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By R. V. Ramani, K. Bode

"IntroductionIn part 1 of this paper, the results of the experimental work in a wind tunnel under controlled conditions were presented. To understand the spatial and temporal behaviour of airborne dust under mine conditions, experiments were performed in the Lake Lynn Experimental Mine of the US Bureau of Mines. The mine airway experiments were designed to study the variations of ambient concentrations along the length of an airway, and the deposition rates and deposition velocities in an airway cross section. Other objectives of the study were to investigate the spatial and temporal behaviour of airborne dust with regard to air velocity and particle density, the size distribution of the airborne dust cloud as it travels in the airway, and the size distribution of the dust particles deposited along the airway. Other objectives of the study were to compare wind tunnel results with In-Mine results, whenever appropriate.Experimental ProcedureLocationA complete description of the Lake Lynn Laboratory Mine of the US Bureau of Mines (USBM) can be found in the US Bureau of Mines Information Circular 8911 (Mattes et al., 1983). This mine provides excellent facilities for dust experiments. The airway dimensions are similar to those of a coal mine airway. The mine has a drift (D Drift, see Figure 1) that extends almost 500 meters without interconnecting airways, stoppings, or bends. The D drift measures approximately 6m in width and 2,25m in height.The Lake Lynn Laboratory Mine· is ventilated by a variable speed, 75 kW (100 hp), 28,3 m3/s (60 000 cfm) blower fan installed on the top of the ventilation shaft. Air velocities of 1, 1,3 and 1,8 m/s, respectively, could be obtained in the D Drift. However, the temperature changes at the surface can cause small variations in air quantities and velocities underground. The low background dust level in the mine reduces the problem of sample contamination from other sources of dust. Extended periods of uninterrupted data collection can also be scheduled. Mine airway parameters which are important for the analysis include the dimensions of the airway and the ventilation data such as the air velocity, and the pressure loss in the drift. All these data were acquired during altimeter surveys which were conducted prior to each of the experiments."

Dec 1, 1995

By H. S. Chiang, S. S. Peng

"SummaryAir velocity distribution surveys on four mechanized longwall faces are presented and discussed. Although there are discernible patterns in the velocity distribution, there are wide differences in distribution in the cases examined. This can be attributed to differences in mining and face geometry. An important factor is air leakage into the goaf.IntroductionOver the past two years several intensive air velocity surveys have been carried out in West Virginia coal mines as part of a programme aimed at understanding dust transport mechanisms on mechanized longwall faces. These are presented here and some of the patterns of velocity distribution are discussed.Ventilation survey methodologyThe air velocity vector consists of three mutually perpendicular subvectors; the vertical subvector is directed toward the roof, one of the horizontal subvectors is parallel to the coal face and the other pointing to the gob. The air velocity which was measured in this paper is the horizontal subvector parallel to the coal face.To record the velocity distributions at the longwall faces, the surveying stations and the surveying points were established as follows.(1) The first station usually was set at the third support from the headentry T-junction. Within the first few supports, the air coming from the intake airway encounters a sharp right-angle turn at the T-junction, and forms a turbulent region close to the coal face. The intake air in this region is also influenced by the motors that drive the face conveyer. But when the air reaches the different. But for the same face the number was the same. The main air-travelling area was divided into six to 12 equally sized sub-areas, and each surveying point was situated approximately at the centre of the sub-area.The velocity survey was carried out by two persons, one reading the meter and the other taking down the readings. A complete survey at one station took about 5-8 min. Usually, the dimension of the cross-sectional area was measured during that time.A third person used the smoke tubes to observe the following: (I) the air movement patterns at different locations; (2) the extent and the direction of the air leakage; and (3) the machines' influence on the main airflow. At the same time he observed and recorded the roof conditions and the compactness of the roof-caving in the gob."

Mar 1, 1989

By A. B. Hooker, R. A. Kearns

Material used for flame safety lamp gauzes should possess the following qualities: (1) mechanical strength, (2) resistance to corrosion under service con- ditions, (3) a relatively high thermal conductivity, (4) a melting point above the ignition temperature of mine gas. Heretofore, three metals have been used for flame lamp gauzes, namely: steel, brass and copper. While each of these has the requisite qualities to some extent, none of them is an ideal gauze material. Copper gauzes lack mechanical strength, especiall, after they have been heated; brass and copper both have comparatively low melting points so that before the gas ignition is reached the gauze metal is attacked and oxidized under the action of the heat. Steel corrodes readily in damp atmos- pheres and if the lamp stands unused under these conditions for some time, the gauzes become unfit for service. Tests of these three materials have shown that steel offers the highest degree of safety. Manel metal, the natural alloy of nickel and copper, is practically noncor- rodible. Its relting point is approximately the same as that of steel. From time to time it has been suggested that this alloy might be a better gauze material than steel, since gauzes made of Monel metal should have all of the good qualities that steel gauzes possess and, in addition, would not corrode appreciably in ordinary mine atmospheres. One advocate of this use for Monel metal, the International Nickel Co., through their development engineer, Mr. Robert J. McKay, cooperated with the Interior Department, Bureau of ines, to the extent of supplying twelve sets of 30 mesh, Monel-metal gauzes for tests. Those gauzes were made up by the same lamp manufactur- er ho rade the steel gauzes used in the previous investigation. Being made by the same processes as the steel gauzes, these Manel-metal gauzes were needlessly given the same anti-rust coating that is always applied to steel gauzes to prevent any corrosion before the gauzes are used. In previous tests the Bureau of Mines had collected data as to the behavior of steel gauzes. In determining the value of Monel metal as a gauze material, it was planned to obtain similar data with Monel-metal gauzes, and thus make a direct comparison with the steel gauzes. In addition, steel and Monel-metal gauzes

Apr 1, 1923

By V. A. Marple, K. L. Rubow

"IntroductionA particle dispersion system has been developed to aerosolize bulk powders or material deposited on surfaces such as filters. The size distribution of the aerosolized particles can then be obtained by sampling the particles with an aerosol sizing instrument such as a TSI aerodynamic particle sizer (APS) or a cascade impactor.Of particular interest is the applicability of using this technique to obtain the size distribution of particles collected on filters. This system has been used to determine the size distribution of coal dust particle collected on membrane filters. The use of this device in this application is quite intriguing in that, if this technique can be shown to accurately determine the size distribution of airborne particles, one can easily collect particle samples using either personal or area particle samplers, determine the mass concentration of the collected material through gravimetric techniques and then use the particle dispersion system coupled with the APS to determine the particle size distribution. With this technique one can determine the particle size of dust in environments, such as underground mines, where it would be inconvenient or impossible to use a particle sizing instrument.Particle Dispersion SystemA prototype particle dispersion system has been developed for aerosolizing particles. A schematic diagram of the system is shown in Figure 1. The system consists of mechanism for advancing the particle sample beneath a dust pickup tube, a venturi for deagglomerating the particles, and a instrument for analyzing the aerosolized particulate matter.The procedure to analyze a particle sample is as follows. The particles to be aerosolized are placed on a sample plate from where the particles are picked up by an air stream flowing upward in a small capillary tube (dust pickup tube). This tube operates as a small vacuum cleaner passing over the sample plate. The air is drawn up the capillary tube at a rate of 2 L/m by a venturi aspirator. This is accomplished by placing the outlet of the capillary tube near the throat of the venturi. The accelerated velocity of the air creates a low pressure in this region which draws flow up the capillary tube. From the outlet of the capillary tube particles are transported to the inlet of the APS where a small fraction are sampled and the aerodynamic diameter of the particles determined. The particles which are not measured by the APS are passed into a HEPA filter."

Mar 1, 1989

By Peter Kovalchik, Ed Kwait, T. Michael Durr

1. BACKGROUND Many research projects on engineering controls for noise reduction in mining were conducted by the former U.S. Bureau of Mines, mostly under contracts from 1972 through 1986.1 Contracts were awarded to determine which mining equipment was responsible for worker noise exposure over the permitted Federal Regulation level of 90 dBA for 8 hours. Dominate noise sources, for each piece of equipment was also determined. The final goal was to develop engineering noise controls for those pieces of equipment and then evaluate them. This was the general approach taken in each research project or contract. This paper deals with what is known about the control of noise from continuous miners, one of the most commonly used underground mining machines. A literature review on continuous miners revealed that the relevant noise sources were divided into four categories: conveyor, cutting head, drive train, and motor noise. The dominate noise sources reported were the conveyor noise and the cutting noise.2 Conveyor noise consists of impacts and scraping at various locations on the conveyor structure. These impacts are associated with discontinuities on the top and bottom decks, impacts between the chain and the tail roller, impacts on the sidewalls at path points of discontinuity under horizontal pivot, and impacts on the flex plate.3 Scraping occurs along the entire length of the chain conveyor. The other dominate source, the cutting of coal, has also been addressed in past research.4, 5, 6 The research presented in this paper only addresses the controls associated with the conveyor. The literature suggests several effective treatments for noise control: retrofitting the chain conveyor with a urethane coating or a urethane sleeve on the chain flights, using a urethane coating or urethane sleeve on the tail roller, constrained layer damping of the decks, and the use of a modified take up plate.7, 8 However, the in-mine durability of these concepts has yet to be proven for the conveyor system. This paper analyzes extensive noise and urethane durability testing on a Joy 14CM9 continuous miner’s conveyor assembly, tail roller, and bottom plate conducted at the Pittsburgh Research Lab facilities. It will also discuss the initial underground testing results for the urethane-coated conveyor flights and tail roller.

Objective Develop a method to warn of hazardous conditions and enhance safety during operation and inspection of vertical mine shafts and hoists. Background Mine hoisting is the most critical and potentially the most dan¬gerous operation in underground mining. Many hazards are related to falls of miners and materials, ground instability, and malfunction or lack of safety devices. All have the potential to cause a cata¬strophic accident. A condition known as "slack rope" is particularly dangerous, expecially if it occurs without the operator's awareness. The most common occurrence of slack rope is in metal or nonmetal mines that have a history of ground control problems near the "haft. Approach Studies of hoist safety features and Mine Safely a and Heath Administration (MSHA) accident statistics suggests a need to measure wire rope tension, conveyance speed and position, and shaft guide misalignment. The most accurate data are obtained directly from the conveyance. Such data could warn of potentially dangerous situations during hoisting operations, as well as during inspections, maintenance, and lowering of large equipment. Figure I shows the concept for monitoring wire rope tension, conveyance position, and guide displacement. Progress to Date A conveyance-mounted system has been developed that monitors rope tension: hoist conveyance position, speed, and load; and shaft guide misalignment. A "flex-beam" load sensor mounted just above the- conveyance protects against slack and tight rope. The measurements are transmitted 60 times a second via wireless equipment from the conveyance to the hoist room. An encoder mounted on the conveyance eliminates the need to correct the position because of rope stretch. Figure 2 shows the encoder wheel and guide displacement sensors being mounted on a cage in¬spection deck. A potentiometer monitors misalignment of the shaft guides during inspections. Information about the cage or skip can be acquired by the shaft inspector using a hand-held indicator. Inspection can also he conducted remotely from a station or hoist room using the wireless link to a laptop computer. Figure 3 shows test results from a simulated slack rope condition. Patent Status NIOSH has a patent pending on this device entitled "Cable-Load¬Sensing Device." For More Information NIOSH is seeking cooperators to assist in further tests of this technology to improve hoisting safety. Additional information can he obtained by contacting Michael J. Beus, Steven Iverson, or Elaine Cullen, Technology Transfer Officer, NIOSH, Spokane Research Laboratory, Spokane, WA 99207 at (509) 354-8000.

Feb 1, 1998

This chapter presents data on occupational exposures from several sources. Data from compliance samples obtained by MSHA inspectors are presented in table 6-1 for coal mine dust and silica dust, and in tables 6-2 through 6-6 for metal fumes. Since these samples were taken for compliance monitoring rather than as part of a survey of the industry, it is difficult to predict how well they indicate actual exposures for all mining operations. In the coal industry, 7.8% of all respirable dust samples during 1986-1995 were above the permissible exposure limit (PEL). Of silica dust samples obtained during the 10-year period, 23.7% were above the PEL in coal, 16% in metal, 10.8% in nonmetal, 9.1% in stone, and 7.6% in sand and gravel. Of metal fume samples, silver samples showed the largest percentage above the PEL-approximately 48% of samples in both metal and nonmetal. Tables 6-7 and 6-8 present data on noise exposures from the MSHA "dual threshold" survey, which was published in the Federal Register [61 Fed. Reg. 66347 (1996)] as part of a proposed rule change for occupational noise exposure in mining. This study examined a group of samples obtained during 1991-1995 and compared the percent of samples that were above two separate specified limits. The first limit was the current noise standard, a time-weighted average of 90 dBA, calculated to include only exposures at 90 dBA and above. The second limit was a time-weighted average of 85 dBA, calculated to include exposures at 80 dBA and above. Tables 6-8 and 6-9 refer to musculoskeletal overload conditions examined in the National Occupational Health Survey of Mining (NOHSM) [NIOSH 1996]. Table 6-8 shows the operational definitions for each condition; table 6-9 shows the percentage of the workforce potentially exposed, by commodity. Across all commodities, a large proportion of workers were exposed to musculoskeletal overloads due to positioning of the neck and back; positioning and motion of the forearms, arms, and shoulders; heavy lifting; and positioning and movement of the lower limbs. The recorded overloads were defined in the survey and did not exceed any NIOSH, MSHA, or OSHA guidelines for musculoskeletal overloads. Further information on this survey appears in appendix A.

Jan 5, 2000

By T. C. Phalen

"Necessary Qualities There mines are located in out of the way places, where timber is scarce and its price high, the problem of protecting it from fire risk is of considerable practical importance. Scarcity of timber for mining purposes is commonly met with in the more arid parts of the country, where, also, mine timbers are most likely to become dry, and inflammable, and where, therefore, the risk from fire is accentuated. Any effective and cheap method of securing adequate protection of such timber at the lowest possible cost is worthy of careful investigation. The use of magnesite cement has been suggested in this connection.The necessary qualities for any substance for the use suggested above may be summarized as follows: It should be resistant to abrasion, and to impact and structural stresses. It should be durable when subjected to the action of the ele¬ments, and stable to any minor derangements of the base on which it is placed. It must be relatively unaffected by changes in temperature, by the action of water, and should adhere to the material on which it is placed, and it should also be free from shrinkage cracks due to setting up the material, Magnesite cement appears to fulfill the various conditions outlined above.The elasticity of such cement is the outstanding quality which makes it desirable for exterior work. It also possesses the ability to adhere to all materials in a marked degree. Its freedom from expansion and contraction after setting is one of the principal reasons for its adhesive quality. It may be applied to old brick or stone walls, even tho they have been repeatedly covered with paint, and the quality of ad-hearing to wood makes it possible to secure a weather-proof protection to the wooden casings and trim of exterior openings. The quality of adhesion and elasticity is easily proved by applying a 1/2"" layer to a 1"" board which has been first covered with metal lath. Such a board, about 8 feet long, may be repeatedly deflected about 3"" in the center without breaking the bond to the board, or causing any cracks to appear in the magnesia cement. The ability to resist fire as well as the possession of insulating qualities are also desirable characteristics of magnesia cement stucco."

Dec 1, 1919

By L. C. IlsLey

One of the details which the engineers observe closely in conducting tests on explosion-proof type equipment at the Pittsburgh experiment Station of the Bureau of Mines is the construction of covers and their means of fastening. These covers are of particular concern from a safety standpoint, because they are very liable to be frequently disturbed for the purpose of making inspection or repairs to internal parts of motors, controllers, etc. When depended upon to permit frequent inspections it is desirable from a practical operating standpoint that the cover be easily and quickly removed and replaced. The position that been taken by the Bureau of Mines is that the questio of ease with which covers will permit inspection should be determined by the mamu- facturer and the user. If the scheme used requires considerable effort, as in the removal of several bolts, it has been assumed that the one going to all this trouble to get the covers off could be trusted to put the cover back in place properly, especially as the approval plate generally calls specific attention to this point. Where a cover is simply screwed in place, fastened by a single bolt, or secured by a hing in or other similar arrangement, the bureau has not been willing to assume that the, cover would in all cases be properly fastened and has required further security by means of a lock or seal. In the course of its investigations the bureau has had occasion to inspect and test five types of covers as follows: (1) Covers having two plane surfaces bolted together. (2) Covers having two plane surfaces held together by lugs, clamps, or other means besides bolts. (3) Covers which have screw thread and fit a similar thread in the compartment. (4). Covers having plane surfaces with the addition of a tongue-and-groove joint at about the center of the plane surface and in general secured by hinge pins or similar means of fastening. (5) Covers in the form of an annular ring accurately fitted to a motor shell which can be rotated around the shell, and thus make easy access to inspection openings in the shell and rotated back again to effectively seal these openings. Tests. have also been made of large bolted covers having a smaller and more accessible cover within the large one, which could be opened for inspection purposes.

May 1, 1930

By C. P. Bowie

"It is well known by all who are familiar with the oil industry in America that the oil pipe line is as fundamentally a part of this industry as the railroad is of most other American industries. The oil pipe line was first introduced about 56 years ago, and since that time has so demonstrated its superiority as a means of carrying crude oil froze the well to the refinery, that this method of transportation has largely superseded all others. This has made possible the building of refineries in or near the large consuming centers, rather than at the wells, which are usually remote from the centers of population.The pipes for conveying the oil are laid on the surface of the ground, or at a depth varying from 18 inches to 3 feet beneath the surface, and the main lines are generally eight inches in diameter. The oil is forced through the pipes by means of pumps operated either by steam or by internal combustion engines. The pump stations are located from 1 1/2 to 90 miles apart, varying with the condition of the country through which the pipe lines extend, and the viscosity of the oil to be handled.Some of the larger pipe-line systems and hundreds of miles in length. It is estimated by the U. S. Geological Survey that the total mileage of oil trunk lines in the United States today is approximately 34,000, and that the gathering systems, which are a fundamental part of the trunk systems, aggregate about 11,500 miles in length, making a total of 45,500 miles.Using those figures as a basis, some general conclusions as to the investment represented will be made. At the time most of the lines were constructed, the average cost per mile based on eight-inch pipe was about $6,500. The cost of the average pump station at that time varied from $130,000 to $250,000. The cost at the present time, on account of the increase in price of labor and materials, would be much in excess of these figures. The fixed investment in pipe lines is, therefore, very extensive. It corresponds closely with the actual cost of the property, and is estimated to be approximately $500,000,000."

Sep 1, 1920

By Bob Thomas, Syd Peng, Quanzhong Gu, Gene Wilson, Gerald L. Finfinger

Identifying the properties of overlying rocks in underground mining operations is important to ensure the appropriate roof support design is used to maintain stability of the mine entries. Recently J. H. Fletcher & Co. developed a monitoring and control system for roof bolters for the underground mining industry. The system records the drilling parameters used during roof bolt drilling and the information can provide insight into the physical properties of the roof strata. The parameters include thrust, rotational speed, torque and velocity and the measurements are collected every 0.1 second during the operation. The drilling parameters were analyzed to determine the application of identifying the strength of rocks being drilled from the measurements. The data was converted into the specific energy of drilling which is a measure of the amount of energy required for removing a given unit of rock during a drilling operation. The laboratory studies completed to date indicate a fairly high correlation between the specific energy of drilling and the unconfined compressive strength of :he rocks that were drilled. Additionally, the drilling parameters were shown to be effective for identifying the presence of fractures or bed separations between rock layers. The thrust, torque and specific energy of drilling were all good indicators for identifying the fractures or separations. Regardless of the drilling parameters used during the drilling experiments, the location of the fractures were identified. In order to determine the application of the drilling parameters for identifying roof rock properties, two series of experiments were conducted. The first series of experiments used three "manufactured" roof layers that had various rock samples embedded in concrete blocks. The rock samples included three types of sandstone, marble, and argillite. Another concrete block was poured with foam inserts to simulate large bedding separations (2 to 8-in). Two other manufactured blocks were constructed using high-strength concrete with cardboard layers embedded to simulate smaller fractures or bedding separations. The size of the cardboard layers varied from 1/8- to 1-in thick. One of the blocks had the cardboard layers embedded at an inclined angle to determine the effect of the orientation on the drilling parameters. A series of experiments was conducted with the rotational speed and the penetration rate held constant and the thrust and torque allowed to vary. The information collected from the experiments is used to determine the application of the drilling parameter measurements for identifying rock properties, fractures and bedding separations.

By Stephen C. Tadolini, Thomas M. Barczak

An underground investigation was designed and implemented by the National Institute for Occupational Safety and Health (NIOSH) to evaluate the effectiveness of pre-driven longwall recovery rooms supported with pumpable concrete roof supports. The pumpable roof supports were instrumented with hydraulic pressure cells to measure support loading. Displacement of the support was also measured to assess the support reaction to ground movements. Movement up to 12.5 m (40 ft) into the roof was measured with sonic probes and multipoint extensometers. Borehole pressure cells were installed in the panel fender and outby pillars to evaluate load transfer. Shield loading was also assessed to determine the impact of the panel fender yielding on face loading. The results show that the yielding of the panel fender produced uncontrollable convergence that caused yielding of the shields and stiff pumpable roof supports when the face was approximately 3.12 m (10 ft) from the recovery room. The shields continued to yield until the face advanced into the recovery room. Convergence in the pre-driven recovery room typically reached 15-20 cm (6-8 inches) once the pumpable supports yielded and shed load through a declining residual load capacity prior to being cut out by the longwall shearer. Roof deformations occurred beyond the 10 m (32 ft) horizon with total deformation ranging from five to 10 cm (2 to 4 inches), indicating that standing support was necessary to help control the span as the longwall face advanced into the pre-driven recovery room. Stability improved as the pumpable supports were cut out, thereby reducing the span between the shields and the outby pillars. In the end, the shields were successfully recovered under stable ground conditions. However, a disconcerting discovery was the load shedding of the outby pillars as the recovery room was mined into. This behavior was not expected with a relatively narrow room that is fully supported with standing support. If this mechanism is indeed occurring, the width of the room and performance of the standing support becomes even more critical. Premature failure of the support can lead to excessive convergence on the shield line that cannot be controlled even with modern shields, resulting in instabilities that can lead to catastrophic results. Understanding the load transfer mechanics associated with the advancement of the longwall face into the pre-driven recovery room and the widening of the room to recover the longwall shields is providing valuable information for the refinement of the NIOSH Ground Reaction Curve for standing and intrinsic roof support system design.

By Kevin A. Rider, Bernard J. Martin

Reaching movements are planned and subsequently executed [1] using visual and somatosensory feedbacks [2], where absence of visual feedback is known to increase endpoint variability [3]. Visual occlusion decreases the ability to make rapid online compensatory movements, which results in initial radial deviations that are highly correlated with radial dispersion at the target. Perturbations of rapid, visually-guided reaches are compensated on-line and result in endpoint dispersions poorly correlated with initial deviations, emphasizing the strong effect of visual feedback in temporally-constrained reaching tasks. In control conditions (no vibration), these uncompensated, rapid reaches serve as estimates of the individual’s intended trajectory. When ride motion is present, trajectories of rapid, visually-occluded reaches provide a measure of the natural biodynamic response of the cantilevered spine-arm-hand linkage. These intended movement trajectories and the biodynamic response (vibration feedthrough) are used to predict the effect of ride motion on the performance of rapid reaching tasks. Goals of this study are to investigate the influences of vehicle motion on human reaching and pointing, and to reveal movement strategies used in visually-occluded reaching tasks. Methods A six degree of freedom human-rated Ride Motion Simulator (RMS) was used to generate a dynamic vehicle environment. Participants performed discrete, rapid pointing tasks to targets presented on three touchpanel displays under stationary and random whole-body vibration. Reach instructions included successfully reaching identical circular targets (Ø = 0.25”) with the right index fingertip as fast as possible. Targets were presented on resistive-touch displays mounted approximately 60 cm from the participant’s nasion. The touchpanel displays were located in the forward and lateral directions at eye level, and forward at 45° of elevation. These displays measured the spatial error of the reach destination. A ten-camera VICON motion capture system recorded the upper body kinematics of the participant. Reflective markers were placed on the participant’s torso, head, and arms. Initial kinematics of the fingertip (i.e. time and magnitude peak tangential velocity) and tangential velocity at target were used to estimate the planned endpoint of the reach. Results and Discussion Ride motion resulted in increased endpoint variability compared to reaches performed in the stationary condition. Reaches to the elevated touchpanel consistently resulted in the largest variability across all motion conditions, suggesting that a vehicle occupant would not be capable of accurately activating a control in that location. Principal axes of endpoint ellipses were along and perpendicular to the direction of fingertip movements. Example graphs of endpoint variability with ellipses containing 95% of the data points are shown in Figure 1. These ellipses

Jan 6, 2006

By 1928-05-01

"The carnotite ores of Colorado and Utah were the chief source of radium for many years. In the fall of 1922; news of the very rich deposits of radium ore discovered in the Belgian Congo caused almost complete cessation of domestic production. The Congo ore contains pitchblend and a number of alteration products, including the new minerals bequerelite, curite, kasolite, stasite, and dewindite. Reports indicate that a considerable amount of ore containing over 50 per cent U308 has been produced. It seemed certain that radium could be extracted from the Congo ore at a much lower cost than from the relatively low-grade carnotite; and in order to preserve their elaborate marketing organizations, the larger American producers made agreements with the Belgium syndicate to market the foreign product in this country. The price of radium has been held at $70 per milligram. Extraction of radium from carnotite by the usual methods does not appear to be profitable at that price. With one exception, all American firms ceased production after their accumulated stocks of ore had been treated. The United States Radium Co., which continued to treat carnotite for several years, has recently shut down its plant.Although there is probably a large profit derivable from the treatment of Congo ore, it is not likely that the Belgium Syndicate will meet sufficient competition to cause any reduction in the present quotations. It is not known how long the high-grade ore will last, but when exhausted the price of radium may be expected to advance. According to K. L. Kithil3 it is the opinion of interested parties that, so far as can be forseen, radium will in future years never again be obtained at the present low price.Carnotite contains the valuable element, vanadium, which is not present in the Congo ore. The recovery of-vanadium is an important factor in the cost of extracting radium from carnotite, and under favorable conditions carnotite can be profitably treated for the vanadium alone. Vanadifecous sandstones are found ranging all the way from roscoelite (vanadium mica), containing little or no uranium; to a high-uranium and low-vanadium carnotite. Ores containing 2 per cent or more U308 are classed as radium ore and are sold on the basis of the uranium content, which is proportional to the radium content. Only in exceptional cases have ores been evaluated on the basis of both the radium and vanadium contents, although the presence of vanadium has undoubtedly influenced the price of radium ore."

May 1, 1928

This chapter explores leadership behavior in a life-threatening situation-fire in a coal mine. Previous chapters have discussed the database of interviews with miners who escaped from underground fires. Researchers raised questions such as: (1) Who led the miners out of the mine? (2) Did leadership make a difference in the escapes? (3) Was the escape leader the hierarchal leader? (4) What, if any, characteristics did the escape leaders possess? Subsequently, researchers analyzed the mine fire database from a group interaction perspective to address the leadership questions posed. Leadership has been one of the most researched topics of human behavior in the twentieth century. Studies have ranged from individual characteristics of leaders, to situational leadership, to interaction of leader and follower, suggesting different leadership techniques for different followers. The question that emerges here is: Are there different types of leadership that "fit" different kinds of situations? In a crisis situation like that examined in the present study, such information about leadership may significantly improve the chances of escape. To address these issues, the study team looked at the formal authority structure before each fire, considered leadership behavior or lack of leadership during the escapes, and examined those conditions associated with the emergence of leadership. According to Bardo [ 1978], "Emergent behaviors are those forms of action, and the norms, values and beliefs governing those actions, that rise out of the disaster situation." This chapter discusses previous studies in the area of crisis leadership and examines the emergent behaviors of leaders under duress during the mine fire escapes. Previous Studies The research on leadership during emergency situations has consisted mainly of simulation and field studies, with the principal concern being escape from building fires. During the 1980s, Hayashi [1988] created a computer simulation model to evaluate leader behavior in a fire. Although his purpose was to aid in planning for disaster prevention, his findings are relevant because they address the issue of situational leadership in crisis-where a leader changes his or her behavior to fit the situation. Essentially, his simulation model was designed to judge the actions and thinking of leaders. The simulation was tried by 101 subject/leaders 4 times each. The simulation consisted of amaze containing the leader, an informal leader, and 50 evacuees. Interestingly, the results indicated that the leader's actions were not dictated by circumstances. Any differences in behavior were attributed to the individual characteristics of each leader. The study also showed that the worse the situation became, the less individual

Jan 1, 2000

By W. W. Adams

Fatal accidents at coal mines throughout the United States numbered 170 during May , 1923 , according to reports received by the Bureau of Mines of the Department of the Interior from State mine inspectors . The production of coal during the month was 54,649,000 tons ; hence , the fatality rate was 3.11 per million tons . In May last year the fatality rate was 4.22 , based upon an output of 20,636,000 tons , including 35,000 tons of anthracite . Because of the miners ' suspension of work last year, the production of fresh-mined anthracite in May 1922 was completely stopped ; in May 1923 the anthracite output was 9,573,000 tons . For bituminous mines alone , the fatality rate in May 1923 was 2.67 per million tons as compared with 4.17 in May a year ago . During the first five months of the present year , the fatality rate for bituminous mines was 3.55 per million tons , based on 809 deaths , as against a rate of 3.70 for the corresponding months last year, based on 611 deaths . The rate for 1923 represents a reduction of 4 per cent . For anthracite mines alone the five-month fatality rate for 1923 was 5.36 per million tons , based on 228 fatalities , as against a rate of 6.72 for the first five months of 1922 based on 147 fatalities . The rate for 1923 represents a reduction of 20 per cent . The average fatality rate for May during the past ten years ( 1913-1922 ) is 3.37 per million tons for bituminous coal mines and 6.02 per million tons for anthracite mines . A gas explosion , due to open light , occurred on May 6 at Aguila , Colorado , and resulted in the death of ten men . This explosion brought the number of "major disasters " in 1923 to 4 , with a loss of 145 lives . During the first five months of last year, there were 7 similar disasters with a loss of 82 lives . The year's record to date for all fatalities from all causes shows a reduction in the fatality rates per million tons from haulage accidents and from falls of roof and coal , an increase in explosions of gas and dust and in accidents due to electricity , and no change in the fatality rate from explosives . The reductions in haulage accidents and in falls of roof and coal are particularly gratifying, because these two classes of accidents usually cause about 65 per cent of all fatalities in the coal-mining industry . Reports of Investigations , Department of the Interior , Bureau of Mines .

Jun 1, 1923

By William H. Pomroy, Shail J. Butani

"This Bureau of Mines publication presents a statistical analysis of official U.S. Mine Safety and Health Administration (MSHA) metal and nonmetal mine fire reports from 1950 through 1984, plus accounts of selected nonreportable fires (less than 30 min and no injury) and opinion data on fire hazards from mine safety directors. Fires were analyzed by time trends, ore type, ignition source, burning substance, location in mine, equipment involved, means of detection, duration, number of injuries and fatalities, mining method, and successful extinguishing agent. The leading ignition sources were electricity in underground fires and engine heat in reported surface fires. The most frequent burning substance was combustible liquids for all nonreportable fires, reported surface fires, and reported underground fires from 1978 to 1984. For underground fires reported prior to 1978, timber was the leading burning substance. Mobile equipment was the type most frequently involved in both underground and surface fires. Underground fires occurred most often in haulageway or drift areas, and reported surface fires occurred most often in plant and mill buildings, while most nonreportable surface fires occurred in other areas. The most common methods of extinguishment were water hose lines for reported fires and dry chemical handportable extinguishers for nonreportable fires. INTRODUCTIONIn support of the Bureau of Mines research on mine fire protection, two separate baseline studies on mine fire incidents have been prepared. The first study addressed coal mine fires. The results of that study are summarized in the Bureau's Information Circular 8830, ""A Statistical Analysis of Coal Mine Fire Incidents in the United States From 1950 to 1977."" This second report, addressing metal and nonmetal mine fires, is the companion to IC 8830. Together, these two reports provide a comprehensive, factual summary of the mining industry's fire experience.This report, like IC 8830, covers (1) MSHA fire incident reports, (2) records of nonreportable fires (less than 30 min and no injury), and (3) mine safety directors' opinions on fire hazards. The MSHA fire reports provide the most reliable, objective, and accurate historcial record of the major noncoal mine fires that occurred during the study period (1950 through 1984). However, MSHA fire reports alone understate the true magnitude of the fire problem because metal and nonmetal mines have been legally required to report fires to MSHA only since 1968. Although some fires were reported prior to 1968, doubtless a great many were not. Also, the reporting regulations that took effect in 1968 specify that only fires lasting 30 min or longer or involving an injury need to be reported. MSHA fire reports are thus limited in scope by MSHA's legal authority. In order to provide a more comprehensive data base, it was also necessary to gather and analyze mine company records of nonreportable fires."

Jan 1, 1986

By V. A. Marple, K. L. Rubow, B. A. Olson

"A virtual impactor personal aerosol sampler (VIPAS) bas been developed for use in diesel-equipped mines to collect size-separated samples or diesel exhaust and mine dust particles in the respirable size range. This sampler consists or three stages: a 10-mm cyclone, a 0.8- µm cutpoint virtual impactor and two filter cassettes. The large non-respirable particles are removed in a 10-mm cyclone. The respirable particles from the cyclone pass through a 0.8- µm cutpoint virtual impactor with the particles larger than and smaller than 0.8 JI.ID (mine dust and diesel exhaust particles, respectively) collected on two 37-mm filter cassettes. The sampling How rate is 2 L/min so as to be compatible with the standard respirable dust sampler and personal sampling pumps. INTRODUCTIONThe mining industry has been increasing its use of diesel-powered equipment in underground mines, especially in coal mines. With this increased usage has come an increased concern regarding the health effects associated with inhaling diesel exhaust particles since they are in the respirable particulate size range. A Mine Safety and Health Administration (MSHA) advisory committee directed MSHA to develop a sampling protocol for diesel exhaust particles to assess worker exposure (MSHA, 1988). In addition, the National Institute for Occupational Safety and Health (NIOSH) has recommended that ""whole diesel exhaust be regarded as a potential occupational carcinogen as defined in the Cancer Policy of Occupational Safety and Health Administration"" (NIOSH, 1988).Past work, with cascade impactors in both the laboratory and in underground coal mine studies, has shown that the diesel exhaust particles are primarily in the less than 0.8 µm aerodynamic diameter size range whereas the coal dust particles are primarily greater than 0.8 µm (Marple et al., 1986; Cantrell, 1987; Rubow and Marple, 1988; Rubow et al., 1990a, b; Cantrell and Rubow, 1991). This work has been jointly performed by the University of Minnesota and Bureau of Mines researchers over the last 8 years. Typical particle size distributions for a diesel-equipped coal mine and an all electric coal mine are shown in Figure l. Note that the particles in the < 0.8-µ.m mode are the diesel exhaust particles (seen only in diesel-powered mines) and the particles in the > 0.8-µ.m mode are coal particles (seen in both the electric-powered and diesel-powered coal mines). Chemical mass balance analysis was also performed on these two size fractions to determine the degree of particle cross-contamination in the two modes. This analysis showed less than 10% cross-contamination, i.e., in coal mines Jess than 10% of the diesel exhaust particulate was collected in the coarse fraction and Jess than 10% of the mine dust was collected in the fine fraction (Cantrell, 1987; Rubow et al., 1990a)."

Dec 1, 1996

By Joseph A. DeCarlo

This information circular presents data for 1960 on individual coke plants in the United States. All oven-coke plants in existence, whether active or idle at the close of the year, are listed in this report. With two exceptions, only beehive plants on which reports were submitted to the Bureau of Mines are included. Although the Bureau was unable to obtain reports from these two active plants in 1960, the number of ovens and annual coke capacity are estimates, based on previous reports filed with the Bureau of Mines. Accordingly, there were 73 oven-coke plants in existence at the end of 1960 of which 70 were active. Reports were received on 44 beehive plants of which only 13 were in operation at year end. Oven-coke plants are listed alphabetically by States in table 1 which shows the name and address of the company, location of plant, classification or type of ownership, number and type of ovens, annual coke capacity, and coal-chemical materials produced. The terms "furnace" and "merchant" plants used in this and all other reports of the Bureau of Mines designate the type of ownership. Furnace plants are those owned by iron and steel companies whose main business is production of coke for use in their own blast furnaces. Merchant plants are those associated with chemical companies, gas utilities, as well as those that produce for sale on the open market metallurgical, industrial, and residential-heating grades of coke and coal-chemical materials. The annual coke capacity shown for each plant in this table is based on the minimum coking time necessary to produce coke with qualities suitable for its intended use. Therefore, the annual capacity of a plant may change from year to year, depending on age and condition of ovens, the character and quality of coal carbonized, the grade of coke required, and other factors. For example, annual coke capacity of a coke plant primarily making foundry coke with longer coking cycles would be lower than one making blast-furnace coke. Also some ovens must be operated at slightly lower than normal flue temperature because of the condition of the refractories and maximum capacity would be somewhat lower than designed capacity. For this reason, the capacity reported to the Bureau of Mines may differ from the designed or rated capacity estimated by the coke-oven builders at the time of construction. For example, if the generally accepted standard coking rate of one inch per hour were used on all ovens to calculate total annual capacity, it would have been 80.4 million tons or 2 percent higher than the capacity reported to the Bureau. Because of the factors mentioned above, the annual coke capacity shown in this table is believed to be a reliable measure of the practical operating capacity of the industry.

Jan 1, 1961

By C. R. Hopkins

"The petroleum refiners of the United States made rapid strides in fuel economy in 1928, when the calculated number of B. t. u. required to refine a barrel of crude petroleum was 637,000 as compared with 698,000 in 1927, a decline of 9 per cent. The major portion of this indicated increase in fuel efficiency was undoubtedly due to the gain in use of the pipe-still and the continuous process of refining as contrasted with the old batch method of distillation.,Almost without exception the new refineries are equipped with one or more pipe-stills, and many of the older plants are changing over to this type of equipment. The refineries of the Appalachian area are among the oldest and most conservative of the country, yet a number of them dismantled their old shell and cheese-box stills and installed pipe-stills in 1928.The more extensive use of the pipe-still was not the only way in which the refiners lowered their fuel requirements in 1928. The use of heat-exchangers, in which the heat from hot vapors to be condensed is used to heat the incoming cold oil, also increased in 1928, and in addition the refiners have kept abreast of the times in saving fuel through better insulation and improved burners.For the year 1928 the refiners burned as fuel the following: 46,400,000 barrels of-fuel oil; 3.362,000 short tons of coal; 114,950,000,000 cubic feet of natural gas; 77,313, 000,000 cubic feet of refinery or still gases; and 470,000 short tons of coke, in addition to a little electricity. This fuel was calculated to have produced a grand total of 582,156 billions of B. t. u., as compared with 578,784 billions B. t. u. in 1927.The total of 46,400,000 barrels of fuel oil burned represents an increase over 1927 of about 5,000,000 barrels, which made it about equal to the 1926 figure. Practically all of the refinery districts used more oil as fuel in 1928. The most important increase - over 2,000,000 barrels - occurred in the East Coast district, where greatly expanded throughput of foreign crude was instrumental in increasing fuel-oil production.The use of coal for fuel at refineries again registered a material decline, the total for 1928 of 3,362,000 short tons being 33 per cent below that of the previous year. The outstanding decrease in the use of coal in 1928 occurred in the East Coast district, where coal was probably replaced by oil. The use of coal for fuel at refineries in the States of Ohio, Pennsylvania, and West Virginia declined materially in 1928; the substitute used in the majority of cases was oil burned in pipe-stills."

Nov 1, 1929