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By Donald P. Blasko

The Bureau of Mines investigated two areas of the Alaska Peninsula (Iniskin Peninsula and Becharof Lake) where oil and gas seeps were known to occur in an attempt to determine (1) whether the seeps are still active and (2) the amount of bitumen contained in the drainage leaving the seeps. The oil, water, oil-water, and gas were sampled and analyzed. Five oil samples were taken at four separate oil seeps. The API gravity of the oil ranged from 8.90 to 21.4° and the sulfur content was from 0.12 percent to 0.59 percent. Four of the seven gas samples obtained revealed fair high caloric values, ranging from 746 to 924 Btu/cu ft, and averaging 844 Btu/cu ft. A total of 44 water samples were obtained. Bitumen content of the water was as high as 9,773 mg/l at one seep, but averaged less than I mg/l where the seep-drainage water entered the Gulf of Alaska.

Jan 1, 1976

By Edward F. Divers

This Bureau of Mines report describes the results of a laboratory evaluation of a mechanical dust collector (scrubber) suitable for underground coal mine applications. The scrubber was developed by the Mining Research and Development Establishment (MRDE), Burton-On-Trent, England, and tested at the Federal Bureau of Mines Pittsburgh Mining and Safety Research Center (PMSRC) as part of a technical exchange program. The 2,000-to 8,000-cfm scrubber consists of an electrically driven vane-axial fan, water sprays, a 2-inch-thick, loosely knit stainless steel wire filter panel, and a water droplet eliminator. The dust-laden airstream travels axially through the water sprays, filter panel, fan, and eliminator. Slurry water drains into a sump where a belt-driven pump recycles it back to the sprays. Fractional breakdown tests showed a maximum dust collection efficiency of 96 pct of respirable-size (less than 7 µm) coal dust at 2,000 cfm. Efficiency decreased with increasing airflow; for example, to 91 pct at 6,000 cfm. Following these tests the original wire panel was replaced with a similar but smaller panel developed at PMSRC; a marked improvement in dust collection occurred at all flow rates tested.

Jan 1, 1979

By Douglas Pollock, John Organiscak

Water spray characteristics, including droplet size and velocity, airborne dust capture potential, and induced airflow quantity for various spray nozzle designs were evaluated to provide basic information for improving spray applications. Water droplet size and velocity characteristics were initially measured by a Phase Doppler Particle Analyzer (PDPA) for hollow cone, full cone, flat fan, and air atomized spray nozzles at similar operating parameters. Airflow inducement and dust capture experiments were also conducted under the same operating parameters to examine any salient features of the spray nozzle type, droplet characteristics, induced airflow, and airborne dust capture. Test results indicate that there are trade offs between airflow inducement and dust capture efficiency. A spray nozzle with a wider discharge angle was observed to induce more airflow, but at reduced dust capture efficiencies. Increasing spray nozzle fluid pressure(s) generally reduced water droplet sizes with concurrent increases in droplet velocity, airflow inducement, and airborne dust capture. Placing a three-sided barrier around the spray nozzles normally reduced spray air induction and increased dust capture efficiency. A direct relationship between airborne dust capture efficiency and spray input power normalized per unit of airflow induced was ob- served. This information can be utilized to improve the performance of water sprays for reducing airborne dust levels.

By F. C. Turin, K. M. Cornelius

1. INTRODUCTION Underground mining in the USA has undergone significant change in the past 20 years. Two key elements have been increased mechanization and a more educated work force in spite of these changes, many jobs continue to be labor-intensive and repetitive in nature. They entail tasks that, performed over time, can take a toll on the soft tissues and joints. The problem may be compounded by an aging mining workforce. In 1996 the mean age of the coal mining work force was 45 years and the median total years of experience was 20 (NMA 1998). As a person ages, the body resilience to chronic wear and. tear is reduced, which may cause a worker to pay an increasingly higher health price for performing the same task. Mining companies, like many others, are becoming more aware of cumulative effects to the worker as reports of these types of injuries rise. . Conducting a job analysis is an important step when considering a job redesign or modification to reduce worker cumulative trauma exposure. A basic approach to job analysis is to examine the types of aches and pains reported, the tasks performed, and work site conditions. The US Bureau of Mines (USBM) conducted an evaluation of roof-bolting tasks performed at an underground coal mine concerned about early warning signs of cumulative trauma. This evaluation will comprise the primary focus of this chapter. The approach used for the roof-bolting race study may be applicable to other work environments. 2. FUNDAMENTALS OF CUMULATIVE TRAUMA EXPOSURE Musculoskeletal injury is a term used to describe a wide range of soft tissue disorders that affect the nerves, tendons and muscles. Common examples include lower back pain, tendinitis and carpal tunnel syndrome. The majority of these injuries are not the result of sudden mishaps, but usually develop gradually from repeated wear and tear. Symptoms may not appear immediately, but can take weeks, months or even years. Symptoms may result from many types of activities, performed at work or at home, and it is often difficult to attribute there to a single event. In fact, it is more common to identify factors that may have contributed to the development of the condition. The terms repetitive strain injuries or cumulative trauma disorders (CTD) have been commonly used to refer to disorders that have occurred due to work-related activities. Three main risk factors contribute to CTD: force, repetition and awkward postures. Any one or combination of these may contribute to the development of CTD. Therefore, the design of equipment in conjunction with the required tasks should be evaluated when attempting to reduce these risk factors. Examining the layout of the work area to help identify tasks which may contribute to cumulative trauma is necessary. The following list (Putz-Anderson 1988), describes ergonomics concerns that, overall, should be minimized at the work area: • Crowding or cramping the worker: a work area layout may unnecessarily constrain movements of the worker. • Twisting or turning: placement of tools and materials may require the worker to twist the spine to fulfill the requirements of the job. • Repeated reaching motions: the layout of the work area may require the worker to lean to reach and grasp the necessary tools and controls. • Misalignment of body parts: the arrangement of the work area may require the worker to frequently have one shoulder higher than the other or have the neck or spine bent to one side. While many of these concerns are a function of equipment design and environmental conditions, making workers aware of these issues may help them to adapt their work habits to reduce risk of injury. Additionally, this information is useful when conducting an ergonomic evaluation of a work area and associated tasks. 3. UNDERGROUND MINING ENVIRONMENT The underground mining environment is a unique challenge. It is more difficult to develop controls for an underground mine site as compared with a factory setting where equipment and facilities can be more easily designed to reduce worker force, posture and repetitive exposures. In an underground mine, workers are required to perform labor-intensive tasks that often cannot be avoided due to environment constraints. The dynamic nature of the environment does not allow easy implementation of mechanical assists to reduce force exposure. Many of the tasks performed by workers are repetitive. Restrictive work areas due to low ceiling height, low lighting levels and large pieces of equipment cause workers to perform these tasks in postures that are not desirable. Designers of underground mining equipment can control how a machine will function, but not the environment in which it will be used, Hazards in an underground mine cannot be completely removed by redesigning the system. There are many hazards and information sources that must be continually monitored by workers including their position in relation to large pieces of mobile equipment and unpredictable geological anomalies. Thus, immediate dangers may take priority over awareness of ergonomic considerations while performing a job. 4. CASE STUDY OF UNDERGROUND ROOF BOLTING 4.1. Roof Bolting and Cumulative Trauma Exposure In an underground coalmine, after an area is mined it is necessary to support the roof to keep it from collapsing. Since 1950, the primary method for supporting the mine roof has been installation of roof bolts. Long bolts installed into the roof compress the layers of strata achieving a uniformly distributed support anchorage. Roof bolts, typically 6-8 feet long, are installed by workers using large roof-bolting machines. There are different types of machines

By Steven J. Page

The increasing concern for dust control in metal and nonmetal mining prompted the Bureau of Mines to investigate the dust suppression technique of applying foam to face drilling and crushing operations in a gypsum mine. Previous attempts to achieve effective dust control on face drills using other techniques, such as dry collection and boom-mounted external water sprays, have not been successful. The largest problem encountered when using these techniques has been the lack of sustained effective dust capture at the face. Results for two faces studied show that injection of the foam through the drill steel provided an average dust reduction of 95 pct. However, the use of foam on the crusher provided no dust reduction at the operator's position (because of good local ventilation) and a 27-pct average dust reduction for two feeders at a location midway between the operator's position and the crusher. The conclusions obtained from this study are that (1) foam can be an extremely effective dust control technique when used on localized dust sources such as dust generated within a blasthole during drilling, and (2) foam is only marginally effective as a dust control technique when used on nonlocalized dust sources such as dust generated during the feeder-and-crusher operation.

Jan 1, 1981

By R. V. Ramani, C. B. Manula, A. Owili-Eger

All the input data required to operate the ventilation simulator except the title, are entered with I10 or F1O.O formats. The title is entered with (18A4) format. The floating points are designated by (FP) and the integers by (INT)

Jan 1, 1975

By M. J. Mechling, H. J. Sloman

"INTRODUCTION The practicability of providing safer methods of underground transportation for mine workmen and officials has been given much thought in recent years by advocates of safety, health, and efficiency. Numerous fatalities and disabling injuries have resulted from falls of roof, personal contact with both electrical conductors and timbers, and improper position or conduct of passengers while they were in man-trips. Unfavorable atmospheric exposure due to high-velocity air currents and sudden changes in temperature may adversely affect the health of mine workmen if they are transported in unprotected Cars.Recent installations of protected, man-trip cars, wherein such cars have been utilized to haul men to and from their stations in the mine, have resulted in a large reduction in exposure to haulage hazards. Where special-built cars have been used the probability of injuries to men during transportation in man-trips has been decreased, and the morale of the employees has been improved. A number of man-trip cars are in service in the coal-mining regions; those described in this circular are typical examples of a few installations."

Feb 1, 1948

By I. F. Avery

The productive capacity of natural gasoline and cycling plants in the United States on January 1, 1956, was 46.4 million gallons per day, according to the Bureau of Mines r biennial survey. This represented an increase of 7.5 million gallons per day during the 2-year period since the last survey. In the previous 2-year period, capacity had increased 4.8 million gallons per day. The total number of plants increased from 535 on January 1, 1954, to 600 on January 1, 1956. The number of operating plants increased from 521 to 571. The major increases in capacity took place in Texas, Louisiana, New Mexico, and Oklahoma, with the absorption-type plants continuing to account for most of the new construction. All types of plants showed construction gains with the exception of charcoal which showed no change.

Jan 1, 1957

By Seth T. Reese

THE INJURY DATA and experience at coal mines in the United States for 1949 are presented in this bulletin under the following general sections: (A) General injury experience: 1. Tables 1 to 18: Contain current data on number of injuries, frequency and severity rates, average severity, major and detailed classification of injuries, and distribution of injuries by part of body and days of disability. (B) Selected injury data: 1. Tables 19 to 45: Contain current data on various selected factors of operating and physical conditions that have contributed to the occurrence of injuries at bituminous-coal mines and at Pennsylvania anthracite mines. These factors include extent of hand and machine loading, size of mine by employment and production groups, thickness of bed or vein, duration and number of work shifts, rating of mine as gassy, or nongassy, types of cap lamps, ventilation, explosives, and combinations of factors that influence mine safety. (C) Injury experience by States: 1. Tables 46 to 52: Contain current data on number of injuries, frequency and severity rates, average severity, major classification of injuries, injury records of all coal-producing States, and the relative standing by States according to injury-frequency rates. (D) Appendix: 1. Tables 53 to 84: Contain certain injury data for the years 1943-47 that have not heretofore been published.

Jan 1, 1953

Objective The Roof Monitoring Safety System (RMSS) introduced in late 1997 has been modified to improve durability and simplify the measurement of roof or rib movement. Greater knowledge about the stability of roof and rib rock is expected to result in safer conditions for the mine worker. Developing safety tools and methods is part of NIOSH’s mission to prevent work-related injuries and deaths. A mining operation with a proactive approach to control ground conditions is in a better position to make essential decisions related to mine development and take remedial actions in the event of roof or rib falls. Background Falls of roof and rib rock are responsible for a high number of mining injuries and fatalities. Visual inspections and the sound of breaking or cracking rock are often relied upon to indicate unstable roof conditions. In some mines, roof bolt holes are checked for separations or gaps with a scratch tool; this provides further information about the stability of the roof. Monitors such as the RMSS can confirm roof movement that was surmised only by eye or ear, or potentially detect movement that was missed by these means. An added safety feature of the RMSS is the capability of locating the measurement station at a distance from where the monitor is placed in the roof. Stone and other industrial minerals are the basic raw materials used in construction. Continuing economic growth and highway building in particular have resulted in record demand for stone. This high demand, coupled with increasing constraints on surface mining, have caused many new .stone mines to be developed underground. These new mines in many cases will be employing new and inexperienced miners who may not have the ability to recognize unstable ground conditions using just sight and sound. From the initial introduction of the RMSS, it was anticipated that improvements and modifications would occur. We correctly assumed that, once miners and operators were exposed to the monitor. suggestions would evolve. In one case, the RMSS was used to determine if an impact hammer used on a mechanical scaler was causing or increasing instability of roof rock. The test showed little or no movement resulted from this method of scaling under the particular conditions at this mine. In another instance, an operator needed to monitor the stability of a haulageway in a benched area that had very high ceilings. The RMSS requires a 2-in-diam hole for installation; however, only roof bolt holes of a smaller diameter were available, and it was impossible to drill new holes at that height. The operator attached a 2-in collar at the roof line and was able to install the RMSS. These examples illustrate the potential for further enhancements of this tool. NIOSH continues to seek additional candidate sites for use of the RMSS. As usage increases, we anticipate further refinements and additional applications. How It Works The new RMSS, cable, and read-out box are shown in figure l. The read-out box has a green light that stays lit until battery replacement is needed. The red light is set to conic on to indicate movement. The red light can he adjusted for increments of movement as low as .010 inches. The RMSS red light should not be considered as a primary or only means of determining the stability of roof or rib. It is intended to serve as a supplement to existing methods or procedures for determining stability. An overview of the RMSS components is shown in figure 2. Movement of rock layers within the mine roof is measured relative to a fixed-point calibration at the monitor housing. The housing contains a potentiometer, roller switch, plastic rack, and spur gear. Movements are detected by the transfer of electromotive forces

Jun 1, 2000

By G. E. McClelland

During the 1970?s, the Bureau of Mines investigated a particle agglomeration technique for improving the flow of leaching solution through heaps of clayey or crushed, low-grade gold-silver ores, wastes, and tailings. This technology has been adopted on a broad scale by the precious-metals-processing industry. This report presents information on five commercial operations that have benefited from agglomeration technology and that represent a cross section of current heap leaching practice. The technology is cost effective because of decreased leach times and improved precious metal recoveries.

Jan 1, 1983

By Joseph A. DeCarlo

This publication presents a tabulation of oven-and beehive-coke plants in the United States as well as pertinent statistical data on the number, capacity, kind, and condition of the ovens. Coke is an essential raw material in pig-iron production and the coal-chemical materials produced during the coking process are used to manufacture fertilizers, paints, plastics, synthetic rubber, explosives, detergents, nylon, pharmaceuticals; dyes, and many other products necessary to the high standard of living in this country. In table 1 the oven-coke plants in the United States on December 31, 1959, are listed. This shows the name and address of the company, location of the plant, classification or type of ownership, number and type of ovens, annual coke capacity, and the coal-chemical materials produced. The "furnace" plants I are those owned by iron and steel companies and most of them are integrated with blast furnaces to facilitate transportation of raw materials for steel making. Furnace plants represented 87 percent of the total coking capacity and their average annual capacity was 1. 3 million tons. The "merchant" plants are those associated with chemical companies, gas utilities, as well as those that produce metallurgical, industrial, and residential-heating grades of coke and coal-chemical materials for sale on the open market. These coke plants are smaller than furnace plants and in 1959 their average coke capacity was only 520 thousand tons annually. On December 3l, 1959, there were 54 furnace and 20 merchant oven-coke plants in existence in the United States.

Jan 1, 1960

By B. E. Blair

Accelerometers of various types, velocity gages, and a displacement meter are shown to give reliable date when measuring seismic pulses generated in rock by the detonation of explosive charges. Displacement and velocity records are differentiated and shown to compare with velocity and acceleration records. Velocity and acceleration records are integrated and shown to compare with displacement and velocity records, respectively. The experimental data are shown to satisfy the scaling laws developed by dimensional analysis. Propagation laws are developed for peak amplitudes of the first pulse of displacement, velocity, and acceleration. These propagation laws are shown to be independent of the gage used to obtain the data. As a result, the usefulness of the various gage types can be extended. For example, accelerometers, because of their high frequency and amplitude limits, can be employed near the shot to measure acceleration and/or velocity (by integration), Velocity gages, because of their higher sensitivity, can be employed at relatively larger distances from the shot to measure velocity and displacement or acceleration (by integration or differentiation).

Jan 1, 1954

By J. N. Geyer

In coal mining it is essential to maintain the roof over working places and roadways in the safest condition possible. To this end every effort should be made to win the coal with the least injury to the immediate roof strata, as broken strata certainly present a greater hazard than a sound roof. The chief causes of injury to the roof in mining are inherent in the method of development and recovery and the blasting practices followed. The first, which may produce severe stresses in the roof, weakening it over comparatively large areas has been discussed in previous papers3 and will not be treated here. The second cause is local in effect, but with the same general method of blasting being employed throughout a mine having uniform roof conditions, similar effects on the roof may be expected. In the Bureau's investigation of accidents caused by falls of roof and coal it has been necessary to make a careful study of blasting practices and their effect on the immediate roof. The purpose of this paper is to show the effect of the observed methods of blasting on different types of roof and to point out methods that have been proved to be least harmful to the roof strata. The data used in preparing this paper have been compiled from reports made by Bureau engineers engaged in the study of falls of roof and coal, and embrace mines opened in the Pittsburgh coal in Ohio and Pennsylvania and the Pittsburgh and Sewickley coals in West Virginia.

Jan 1, 1933

By N. B. Greninger

The U.S. Bureau of Mines conducted explosion tests on various full-scale cementitious bulkheads used in abandoned mine areas to evaluate the ability of the bulkheads to withstand gas explosion over-pressures of 20 psig. Tests were performed on 120-ft2 solid-concrete-block seals of varying thicknesses and designs. Of the seven solid-concrete-block seals tested, only the standard-type seal, having a 16-in thickness, keyed at the floor and ribs, all joints mortared, wedged at the roof, and a center pilaster, maintained its integrity when subjected to a 20-psig pressure wave. After being subjected to repeated explosions, the standard seal only showed a small hairline crack and had an air leakage of 87 ft3/min at a pressure differential of 1 in H2O. Tests were performed on cementitious foam seals of varying thicknesses and compressive strengths. Explosion tests have shown that 200-psi strength, 4-and 8-ft-thick seals can withstand a 20-psig pressure wave. At a pressure differential of 1 in H20, no air leakage was detected through the 200-psi, 8-ft-thick foam seal.

Jan 1, 2010

By K. E. Stanfield

The Bureau of Mines oil-shale program, initiated in April 1944, is concerned with the future utilization of oil shale in the United States as a source of liquid fuel. One phase of this program is the accumulation of information on the location, extent, and richness of oil-shale deposits that may be suitable for processing to oil and related products. A previous publication5/ presented, by means of logs, the oil yields of sections of the largest known oil-shale deposit in this country -the shales of the Green River formation occurring in adjoining portions of the states of Colorado, Utah, and Wyoming. This present supplementary report presents the oil yields of cores from 48 holes drilled during 1952 to 1954 in the Green River formation of Colorado, which is distributed over an area of more than 1,000 square miles. The oil yields of the 48 oil-shale sections, involving the assay of 9,651 samples, representing 22,451 feet of core, are presented by means of logs. In addition, the oil yields of samples comprising the maximum sections of 15 gallon-per-ton shale are tabulated for these cores and for the 28 cores from Colorado which were logged in the previous report.

Jan 1, 1957

By E. J. Ostrowski

The Bureau of Mines experimental blast furnace, Bruceton, Pa., was operated to determine the feasibility of replacing part of the burden coke requirements with solid fuels injected directly into the smelting zone. This report describes preliminary experiments and information resulting from such an interchange. These experiments were made on an all-sinter burden, using two parts sinter to one part fuel. Anthracite fines were injected intermittently into the smelting zone of the furnace by compressed air through auxiliary tuyeres. The interval between injection periods ranged from 5 to 15 minutes. A maximum of 28 percent of the stock column coke bur-den was replaced. This percentage replacement caused furnace malfunction, but the overall percentage of these experiments indicated that 20 to 24 percent of the coke requirements could be replaced in this way. The metallurgical coke consumption was decreased by as much as 377 pounds (25.3 percent) per ton of pig metal produced. Buckwheat anthracite No.4 was used in this experiment; however, such solids as bituminous coal or coke fines also could be used. The frequency as well as the quantity of solid fuel injected has a direct bearing on furnace performance.

Jan 1, 1960

By Joseph J. Wallace

1. The Pocahontas No. 3 and Pocahontas No. 6 ere the two most important beds in Mercer County from the standpoint of present production and recoverable reserves. 2. Known measured and indicated reserves in all beds, based on a minimum thickness of 14 inches and 1,800 tons per acre-foot of coal in place, are estimated at 188 million short tons as of January l, 1953. Of this total, 80 million short tons are in beds 28 inches and more thick. Areas in each bed were omitted from the estimate because available data relative to the bed characteristics are too meager to make an estimate that conforms with the definitions of measured and indicated coal adopted for this study. Should future drilling or development prove reserves in these areas, such reserves should be added to the total estimated reserves.

Jan 1, 1954

By J. W. Montoya

The chemical, mineralogical, and physical properties of berylliferous tuff from Spor Mountain, Utah, were determined in laboratory research conducted by the Bureau of Mines. Two beryllium-bearing minerals--a hydrated form of bertrandite and a berylliferous saponite--were found to be present in the tuff. Each of the beryllium mineral forms predominated in samples from certain areas of the district. The possibility of practicably separating the beryllium minerals from the tuff by physical methods appears remote because of the complexity and nature of the mineral associations. Beryllium mineralization in the region occurs in a rhyolitic tuff usually overlaid by unaltered rhyolite. The beryllium is distributed erratically throughout the tuff in individual mineralized bodies. Many samples of mineralized tuff were examined and found to contain from 0.2 to 1.8 percent BeO. All samples contained a hydrated form of bertrandite that was nearly cryptocrystalline, whereas samples from one area also contained a berylliferous saponite of variable beryllium content. Neither mineral has been previously reported, although related bertrandite-like minerals have been found in the U.S.S.R. and berylliferous minerals of the montmorillonite group have been synthesized. Most of the minerals of the tuff are finely crystalline, intergrown, and have similar densities. The complexity and intimacy of the mineral associations, the difficulty of achieving effective mineral liberation, and the highly friable nature are characteristics that lead to the conviction that tuff of the type examined cannot be effectively concentrated by gravity, magnetic, electrostatic, attrition, washing, sizing, or flotation methods.

Jan 1, 1962

By John S. Berber

Bureau of Mines research on the production of phthalic anhydride by vapor phase oxidation of hydrocarbons began at Morgantown, W. Va., in October 1960. For these preliminary studies, a known solution containing naphthalene and 1-methylnaphthalene was used. In another series of tests, pure naphthalene was dissolved in different solvents. A series of investigations was conducted with the neutral oil fraction (210° to 270° C) of a low-temperature tar obtained from the Texas Power and Light Company's low-temperature carbonization unit. The neutral oil fraction was first dealkylated, and then the product was distilled into 210° to 250° C and 250° to 270° C fractions. These two fractions were then subjected to catalytic vapor phase oxidation. Although a number of catalysts was tested, most of the experiments were made using 9 percent vanadium pentoxide on inert alumina. The activity of this catalyst did not decrease after 162 hours at 1,030° F. Studies were conducted over a temperature range from 610° to 1,180° F. An operating temperature of 1,030° F yielded 47.50 percent of phthalic anhydride from a synthetic mixture of naphthalene and 1-methylnaphthalene. The highest yield (84.40 percent phthalic anhydride) was obtained with naphthalene dissolved in toluene and also with naphthalene dissolved in methyl ethyl ketone. With a 210° to 250° C fraction the highest yield of phthalic anhydride obtained was 12.32 percent at a temperature of 900° F, and, with a 250° to 270° C fraction, it was 15.80 percent at a temperature of 875° F.

Jan 1, 1963