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By Thomas O. Glover

Economic growth in Cameroon was restricted during 1987 owing to a decrease in oil prices during 1986 and the falling value of the U.S. dollar in 1987. The yearly budget reflected a significant drop in revenues and expenditures. Taxes on imports and some exports were increased to raise revenue. The shortage of revenue and the lack of liquidity had restricted the Government's operation and investment programs. External public debt reached an estimated $3.2 billion2 representing 20.8% of the estimated gross domestic product (GDP) of $12.3 billion. The Government operated at an 8.7% deficit in 1987. The GDP growth rate had fallen to 0.6% in 1987 and was projected to remain at that level in 1988. Economic planning was managed through a series of 5-year plans. The country was under the sixth 5-year plan, which would end in 1991. The plan stressed balanced development as one method to achieve the desired results. Sources of revenue for the year included direct and allied taxes (31.2%). customs collections (31.0%), and oil royalties (28.0%). The oil royalties in 1987 represented a larger part of the budget than in 1986. In January 1987, eight African oil exporting countries, including Cameroon, formed the African Petroleum Producers' Association (APPA). The eight countries represented about 80% of the total African oil production of 5.2 million barrels per day in 1987. Cameroon produced approximately 3% of total APPA oil production. Cameroon produced approximately 68.5 million barrels of crude oil in 1987, slightly less than that produced in 1986. Earnings also fell in 1987, not only from falling production, but also from declining oil prices and the fall in value of the U.S. dollar on world markets. Oil had been the main stimulus of Cameroon's growth throughout the past decade, and a key generator of export revenues and financial resources for the economy.

Jan 1, 1989

By J. S. Miller

The Bureau of Mines developed methods for fragmenting oil shale formations with chemical explosives to prepare oil shale for in situ retorting and to evaluate the extent of fragmentation by various methods. A combination of three methods of explosive fracturing was used: (1) The displacement and detonation of a liquid chemical explosive in a natural fracture system; (2) the use of 60-pct dynamite to relieve stress conditions of the rock around the wellbore; and (3) the use of pelletized TNT in a series of wellbore shots to fragment the oil shale. An attempt was made to seal or grout around the fragmented oil shale by drilling 67 holes and displacing approximately 200 bbl of neat cement in the fracture system to aid in evaluating the fragmented zone, Seven methods were used to evaluate the fragmented zone. The primary evaluation method was the recording of seismic waves to delineate the fragmented zone with results based on arrival time, amplitude, and record quality. The data from the seismic measurements indicated the extensively fragmented zone was approximately 95 ft in diameter and 70 ft thick. Results of the remaining methods of evaluating fractures--including air-flow tests, coring of the expected fragmented zone, elevation measurements to indicate overburden crowning, and surveys with high-speed camera, impression packer, and down-hole camera--agreed with and supported the data from seismic measurements. Further field application of explosive fracturing and evaluation methods is needed to determine when a zone of oil shale is effectively fragmented for efficient in situ retorting.

Jan 1, 1973

By J. J. McClelland

The face area of a continuous mining section of an underground coal mine is the most dangerous area in the mine. Workers there are at nearly twice the risk of death or disabling injury than those in other parts of the mine. Despite extensive training of miners and improved equipment and mining methods, fatality and injury rates have remained constant during the past few years. It is clear that advanced research is necessary to apply new technologies or modify work procedures to solve the continuing problem of the loss of lives and the numerous injuries at the mine face. A U.S. Bureau of Mines research program is developing a Reduced-Exposure Mining System (REMS) that relocates personnel operating the mining machinery at the face away from the hazards to a nearby haven of relative safety. REMS uses proven technologies of computer automation and robotics to help control the operation of the mining machines at the face, while personnel are moved away to a nearby, safe control room. In this protected, well-ventilated environment, they can supervise or control the operation of the machines while viewing what the machine is doing using video and graphic display monitors.

Jan 1, 1994

By R. E. Peterson

The Bureau of Mines has undertaken beneficiation investigations on samples representing large tonnages of oxidized taconites in order to develop and evaluate applicable technologies for producing iron oxide concentrates from these materials. Information derived from the investigations can be used to expand the domestic iron raw-materials base and to establish the impact these taconite deposits may have on future national iron ore requirements. This report describes a process that includes reduction roasting, magnetic separation, and flotation, applied to a western Mesabi Range sample taken near Coleraine, Minn. The sample contained 36.8 pct iron and 46.0 pct silica, and represented 0.5 to 1 billion tons of crude material. An average of 962 lb/hr of oxidized taconite was roasted with 33 lb/hr of lignite in a 35-foot-long rotary kiln. Pulverized lignite and natural gas maintained the kiln operating temperature at 850° C. The roasted taconite was beneficiated by magnetic separation and cationic flotation in a 900-lb/hr pilot plant and yielded a product analyzing 68.1 pct iron and 4.6 pct silica, and containing 82.4 pct of the iron in the kiln feed.

Jan 1, 1981

By R. C. Ellman

Significant effects of in-the-mill drying on the pulverization of lignite have been demonstrated in an experimental impact-type pulverizer by tests in which the primary variable was temperature of the sweeping gas. Reduced power requirements and increased product fineness were observed as the extent of drying was increased. Pulverization with in-the-mill drying was studied in comparative tests made on lignites from five mine sources. All of the lignites responded in a similar manner; however, differences in the pulverizing characteristics of the individual lignites were significant. For one lignite, the electrical input to the pulverizer driving motor with minimum sweeping-gas temperature was 33.9 kw.-hr. per ton, decreasing to 14.0 kw.-hr. per ton with a sweeping-gas temperature of 658° F.; for another lignite, at comparable operating conditions, the electrical input decreased from 59.1 to 19.0 kw.-hr. per ton. In the first case, the product fineness increased with increasing temperature of sweeping gas from 65.2 to 79.2 percent through a 200-mesh sieve, and in the second case, from b2.7 to 86.9 percent through a 200-mesh sieve.

Jan 1, 1962

By J. L. Konchesky

The Bureau of Mines investigated the main reactions of coal gasification by steam. Coal (minus 70 mesh)-water slurry was fed into a laboratory-scale tube-coil gasifier to show the effects of residence time (0.6 to 1.1 sec), steam-coal ratio (2 to 11 lb/lb), and temperature (1,600° to 1,900° F), at about 1-1/2 atm. Devolatilization was very rapid: about 70 pet of the coal's volatile matter evolved within 0.6 sec at about 1,600° F; temperature increase released more volatile matter until about 95 pct was driven off at 1,900° F. Devolatilization was not increased by raising steam concentrations and was only slightly increased by extending residence time. Below 1,600° F, there was no evidence of the steam-carbon reaction, but at 1,900° F more than 80 pct of the coal's fixed carbon was converted to gas; the conversion was increased Significantly by increased residence time and steam concentration. The water-gas shift reaction occurred simultaneously with devolatilization, and the steam-carbon reaction, approaching equilibrium with increased residence time and temperature and decreased steam concentration.

Jan 1, 1969

By David Hoadley, Kenneth R. Maser, Ashok B. Boghani, James E. Billar, D. Randolph Berry, Mackenzie Burnett, Robert H. Trent

APPENDIX E REFUGE CHAMBER LEAKAGE ANALYSIS Variables p 0 = pressure outside of chamber (psi) p = pressure inside of chamber (psi) i q = leakage flow rate (cfm) Q= total leakage (cu ft. ) V = chamber volume (cu. ft. ) t= time (minutes) p= nominal atmospheric reference pressure (14.7 psi) A= chamber leakage coefficient (cfm/psi)

Jan 1, 1976

By A. F. Colombo

The differences in starch demands that occur between ore types subjected to the anionic flotation of calcium-activated gangue and the range of residual starch contents expected to give the best metallurgical response were evaluated on a batch and pilot plant basis. A phenol-sulfuric acid procedure was used to determine the residual starch contents in the iron ore slurries. Three pilot plant campaigns on samples from a group of orebodies located on the Marquette Range, Michigan, showed that in the rougher flotation pulp, a residual carbohydrate range of 25 to 100 ppm, which is equivalent to a 0.5 lb/short ton spread, yielded optimum results. Concentrates assaying at least 65 percent Fe and less than 6.0 percent Si02 and containing more than 80 per-cent of the recoverable iron were obtained. Both batch and continuous research indicated that starch adsorption is calcium modulated and relatively unaffected by the fatty acid additions employed. Residual starch level is not a critical control factor, but the addition must be adjusted to about 25 to 50 ppm residual starch and kept in this range for most effective utilization of the added tapioca flour.

Jan 1, 1972

By John C. Edwards, Gerald S. Morrow

The U.S. Bureau of Mines has constructed a smoke chamber and developed sensitivity tests for smoke detectors. Response of ionization- and optical-type commercially available smoke detectors have been investigated. Six smoke detectors were measured with respect to visually obscuring smoke characterized by a corresponding optical density for smoldering and flaming coal combustion in the smoke chamber. It was determined that for one type of ionization smoke detector the alarm time was nearly equivalent to that of an odor monitor's alarm for smoldering coal combustion experiments and earlier for flaming coal combustion experiments. The experiments showed that an average CO concentration of 5 ppm corresponded to an optical density of 0.022 m for smoldering and flaming-1coal combustion. Two of the commercially available ionization-type smoke detectors were more responsive to flaming than smoldering coal combustion at an optical density of 0.022 m, whereas the optical smoke detectors-1showed the opposite trend. The responsive characteristics of the detectors evaluated with respect to known smoke conditions in the smoke chamber shows their potential for use as mine fire sensors or part of a mine atmospheric monitoring system to improve mine safety.

Jan 1, 1995

By T. H. Ueng, S. D. Thompson, Y. J. Wang

"INTRODUCTION There is considerable interest among mining companies and government agencies in finding new ways to better control the generation and dispersion of respirable coal dust. Mine face ventilation generally has been considered the most effective tool for lowering the concentrations of respirable dust. Unfortunately, only a limited amount of research has been undertaken to determine the relationships between the various face ventilation parameters and the dust dispersion patterns. If these relationships are better understood, more efficient ways of controlling dust generation and dispersion may be evident. A practical and economical way to gain more in-depth understanding of these relationships is through scale model studies. Past research (1, 2, 3) has substantiated that scale models can be successfully used in studying ventilation parameters provided that geometric, kinematic, and dynamic similitues are maintained.To better understand face ventilation and dust dispersion, a 1/5 scale model of a room and pillar mine face was constructed. This model was then used to conduct experimental investigation on two ventilation parameters. The parameters which were examined included the line brattice distance from the face and the air quantity. The values of these parameters were varied in numerous experiments so that their effects on dust dispersion could be measured.The experiments of this study were unique in two ways. First, two different substances were used to simulate the respirable coal dust. Experiments were conducted with a tracer gas (methane) serving as the dust simulation substance. They were then repeated with an actual dust (Arizona Road Dust) replacing methane as the dust simulating substance. In previous scale model studies, the dust simulation substance has been limited to a tracer gas such as methane. The validity of using a tracer gas to simulate dust is not clearly known yet. The other unique thing about these experiments was the number of points measured for methane (or dust) concentration. Twenty five points in a grid arrangement were measured during each experiment of this study. The number of points measured for concentration in other studies has generally been very small (often 2 to 4). It is difficult to obtain an accurate assessment of how the ventilation parameters are effecting dust dispersion when only 2 to 4 points are measured."

Mar 1, 1989

By Richard H. Oitto

The Bureau of Mines and Kaiser Steel Corp. are cooperatively demonstrating the shield-type roof support in the longwall mining of a 10-1/2-foot-thick coalbed at the York Canyon mine near Raton, N. Mex. As part of the demonstration, the shield supports have been successfully and safely moved from a completed longwall face to a new longwall face. When these shield supports were moved in 1976 they were the largest and heaviest longwall support system in use in U.S. coal mines. The shield supports, preparations for moving them, the actual move, and installation at the new face are discussed.

Jan 1, 1977

By Oliver Bowles

"On January 15, 1920, a man was killed by falling rock in a limestone quarry in Utah. The quarry was visited by a representative of the Industrial Commission of Utah, and the cause of the accident was carefully determined. The facts concerning this accident, and the Bureau's recommendations are given in some detail in the hope that quarrymen may take steps to avoid dangerous conditions, and thus reduce the liability to accident.The quarry rock face is about 180 feet high, and the ledge is covered with an overburden of 10 to 20 feet of a mixture of soil and boulder. The rock face is not vertical, but slants at an angle of about 45° to 50°, while the stripping face stands at an angle of 70° to 80°, and is flush with the rock face. The conditions are shown in the accompanying figure.The man who was killed was employed in breaking rock at the foot of the quarry wall. The sun had been shining on the stripping face, and probably thawed the ground, thus permitting a boulder weighing about 20 pounds to roll down the slanting face, striking, and almost instantly killing the workman who was working with his back toward the quarry wall."

Feb 1, 1920

By MINE SAFETY BOARD

Mining men and State officials in their efforts to make mining safer are confronted from time to time with various complicated questions . In recent years these questions have tended to increase with the use of mechanical appliances . Twenty- five years ago electricity had hardly been used in mines . To - day there is practically no other means of transmitting energy employed in American mines , except in metal mines where compressed air is used for drilling and small portable hoists . In coal mines there has been a great increase in mechanization and a tendency to concentrate the mining operation , which has resulted in attendant difficulties of ventilation and intensified the danger of coal - dust in bituminous and subbituminous mines . The answers to the questions which arise are not always obvious , as is shown by the fact that even able , experienced men differ in their opinions of the best way to overcome a particular mining hazard . The difficulty of solving the problems is also demonstrated by the varying provisions in mine safety rules , both in official regulations and in the rules of those mining companies who are endeavoring to do more than follow the letter of the law . Such differences of opinion on the best way of meeting a difficult situation are often found within the same mining organization , and so it has been in the Bureau of Mines .

Nov 1, 1929

By E. A. Johnson

The Bureau of Mines determined the effect of slag additives on copper solubility and the distribution of typical impurity elements between copper-iron matte and silica-saturated fayalite (iron silicate) slag at 1,250° C, in CO-C02 atmospheres with 10 pct S02. Small additions (<10 pct) of CaO, A1203, and MgO were shown to decrease copper solubility in the slag. Additions of CaO increased the solubility of bismuth and tellurium in the slag at lower matte grades. Al 20 3 additions decreased the solubility of tellurium in the slag, but increased that of bismuth at lower matte grades. A1203 additions increased the solubility of both at lower matte grades. With no additives, matte grade did not affect the solubility of bismuth in slags, whereas increasing matte grade resulted in a slight decrease in the tellurium solubility in slag. Additions of CaO or As203 alone to the slag increased selenium and arsenic distribution to the matte, but had little or no effect on antimony distribution; whereas small MgO additions decreased antimony distribution to the matte but had no effect on arsenic and selenium distribution. In slags with no additives, selenium distribution to the matte increased with increasing matte grade, while that of antimony decreased and that of arsenic was unchanged.

Jan 1, 1982

By Scott Turner

We have gathered here in Pittsburgh, at the Nineteenth Annual Safety Congress, in order to advance the cause of safety. Although my personal inter- est in this movement alone would have been sufficient to insure my presence here, specific duties have brought me to this congress in an official capacity. My responsibilities as the head of the United States Bureau of Mines, which makes the promotion of safety in the mining and allied industries one of its major activities, necessitate my attendance. My obligations as president of the Joseph A. Holmes Safety Association, an organization closely allied to the bureau and which has for one of its objects the public recognition and reward of outstanding achievements in safety, require that I be present at this time. This morning, my job is to present the Holmes safety certificate to the Lehigh Portland Cement Co. First I will remind you of the nature of this safety association, named in honor of the first director of the Bureau of Mines. After Mr. Holmes' death, in 1915, a number of influential organizations banded together to commemorate the name of this leader in the conservation of human life. In order that you may appreciate the significance of an award by the Joseph A. Holmes Safety Association, I shall read you the list of 24 national organizations which comprise its membership:

Nov 1, 1930

By C. J. Tsai, K. Olson, R. Caldow, B. Cantrell, D. Y. H. Pui

"The amount of new surface generated during fracture of coal and rock has been found to be directly proportional to the amount of energy associated with the fracture. To confirm this and study the relationship between the new surface and dust aerosol formed during fracture the U.S. Bureau of Mines has conducted experiments simulating fracture as it occurs during mining operations. In addition to fracture energy measurements, fragment size distributions were determined over a size range of 0.5 microns to 1 inch. The size measurements accounted for the mass of the original sample. Material studied was Illinois No. 6 coal, coal from the Taggart seam of Virginia, and Berea Sandstone. The experiment employed a chamber to confine sample fracture fragments and a combination of particle measuring systems. Measurement methods included aerodynamic sizing of entrained dust aerosol with optical counting and inertial impaction techniques, physical sizing of non-airborne fragments by sieving, and optical sizing of fragments using optical counting and scanning electron microscopy.Results of the size distribution measurements indicate that a single size distribution law does not adequately describe the entire fragment size distribution. A combination of two size distribution laws, such as the Rosin-Rammler and log-normal distributions, is necessary to describe this size distribution. New surface created by fracture is primarily found 1n particulate between 2 and 50 microns with a mean size of 12 microns. It was also found that the specific energy of fracture can be linearly related to the .concentration of aerosol in this size range.Only a .small fraction, between 0.02 and 0.16 pct, of the dust particles .generated during fracture become airborne. The remainder adhere to the new surface created during fracture. An entrainment fraction, defined for a given particle size is the ratio of the mass of airborne particles to that of the total particles generated in that size range. This experimental entrainment fraction is related to the Van der Waals adhesion force and the electrostatic forces operating during fracture and is used to derive empirical parameters to model these two processes."

Jan 1, 1989

By J. G. Groetsch

Municipal solid waste (MSW) contains about 8 pct ferrous metallics which are typically recovered by magnetic separators at municipal resource recovery facilities. At present, only about 182,500 ton/yr of MSW magnetics are being processed for remelting out of the -11 MMton/yr discarded. The Sn in MSW magnetics makes this material difficult to recycle to the steel industry. An electrolytic method using stannous fluoborate-fluoboric acid for removing Sn was evaluated. Optimum concentrations of Sn in the electrolyte and the effects of impurities such as Al, Cu, Fe, and Pb on Sn recovery from the electrolyte were evaluated. Operating parameters were 60 to 65 g/L fluoboric acid and 20 to 30 g/L Sn using stainless steel electrodes at a current density of 60 A/ft2. Tests were run with postconsumer scrap4 and MSW magnetics. Good results were obtained with MSW magnetics, after pretreatment by reshredding, magnetic separation, and air classification. The final Sn content of the MSW magnetics scrap was reduced from 0.18 to 0.017 wt pct. Heat and solvent treatments of MSW magnetics to remove coatings did not significantly improve Sn removal beyond that obtained with material that had only been mechanically pretreated.

Jan 1, 1984

By Bernard Porter

Differential thermal analysis measurements were used to characterize a phase diagram for the KF-YF3 system at compositions between 0 and 60 mole-percent YF3, A congruent melting point compound was observed at 3KF?YFs, which melts at 965°±5° C and undergoes solid-state decomposition at 725°±2° C. One peritectic and two eutectic compositions are also defined.

Jan 1, 1969

By John J. Gordon, Neil K. Cooperrider

The International Organization for Standardization (ISO) standard ISO 2631 [1,2] provides three methods for evaluation of human exposure to vibrations that contain occasional shocks or impacts. Part 1 of the standard specifies the running r.m.s. or maximum transient vibration method (MTVV) and the fourth power vibration dose value (VDV). Part 5 of the standard provides a method of computing the stress in the lumbar spine for humans exposed to multiple shocks. Alem et al [3] have reported application of these methods to data for tactical ground vehicles. This paper reports and compares VDV and spinal stress evaluations of more than 90 hours of vibration and shock measurements on North American locomotives engaged in through freight operations. The measurements evaluated in this paper were obtained for full crew shifts on 19 freight locomotive runs on mainline track in locations from New York to California. The shifts ranged in duration from 187 minutes to 497 minutes. The average speeds for the shifts were from 21.0 mph to 54.6 mph. All measurements were made on locomotives hauling freight trains in regular revenue service. Data Acquisition and Processing The results reported here were computed using test data acquired from a tri-axial seat pad, accelerometer at a sample rate of 400Hz with an anti-aliasing filter corner frequency of 100Hz. The VDVs and the lateral and longitudinal spinal stress values were computed directly from the acquired test data according to the procedures specified and described in [1] and [2]. The vertical spinal stress values were computed by converting the as-acquired test data to a sample rate of 160Hz for input to the vertical spine model, as required in [2]. The conversion of the test data from the as-acquired sample rate of 400Hz to the required sample rate of 160Hz involved up sampling or interpolating the test data to an equivalent sample rate of 800Hz, band limiting the resultant data with a low-pass filter corner frequency of 60Hz and finally down sampling or decimating the 800Hz data to a sample frequency of 160Hz. Discussion The vertical VDVs computed according to [2] for the 19 shifts ranged from 2.68 to 9.33 m/s1.75. In all but one case, the vertical values were greater than the values for the lateral or longitudinal directions. Note that the health guidance in [1] puts the lower boundary of the health guidance caution zone at a VDV value of 8.5 and the upper boundary at 17 m/s1.75. The daily equivalent static compression dose computed following [2] ranged from 0.123 to 0.434 MPa. Health guidance provided in [2] states that there is a low probability of an adverse health effect if the daily dose is less than 0.5 MPa.

Jan 6, 2006

By H. V. Makar

Ferrous fractions recovered from urban refuse were tested by the U.S. Bureau of Mines, U.S. Department of the Interior, to assess their character, composition, and response to detinning. This research was conducted as part of the Bureau&apos;s efforts to conserve mineral resources through recycling. Results from laboratory and pilot tests established the technical feasibility of detinning scrap recovered from household refuse. Detinned products containing less than 0.03 percent residual tin were produced. Concentrations of other residual elements were determined and compared with those in existing commercial scrap and with proposed specifications. Studies were extended to include evaluations under commercial conditions. Preliminary tests on 5-1/2 short tons of ferrous fractions revealed the need for adequate shredding to assure optimum metal-caustic contact and minimize retention of tin-bearing rinse solution in the final baled product. The washed metal product contained 0.06 percent residual tin. Subsequent melt tests showed a yield of 91.4 percent. Based on these studies, full-scale commercial demonstrations showed that detinned bales equivalent to No. 1 bundle quality could be produced in carload quantities. Up to 5 pounds of tin per gross ton of scrap is con-currently recovered during the detinning process.

Jan 1, 1979