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By Ryan G. Wade, Mark K. Larson

Researchers at the National Institute for Occupational Safety and Health (NIOSH) are studying the time-dependent response of rock stressed along weakness planes. The objective is to improve mine safety by reducing the number of roof falls that cause injuries and fatalities to miners. Researchers conducted direct-shear tests on mudstone overcored from a coal mine roof. Results showed frictional strength weakens with displacement along the shearing planes. Results of direct-shear creep tests also showed time-dependent deformation. Test measurements were compared to classical analytical models. Results suggest that risk of roof collapse may be decreased through layout and support design to reduce creep rates and thus slow progression toward failure.

By B. J. Scheiner

The Bureau of Mines investigated an electrolytic technique for oxidizing ore slurried with brine to determine its effectiveness in extracting mercury from ores. Optimization experiments conducted on a 2.5-pound-mercury-per-ton ore resulted in 95 percent mercury extraction with operating conditions of 30° C, a current density of 0.5 ampere per square inch, and a treatment time of 3.5 hours. Corresponding power consumption amounted to 38 kilowatt-hours per ton of ore treated. Amenability experiments conducted on 62 different mercury ores resulted in 90 to 99 percent mercury dissolution from ores containing 0.6 to 20 pounds of mercury per ton of ore, Treatment time ranged from 1 to 7 hours in 4-to 20-percent sodium chloride solutions, and power consumption was in the range of 10 to 50 kilowatt-hours per ton of ore. Pilot mill experiments in a 100-to 200-pound-per-hour extraction plant have been in progress to quantify power and reagent requirements. These data will be reported in a subsequent publication.

Jan 1, 1973

By Henry M. Harris, George J. Carter, Karle G. Strandberg

Field and laboratory data were used to evaluate sand from 13 dune locali­ties along the Oregon coast between Fort Stevens and Coos Bay. Chemical and glass button analyses were used to classify the 65 samples into two geographic zones of decreasing iron concentration. These data and data from preliminary magnetic beneficiation tests indicated that the sands between the Umpqua River and Coos Bay contained fewer impurities and were easier to beneficiate than those from areas between Fort Stevens and the Umpqua River, In addition, the dunes in the Umpqua River-Coos Bay area are the largest along the Oregon coast and are most favorably located with respect to existing transportation facilities, A composite sample from the Rogers area on the west bank of the North Slough of Coos Bay was selected for detailed mineral dressing studies. Splits from this sample were treated by various processes including magnetic separa­tion, electrostatic separation, flotation, and acid leaching, Concentrates were evaluated in glass melting tests. Test results indicated that high intensity magnetic separation followed by acid leaching would produce a prod­uct that could be used in amber and clear container glass batches. Good glass sand concentrates were also prepared by froth flotation; however, recovery was low and reagent costs excessive.

Jan 1, 1964

By J. A. Woolf, E. S. Leaver

"So much has been written on the ore genesis of the Comstock Lode that it is sufficient here to describe the original ores as consisting chiefly of quartz and some calcite. In places .a considerable amount of pyrite appears. There are also minor amounts of galena, chalcopyrite, and sphalerite. The chief silver minerals are argentite, stephanite, and polybasite. Most of the gold is native.A composite analysis of the ore; obtained by averaging the results from the early workings of several mines, is shown in Table 1.The history of the milling of Comstock ores from the Sixties to date would call for a review of the development of precious-metal metallurgy. However, a brief outline of the methods employed will give a better basis for studying the difficulties that are to be expected in the final retreatment of the remaining residues.In the earliest days amalgamation was accepted as the main recovery method. The large tonnage of good-grade ore from the Comstock Lode attracted the best metal¬lurgical talent of the time, and developed there the highest skill in the art of amalgamation, including the auxiliary chemical preparation of the ores and the mechanical operation. In this work 0.5 to 6.0 pounds of mercury was used per ton of ore. The mercury loss probably averaged about 2.5 pounds per ton of ore treated.The better grade of tailings from the ores milled were impounded, often near the works. In many cases the same tailings were reworked several times before they entered the main outflow waste stream that had its course down Seven Mile Canyon to the Carson River. Following the water course down the canyon the tailings were re-impounded, and, as metallurgy advanced, new plants were erected to again re-treat them. The most successful of these plants used forms of amalgamation, chlorination, and cyanide methods of recovery. To the tailings from the re-treatment plants, smaller tonnages of untreated tailings produced by milling lower-grade ore are continually added.Along the stream bed down the canyon a series of blanket runways were constructed, and each was independently operated for years by one or more men. All, or the heavier portion of the stream flow was run over the blankets; in this way much of the heavier concentrate, mostly sulphides and mercury products, was reclaimed.All of the residues finally entered the run-off flow along with the sewage from Virginia City and waste from the gas works, saw mills, and stables. As the accumulation along the bed of the Carson River became a nuisance, the lighter and better grade of tailings was impounded on the banks of the river. This accounts for the large dumps locally known as ""Douglas tailings."""

Jul 1, 1928

By G. R. Hopkins

A fifth successive record in fuel efficiency was established by the petroleum-refining industry in 1937, when the average heat requirement per barrel of crude oil run to stills was 554,000 3. t. u. compared with 597,000 3. t. u. in 1936. These amounts represent roughly the heat ob¬tainable from a tenth of a barrel of fuel oil (see fig. 1). Innumerable factors can influence heat consumption at rep aeries, but only a few are important enough to be subject to statistical measurement. Foremost of these is the ratio of runs to capacity, it being generally recognized that because the fuel demands for other than stills are comparatively inelastic, total heat requirements do not increase in direct proportion to crude runs. Thus, in 1937, crude runs of 1,183,440,000 barrels wore 11 percent higher than in 1930; the average operating ratio was 93 percent compared with 79 percent in 1936; but the total heat requirement increased only 4 percent, or from 639,149 billion 3. t. u. in 1936 to 663,966 billion 3. t. u. in 1937.

Jan 1, 1938

By W. E. Lamont

The Bureau of Mines studied the potential for recovering marketable barite concentrates that meet oil well drilling-mud specifications from barite waste pond materials. The study was conducted as part of the Bureau's research program to conserve domestic mineral resources by advancing mineral recycling technology. The samples, which were obtained from Georgia, Nevada, Missouri, and Illinois, varied widely in character, particle size, and barite content. The samples were treated with sodium silicate and sodium cetyl-stearyl sulfate and then floated. The flotation flow sheet was modified to reject barren coarse and/or slime fractions from some of the samples, and one feed sample had to be scrubbed prior to flotation to remove residual flotation reagents. The cleaned concentrates had BaS04 contents ranging from 95.0 to 96.9 percent, while recovering 81.2 to 95.9 percent of the barite in the feed. The specific gravity of each concentrate exceeded the specification of 4.2. The removal of the flotation reagents from the flotation concentrate was not investigated.

Jan 1, 1982

Reduce respirable dust in coal preparation plants. Approach Survey a representative sample of coal preparation plants to (1) determine the dust problem extent, and (2) evaluate practical dust control techniques. Coal Preparation Plant Survey Results Of the approximately 500 coal preparation plants in the United States, 18 plants in six States were randomly selected for survey. The raw hourly capacities ranged from 250 to 1,250 tons, plant construction dates ranged from 1037 to 1982, and usually 3 to 10 workers were required per shift. Respirable dust concentrations within the plants surveyed ranged from very good (less than 1 mg/m') to very poor (14 mg/m3). No relationship was found between dust concentration and age or capacity of the plant. Six of the plants had long-term respirable dust concentrations between 7 and 14 mg/m' in specific areas; six had concentrations between 2.2 and 6.2 mg/m'. Except fur ventilation, only 4 of the 18 plants used operating dust control devices with mixed effectiveness.

Jan 1, 1989

By W. R. Reed, J. A. Organiscak

Surface mining operations use large off-road haul trucks extensively to move material at mining properties. Past research, using the United States Environmental Protection Agency’s (EPA) emissions factors for unpaved haul roads, has shown that haul trucks generate the majority of dust emissions from surface mining sites, accounting for approximately 78%-97% of total dust emissions. Observations of dust emissions from haul trucks show that if the dust emissions are uncontrolled, they can be a safety hazard by impairing the operator’s visibility. This increases the probability for haul truck accidents (See photos above). However, the greatest long-term health hazard of dust generated from hauling operations is due to inhalation of the respirable dust [median diameter <4 micrometers (µm)] and tho­racic dust, which is equivalent to the EPA’s definition of PM10 [particulate matter with a median diameter <10 µm]. Exposure to respirable dust has long been considered a health hazard at surface mining operations, especially if silica dust is present. There are two legislative acts that regulate the air quality for mining operations: the Federal Coal Mine Health and Safety Act of 1969 and the Clean Air Act of 1970. The Federal Coal Mine Health and Safety Act of 1969 established the limits for dust in the work place for health and safety pur­poses. The Clean Air Act of 1970 regulates air emissions from facilities from an envi­ronmental perspective.

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 John N. Murphy

INTRODUCTION It&apos;s a pleasure to be with you and to share in this opening session for the 27th Annual Institute on Mining Health, Safety, and Research. As we come together today, I would like to reflect for just a moment on the past Institutes. Much of the progress in Mine Health and Safety has been the subject of presentations and discussions at the previous Institutes, either by identifying the issues, proposing techniques to overcome the problems, or presenting new findings to address important issues for the mining industry. May I extend our thanks and congratulations to VPI and to the planners and organizers of all of these Institutes for their important contributions to our mutual goals of improving mine health and safety. In the time available today, I would like to accomplish two objectives: 1. Provide my assessment of the current status of the Mining Health and Safety Research Program; and 2. Share with you the continuing need and direction for the Mining Health and Safety Research. MINE HEALTH AND SAFETY RESEARCH TODAY Mine Health and Safety Research was conducted formerly at four geographical locations within the USBM, specifically: Pittsburgh, Spokane, Twin Cities, and Denver. Although the Twin Cities and Denver Centers were closed, we have been successful in relocatingfmaintaining much of the core competencies to conduct a responsive, multidisciplinary Mine Health and Safety Research Program. The current staff at Pittsburgh and Spokane directed toward Health and Safety Research totals 397. This includes Lake Lynn Laboratory, about 60 miles southeast of Pittsburgh, where large-scale mine fire and explosion research is conducted. With the closures and transition we have worked to maintain the core competencies necessary to address Health and Safety issues. In some cases we have made arrangements to provide the requisite core competencies through external organizations. An example is the reassignment of the diesel particulate instrumentation capability from Twin Cities to Pittsburgh. In the case of diesel emission control, due to the cost of relocating laboratories and the lack of the appropriate skills at Pittsburgh or Spokane, the emission control laboratory is now being operated in Minneapolis through a contractual arrangement with the University of Minnesota.

By Marilyn Biviano

This analysis was conducted in response to concerns raised by the Chairman of the Energy and Commerce Committee of the House of Representatives regarding the potential impacts on automobile production, GNP and employment resulting from embargoing south African supplies of the platinum-group metals, platinum and rhodium. This analysis should be considered an addendum to the recently released study in which the direct economic costs I of a U.S. embargo on selected South African strategic and critical materials are estimated. The estimated GNP and employment impacts result when a mineral shortage reduces domestic industrial production and are in addition to the direct costs estimated in the earlier study. The direct cost to this nation resulting from a decision to embargo six south African strategic and critical minerals was estimated at $1.85 billion per year during the 1988-1992 time frame of the study. The estimated direct costs include the costs to U.S. consumers resulting from higher strategic mineral prices, lower mineral consumption, and the cost of developing higher cost domestic mineral resources. A large proportion of the costs result from the shortage of the two metals: platinum and rhodium. The average direct annual cost of embargoing is estimated at about $1.4 billion for platinum and $384 million for rhodium.2

Jan 1, 1988

Investigate and demonstrate the possible use of low-density (45-lb/ft3) foam concrete for sealing abandoned mine openings. Although foam concrete has been used in the construction and geotechnical industries for some decades, it has not previously been used in abandoned mint shaft reclamation. Approach A structurally engineered permanent mine seal for an abandoned coal mine shaft, using low-density foam concrete and other appropriate structural materials, was designed by the Bureau and installed at a field site, the No. 22 Shall in Logan County, West Virginia. The field work was conducted by a private contractor under the supervision of the West Virginia AML Section (WV AML). The No. 22 Shaft is situated immediately adjacent to a frequently traveled county road. A few feet away from the shaft is a branch of Pine Creek. On the other side of the road is a railroad line that serves an active mining operation located several miles away. The shaft measures 9.5 by 9 feet in cross section and is about 300 feet deep. It was used at one time as a dewatering shaft and still contains two large-diameter dewatering pipes with steel drive shafts inside them. Before the beginning of the reclamation, the shaft included a concrete cap a few inches thick, with a 4 - by 5-foot access opening covered by a steel grate. The interior of the shaft was quite humid, and considerable flow of water at the bottom of the shaft could clearly be heard from the surface. Methane tests were conducted that showed the shaft to be clear of this gas.

Jan 1, 1992

By Launa Mallett, Charles Vaught, Kathleen M. Kowalski, Michael J. Brnich

This paper reviews research from the U.S. mining community to define issues relevant to an evolving national and international workforce and to relate them to the emergency response population. The authors further explore and relate the key problem of an aging workforce to the resultant changes and emerging physical and psychological needs of emergency workers. Finally, the authors cite examples from the mine emergency response community and mine rescue experiences to suggest practical recommendations that include both organizational and individual formats, with an emphasis on new approaches to training this changing workforce.

By Doris M. Hyde

Belize remained relatively insulated from the social and political turmoil that affected some of its Central American neighbors. This stability encouraged new foreign investments that assisted the economy, especially in tourism and agriculture. The mineral sector remained limited to construction-oriented materials. Oil exploration efforts continued by the various concessionaires.

Jan 1, 1987

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

Le Système International d&apos;Unités (SI) Equivalents of Common United States Units Length 1 ft (foot) = 0.30480 m Mass 1 lbm (pound mass) = 0.45359 kg Force 1 lb (pound) = 4.4482 N Volume Flow Rate 1 cfm = 0.00047195 m3/s Mass Flow Rate 1 lbm/min = 0.00756 kg/s Pressure (wg at 20° C) 1 inch wg (water gage) = 248.36 N/m2 1 inch wg = 248.36 Pa Power 1 hp (horsepower) = 0.74570 kW Energy 1 ft-lb = 1.3558 J Torque 1 lb-inch = 0.11298 N-m

Jan 1, 1977

By Guy A. Johnson

The Bureau of Mines, through research and development contracts with the FMC Corp., has developed and tested two automatic fire sensing and suppression systems for large haulage vehicles. This work was undertaken to enhance driver safety, to lessen substantial property damage from fires on large trucks, and to increase the vehicle operator&apos;s productivity. One automatic truck fire protection system protects both the engine compartment and brake rid areas, while the second system is used in the engine compartment and fuel tank area. Both systems have redundant fire-sensing capabilities and manual override controls. Upon sensing flames and/or heat, the systems suppress the fire with pressurized, E-C class dry chemical. Both fire protection systems are flexible in design and can be modified for use on other mobile minim&apos; equipment. Prototypes of both systems have been subjected to long-term, in-mine validation testing and actual Eire tests on 100-ton-capacity trucks at the Puma mine near Tucson, Ariz., and the Erie mine at Hoyt Lakes, Minn. During the 4-month endurance test of the first system at Pima, an accidental flash fire occurred in the test truck while it was working in the pit. The system successfully sensed and automatically extinguished the fire with no injury to the driver and little damage to the truck.

Jan 1, 1975

By Jonisha P. Pollard

In restricted vertical working heights such as low-seam coal mines, workers are forced to assume kneeling or squatting postures to perform work. These postures are associated with increased risks for the development of significant knee pathologies such as meniscal tears and osteoarthritis [1,2]. Previous research has shown an increased applied knee flexion moment when kneeling in high flexion [3]. However, these models have been applied to flexion up to 140° which neglects contact between the thigh and calf which occurs beyond 140°. In addition to thigh-calf contact, when kneeling near full flexion there may be additional contact between the heel and gluteal muscles. To the author?s knowledge, heel-gluteus contact has not been previously investigated, but thigh-calf contact has received some attention. Zelle (2007) reported thigh-calf contact forces up to 30% bodyweight (BW) [4]. Zelle (2009) reported decreased knee forces when accounting for thigh-calf contact [5]. Caruntu (2003) reported erroneous force and moment estimations when thigh-calf contact was neglected from models to determine muscle and ligament forces [6]. In this study thigh-calf and heel-gluteus contact forces were quantified to determine their effect on the externally applied flexion moment in high flexion.

Jan 1, 2009

By R. V. Ramani, R. Stefanko, G. W. Luxbacher

1 11 THIS IS A DIGITAL COMPUTER SIMULATION OF THE VENTILATION SYSTEM OF A MINE LOCATED IN WASHINGTON COUNTY, PENNSYLVANIA. THE VENTILATION SURVEY ON WHICH THIS SIMULATION IS BASED WAS RUN FROM OCT. 27-30 AND ON NOV. 2, 1976 IN THE ACTIVE PORTION OF THE MINE INBY COMBS SHAFT AND THE INACTIVE PORTION AFCUND AND INBY DONAHOO SHAFT. FOLLOW-UP SURVEYS WERE THEN RUN ON FEB. 18 AND FROM MARCH 7-11, 1977 TO INCLUDE THE MAIN HAULAGE SYSTEM OUTBY THE ACTIVE AREAS OF THE MINE. ALL SURVEY DATA WAS NORMALIZED TO A STANDARD DENSITY OF 0.075 POUNDS/CU.FT. PRIOR TO THE CALCULATION OF INPUT PARAMETERS FOR THIS SIMULATION. IN THIS RUN, THE ORIGINAL VENTILATION NETWORK HAS BEEN MODIFIED TO INCLUDE NEW INTAKE AND RETURN AIRSHAFTS AT THE WEST MAINS.

Jan 1, 1977

By J. J. Rutledge, C. A. Herbert

In 1916 the Bureau of Mines, the Illinois Geological Survey, and the University of Illinois, working under a cooperative agreement, began an investigation of the subsidence of the surface above coal- mining operations. Four widely separated places where mining was in progress were selected as observation stations-one in the northern Illinois longwall coal field, one in the central Illinois coal field where the room-and- pillar panel system of mining was being used, and two in the thick coal of southern Illinois where the room-and-pillar panel system of mining was employed. These four points of mining activity gave a rather wide opportunity for observing subsidence when different mining methods were used. The four places selected for investigation of surface subsidence due to coal mining have been designated as follows: 1. A, in the northern Illinois longwall coal field, where the coal company has been operating coal mines for 30 years and where long- wall advancing is the only method of mining. 2. B, in Perry County, where panel room-and-pillar mining of a thick seam of coal is in progress and where mining operations have been carried on about 40 years. 3. C, in Macoupin County, where the coal company has been min- ing coal by the panel room-and-pillar method for about 15 years. 4. D, in Franklin County, where the coal company has been mining coal for 10 years by the panel room-and-pillar method.

Jan 1, 1927

By Paul W. Johnson

An evaluation is made of a lime-soda sinter process for extracting alumina from anorthosite. In this process, alumina is extracted by sintering anorthosite with soda ash and limestone, and then leaching the sinter with a dilute sodium carbonate solution. The sodium aluminate solution formed during leaching is separated from the residue and treated with lime in autoclaves to remove dissolved silica before alumina trihydrate is precipitated with carbon dioxide. The trihydrate is then calcined to a-alumina. Two options, dry grinding and wet grinding, are used in the sintering step. For plants producing 1,000 tons of alumina per day for the dry grinding and wet grinding options respectively, the estimated operating costs are $75.40 and $74.36 per ton of alumina when charging $2.50 per ton for anorthosite and $1.00 per ton for limestone. The fixed capital costs based on a Marshall and Stevens chemical equipment index of 240.0 are $76 million and $70 million respectively. This process, when treating anorthosite, using either option, is not competitive under current economic conditions with the Bayer process treating bauxite to produce alumina.

Jan 1, 1968