Costs of materials and equipment are based on January 1976 indexes. Initial capital investment is $34.6 million for the 1.3-MM-tpy mine and $63.5 million for the 2.6-MM-tpy mine. Capital requirements, including deferred investments, were estimated to be $50.8 and $92.2 million, respectively. These figures translate into total capital investments of $38.51 and $34.91 per ton of annual production. Assuming a desired 15-percent discounted cash flow (DCF) rate of return, with Federal income taxes at 50 percent, the required selling prices would be $13.41 and $12.27 per ton for the 1.3- and 2.6 -MM- tpy mines.

Jan 1, 1976

By Christopher C. Jobes, Alan G. Mayton

The National Institute for Occupational Safety and Health (NIOSH) researchers have investigated seat design issues for the occupants and operators of underground mine transport vehicles relative to whole-body vibration (WBV) and mechanical shock. Considering the ergonomic and engineering improvements made to underground mine shuttle car seats, this study has focused on reducing injury risk by improving seating on transport vehicles such as scoops, mantrips, personnel carriers, and rail-mounted locomotives. Similarly, proposed seat design improvements included layering of various types of viscoelastic foam padding to isolate vehicle occupants and operators from adverse health effects of jarring/jolting exposure. This paper discusses the results obtained from laboratory vibration testing of new seat padding materials in seven configurations at the NIOSH – Pittsburgh Research Laboratory (PRL) showing that careful use of configuration and materials can reduce the operator’s exposure to WBV. In addition, results for four mine transport vehicles are presented from field data collection efforts at a Southwestern Pennsylvania mine showing that a NIOSH based design does reduce the operator’s exposure to WBV. Finally, the authors discuss efforts to develop and field test seat padding interventions using new padding materials and configurations based on the laboratory test results [The findings and conclusions in this report are those of the author(s) and do not represent the views of the National Institute for Occupational Safety and Health.]

As part of the overall development of a borehole mining system -develop backfilling techniques to reduce possible damage to the environment caused by either the cavities or the piles of sand tailings that result from in-borehole hydraulic mining. (Borehole mining is described in technology News No. 56.) Approach Water filled cavities at the bottom of boreholes are rapidly and economically backfilled by jetting a sand-tailings slurry underwater. How It Works The Bureau of Mines borehole mining system creates a cavity having a horizontal diameter of more than 50 feet. (See photograph). This cavity can be backfilled with much of the sand previously mined. The sand tailings or similar backfill material is dumped by a front end loader into a slurry mixing tank at the rate of about one ton every three minutes. Enough water is added to give a thick, but pumpable slurry. The slurry is pumped down the borehole through a string of 4-inch pipe to the cavity. After passing through a right elbow at the end of the pipe, the slurry is ejected through a nozzle. The resulting slurry jet emplaces the backfill at the outer wall of the cavity. To obtain an even distribution of fill in the cavity, the nozzle is slowly rotated. The nozzle and pipe assembly is supported from a turntable on the surface. See diagram.

Jan 1, 1981

By Fred Garcia, William P. Diamond

As part of its mine safety and health research program, the National Institute for Occupational Safety and Health, Pittsburgh Research Laboratory, has been investigating the geologic and mining factors influencing methane gas emissions associated with longwall mining. A primary focus of this research has been the consequences of increasing longwall panel dimensions, particularly face width, in gassy coalbeds. Continuous longwall face emission monitoring studies were conducted at two adjacent mines operating in the Pocahontas No. 3 Coalbed, where longwall faces were to be extended from 229 to 305 m (750 to 1,000 ft). Average longwall pass methane emission rates for 229-m (750-ft) wide faces were 61 % (0.072 m31sec (153 cfm)) higher at the VP-3 Mine than at the VP-1 Mine. It was predicted by regression analysis of methane emissions data from 229-m (750-ft) wide faces that extending faces to 305 m (1,000 ft) would increase methane emission rates by only 7%, or 0.009 m3/sec (20 cfm), to 0.144 mysec (304 cfm) at the VP-1 Mine. In contrast, it was predicted that extending faces to 305 m (1,000 ft) at the VP-3 Mine would increase methane emissions by as much as 13%, or 0.029 mysec (65 cfm), to 0.268 dlsec (567 cfm).

Jan 10, 1999

By J. I. Paige

The Bureau of Mines, U.S. Department of the Interior, conducts research on methods to minimize the requirements for mineral raw materials through conservation, substitution, and increased minerals and metals recovery from primary and secondary domestic resources. To further this goal, prior Bureau research had devised a pyrometallurgical technique (RI 8211) to recover metallic iron and copper from molten copper smelting furnace slags by carbothermic reduction. As a continuation of this Bureau research on the carbon injection technique, reverberatory and electric furnace slags were treated in an 800-kva electric arc furnace by simultaneously top feeding the slag and injecting coke breeze into the bottom of the molten bath. Results of the tests show that 76 to 89 pct of the copper and up to 48 pct of the iron contained in the smelting furnace slag were recovered in a metallic phase, while 56 to 82 pct of the carbon injected into the sys-tem was utilized as the reductant. Laboratory evaluation has shown that the metal product, which contains 86.9 to 95.0 wt-pct iron and 1.3 to 4.1 wt-pct copper, will be a satisfactory medium for recovering copper by cementation. The expected energy consumption for the process is 110 kwhr per ton of molten slag treated and 520 kwhr per ton of solid material charged.

Jan 1, 1982

By S. J. Page, J. A. Organiscak

Dust surveys were conducted around small truck-mounted rock drills operating at surface coal mines to assess their airborne respirable dust generation and in-service dust control methods. Of four out of seven drills sampled, respirable dust concentrations measured around the drill deck ranged from 8.68 to 95.15 mg/m with3concentrations ranging from 1.37 to 2.69 mg/m at distances 12.2 to 30.5 m downwind of these drills. The other3drills had noticeably lower respirable dust concentrations measured around the drill deck at or below 1.30 mg/m.3Rotoclone-type dry dust collectors were commonly used with dust being emitted from around the drill deck shroud, collector exhaust, and collector fines dumping. Wind speed and direction was also a factor in the dust concentrations measured on the bench. Dust control modifications were made to three rock drills to reduce their airborne dust emissions. Dust controls tested included water injection into the Rotoclone exhaust, Rotoclone exhaust extension, improved sealing of drilling deck shroud, shrouding the Rotoclone hopper dump process, and wet drilling. These control techniques showed noticeable improvement in dust concentrations measured around the drills. Finally, repositioning the drill operator showed that operator dust exposure can be improved by avoiding the dust cloud around the drill.

Jan 1, 1995