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By Lon Santis, Scott Deitchman, John A. Decker

This report describes three cases of carbon monoxide (CO) poisoning in a manhole, including one fatality, from CO migrating through soil after nearby use of explosives. A municipal sewer project involved installation of new pipes and manholes. Work of the general construction contractor was interrupted when a subcontractor detonated 265 pounds of nitroglycerin-based explosive 40-60 feet south of the manhole to break up underlying rock. A construction worker who descended into the manhole 45 minutes after the explosion collapsed within minutes, and two co-workers descended into the manhole to rescue him. One rescuer retrieved the unconscious worker, and the other rescuer died in the manhole. All workers had elevated carboxyhemoglobin levels. In the subsequent NIOSH investigation, air monitoring was conducted with real-time instruments, and air samples were collected in Tedlar bags. Laboratory analyses of the bag samples collected near the bottom showed 1905 parts per million (ppm) CO, 19.5% oxygen, and 3% carbon dioxide. Direct reading instruments showed progressively higher concentrations as the sensor was lowered into the manhole. Subsequent chamber tests on sample explosive yielded 27 liters CO per kilogram detonated. Based on this value, the surface blast at the construction site may have produced about 3,250 liters (114.8 cubic feet) CO. The CO in this incident most likely was released from the nearby explosion and migrated through soil and fractured rock into the manhole. The blasting and construction industries should be made aware of this previously unrecognized CO exposure hazard associated with surface blasting. The extent of CO exposure from explosives used in construction is not known, and additional information on the extent of CO exposure must be collected. In addition, confined space entry procedures (including monitoring confined space atmospheres before entry) should be observed; CO monitoring of confined spaces in the presence of blasting can prevent future incidents such as this one.

Oct 30, 1997

By Clarence O. Babcock

This report describes the results of a Bureau of Mines investigation in which an elastic inclusion, orebody, or mining system is replaced for purposes of analysis with an imaginary effective inclusion of rectangular shape that encloses the region. This effective inclusion has averaged physical properties that are directly related to the average inclusion stresses. The results were in excellent agreement with those predicted by Donnell's theory (2)2 for an isotropic elliptical inclusion in a state of plane strain for all height-to-width ratios and for all degrees of inclusion hardness from very soft to very hard. The results are exact for an isotropic circular inclusion and provide an easy method of evaluating the stresses obtained by cylindrical inclusions in boreholes when the Poisson's ratio for both host and inclusion equals 0.25 for plane strain or 0.333 for plane stress. The theory was also used to predict the average stresses in 48 finite-element models of single transversely isotropic inclusions of ovaloid and rectangular shape. An interesting and useful observation from this work was that the forces on elliptical inclusions in the x and y directions, produced by applied uniaxial stresses in the y and x directions, respectively, were equal in magnitude but opposite in sign when the applied stresses were equal. This occurred when the plane strain condition was used with a Poisson's ratio of 0.25. The practical usefulness of the method is given by treating a variety of mining problems: (1) The average stresses in an orebody of irregular shape; (2) the average stresses in single-level room-and-pillar mining systems; (3) the average stresses in multiple-level room-and-pillar mining systems; (4) the effect of mining sequence on the average stresses produced during mining; (5) the stresses in cylindrical borehole plug related to host rock stresses as a function of plug hardness; and (6) methods for simplifying finite-element analysis of complex structures.

Jan 1, 1974

By R. R. Sayers

Carbon monoxide poisoning is one of the most widely distributed and most frequent causes of industrial accidents . Carbon monoxide gas is a product of incomplete combustion , and since it is without color , odor , or taste , its presence is frequently unsuspected in many places where it exists . It is an ever present danger about blast and coke furnaces , smelters , and foundries . It may be found in buildings having a leaky furnace or chimney , and in buildings where a gas stove is used without a proper flue connection , such as is often found in tenements , tailor shops , and boarding houses . Hospitals receive annually a great number of victims of poisoning , either by accident or in an attempt at suicide , from artificial illuminating gas . People may be affected by leaks wherever water gas is formed or used . The exhaust gases from gasoline motors under average running conditions usually contain 5 to 7 per cent carbon monoxide , and sometimes as much as 13 per cent ; deaths from running an automobile engine in a closed garage are not infrequent . A similar danger may arise in operating gasoline engines in launches without sufficient ventilation . The gas is formed also in stoker rooms , in gun turrets on battleships , in petroleum refineries , and by the Lablanc soda process in cement and brick plants . In underground mines carbon monoxide may appear as the result of shot firing , mine explosions , or mine fires , and in tunnels where automobiles , coal or oil - burning locomotives are operated . In spite of the common occurrence of carbon monoxide poisoning there appears to be no uniformly recognized treatment for a person overcome by carbon monoxide . In the rescue work of the United States Bureau of Mines , however , a method has been developed which has been supported by laboratory investigation , and has proved successful in practical experience , over a period of years . As outlined in this article , the method is useful in the hands of first -aid men as well as physicians .

Dec 1, 1921

By D. Harrington

We Americans pride ourselves that ours is the greatest, most advanced and most civilized nation in the world, yet statistics indicate that apparently we have little appreciation for the lives and limbs of the people whose efforts give us whatever greatness we possess. The latest available statistics indicate that the annual accident death rate per 100,000 of population is nearly 80 or far higher than that of any other country in the world except Chile; is more than double the rate in Hungary, Netherlands, Ireland, Belgium, Denmark, Italy, Germany, Norway, Sweden, and Finland; is nearly double that of England, Wales, the Union of South Africa, New Zealand, Japan, Australia, and Scotland; and is 25 to 40 percent greater than Switzerland, France and Canada. To make matters worse, from the viewpoint of mining; the latest available statistics give mining by all adds the worst accident rate of the major industrial occupations in the United States, the accident severity rate for mining in 1935 being listed by the National Safety Council as 10.14, the next highest rate being 4.52 for construction, and the average rate for 30 main industries being 1.58. IMPROVEMENT IN SAFETY RECORDS Offsetting to a slight extent this rather dark picture is the fact that starting about 1930 the mine-accident rate in both coal and metal mining began to fall, reaching an all-time "low" in metal mining in 1932 and in coal mining in 1933. It now appears probable that the year 1936 has again lowered the accident rate in coal mining and this, together with the fact that the years 1933, 1934, 193 5, and 1936 all set much lover accident rates in coal mining than ware achieved for any similar previous period, gives hope that the trend is definitely in the right direction, though from the above figures mine-accident severity has a long, tedious road to travel before it becomes as low as the severity rate of all other major industries in the United States.

Jan 1, 1937

By Ronald D. Hill, Elmore C. Grim

1. Potential groundwater pollution resulting from large-scale surface mining in western states should be evaluated. 2. Criteria should be developed for the design, construction, operation, and abandonment of sediment ponds for maximum removal of suspended solids. 3. Methods should be developed for reducing suspended solids discharges from surface mines to less than 30 milligrams per liter. 4. A manual should be developed in the form of guidelines for sediment and erosion control planning and implementation. These guidelines must be designed to cover all phases of the coal surface mining operation (i.e., before, during and after mining). 5. A mine road manual should be developed for use by surface mine operators and State and Federal Agencies. Erosion, water control, and dust control should be covered in detail. 6. The block-cut, head-of-hollow fill, and mountain-top removal methods and techniques of contour mining should be evaluated, designed, and ultimately demonstrated in steep, mountainous terrain. Designs should be prepared for single and multiple seam applications. The demonstrations must show that these mining systems will secure rapid achievement of reclamation require¬ments and provide maximum, safe, economic coal recovery in an environmentally safe manner. 7. A Blasting Manual should be developed for use by industry and enforcement agencies. Stress should be placed on how to keep vibration levels and air blast pressures well below accepted safe criteria. Also, methods that can be used to control noise should be listed. 8. Investigations should be conducted on the effect of continued vibrations from blasting on structures that are near quarries. 9. Additional guidelines need to be developed for sampling and analysis of overburden material and associated coals for trace elements and noxious strata. Samples are extremely sensitive, and the source and manner of their collection, handling, storage, and treatment, as well as the techniques of analysis, need to be developed and described. This information can be used to predict possible future problems.

Jan 1, 1974

By D. Harrington

"A tabulation is printed below showing the accident record of the thirty-two mines of the Anaconda Copper Mining Co., in Montana, for the years 1915, 1916, 1917, 1918, 1919 and 1920, with summation for the six years. The table gives the number of shifts worked by each mine with the serious plus fatal accident rate for each 10,000 shifts worked.It is seen that in a total of 16,460,767.5 shifts worked by employees of the company's mines during this six-year period, there were only 2,005 serious and fatal accidents and of these, 200 were fatalities, leaving 1,805 serious accidents (a serious accident being one in which 14 or more days' time are lost). This is at a rate of 9,120 shifts per serious accident and at a rate of 82,304 shifts per fatality. Hence, if 300 days be adopted as a year's work, there would be one fatality per 274.3 years of work in these mines and one serious accident for every 30.4 years work.According to Bureau of Mines Technical Paper 286, (Metal-mine accidents in U. S. in 1919, by W. W4 Adams), during the years 1916, 1917, 1918, and 1919, the metal mines of the United States worked a total of 210,548,640 shifts and had 39,705 serious accidents. During the same years, according to the following tabulation, the mines of the above-mentioned company worked 11,958,978 shifts and had a serious accident rate of 0.995 accidents per 10,000 shifts worked. If the metal mines of the United States, working these 210,548,640 shifts, had this company's serious accident rate of 0.995 serious accidents per 10,000 shifts, the total number of serious accidents would have been only 21,000 or the country's metal-mine serious accidents would have been reduced nearly one-half, and there would be 18,705 fewer serious accidents than actually occurred."

Jun 10, 1921

By L. C. IlsLey

A specially recommended cable is a rubber-sheathed cable that has met the minimum performance requirements for cables as set forth in Schedule 20. The first in the series of recommended cables was listed September, 1930, by an ad- vance notice published as Information Circular 6375 which described the conditions under which the cable tests were made. Since that time six additional cables have met the test re- quirements and have been specially recommended. The complete list of specially recommended. cables is as follows: Specially recommended-cables Cable Symbol No. Hazacord No. 2 BM1 Stranding 19 x 7 Hazacord No. 3 BM2 do. Hazacord No. 4 BM3 do. Super-service No. 3 BM4 7 x 19 Super-service No. 2 Super-service No.. 4 BM5 do. BM6 do. Super-service No. 4 BM7 7x 7 Manufacturer Hazard Insulated Wire Works (division of the Okonite Co.) Do. Do. Rome Wire Company (division of the General Cable Corporation) Do. Do. Do. The cables listed were all twin (parallel duplex) cables having outer belts or cov- erings of approximately -inch high-grade rubber compound, and have the same general appear- ance and relative dimensions as the types of twin cables in general use. The cables were tested under conditions designed to show their approximate perform- ance under mine conditions. Only twin-type cables have met these test conditions. Although the recommended trailing cables are much better and safer for mine use than other cables and withstand being run over by a 7-ton car, they should not be considered as 100 per cent safe. The 7-ton car test represents the minimum limit requirement effective at present under Schedule 2-C, but the test car operates on a straight and level track, and thus does not represent the more severe conditions of mine use such as might result from uneven and curved tracks.

Mar 1, 1931

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

6. Computer Simulation of Mine Evacuation Analysis of mine evacuation involves manipulating a large number of interacting variables, such as mine passages, hoist movements, miner position etc. , which makes the problem very difficult to solve without use of a digital computer. Therefore, under this contract, a computer program is written to simulate evacuation of a mine under different configurations. The simulation input variables include characteristics of the mine, expressed in the form of a network of passages, properties of hoists and hoisting strategies, initial miner positions, ventilation and compressed air velocities, stench warning system and self-rescuer capabilities. The simulation can be carried out for cases in which there is no fire, i. e , fire drill simulation, or for cases in which there is a fire at any arbitrary location. The fire is assumed to generate gases which impede movement of miners in the gaseous region. However, the thermal effect of the fire on ventilation or on the miners is not considered. The simulation output includes the evacuation time for each miner, movement of the hoists, propagation of smoke through the mine passages and smoke exposure time of the miners. Details of the mine evacuation program are summarized in Appendices A and B Appendix A describes structure of the program and its capabilities. The flexibility of the program and its suggested uses are demostrated by a number of illustrative examples Appendix B is a user' s manual for the program which includes instructions for program use and explanation of the logic employed in the program. A computer listing and printout of a typical output are also included in the Appendix B The ten illustrative examples summarized in Section A 4 (Appendix A) provide proof of validity of the program and help make certain inferences about mine evacuation. The first example, which is simulation of an actual mine fire drill conducted in a mine, has predicted evacuation time within 5 percent of the actual evacuation time in the fire drill, demonstrating thereby, the ability of the program to make accurate predictions.

Jan 1, 1976

The following additional studies are recommended: 1. Future development of the magnetic sludge process should involve optimization of the process variables; these include stirring rate, sludge heating time, aeration rate, method of preliminary sludge dewatering, and the effects of magnetic separation for final sludge processing. In addition, the matter of amenability of coal mine waters to treatment should be considered more thoroughly by means of a more comprehensive literature review as well as actual tests on mine drainage samples. 2. The technique of sludge freezing should be explored further. Emphasis should be given to conservation of energy in the process, and the filtration characteristics of the resulting sludges should be evaluated. 3. Further studies should be conducted on the feasibility of the CO2-addition approach to mine drainage sludge densification. The filtration characteristics of product sludges should be evaluated, and the possibility of recovering and recycling the residual alkalinity of the sludge, either directly as CaCO3 or as CaO plus CO2 after calcination, should also be explored further. 4. Efforts should be made to develop cost data for the conceptual processes involving magnetic sludge formation, sludge freezing, and C02 pretreatment, with the purpose of selecting at least one of these processes for further testing at a field site. 5. The use of inexpensive waste materials (e.g. fly ash, "red dog," slag) as sludge additives, possibly in conjunction with coagulant aids, should be considered further. Results of the present studies have indicated that such materials, employed in relatively large amounts, increase the density and enhance the filterability of sludges from lime neutralization of coal mine drainage. The use of pulverized blast furnace slag in this regard appears particularly promising, and studies should be made to ascertain its availability, relative cost, residual alkalinity, and behavior as an additive in conventional mine drainage treatment processes involving lime neutralization and aeration.

Jan 1, 1971

By C. P. Bowie

During the period of 10 years from January 1, 1908, to January 1, 1918, approximately 12,850,000 barrels of oil and 5,024,506,000 cubic feet of gas were destroyed by fire in the United States, entailing a total estimated property loss of $25,254,200. During this period 503 fires were reported. As is shown in Plates I and II, 310 of these fires were caused by lightning and 193 by all other causes. The losses from the fires caused by lightning were estimated to be $11, 148,000, and from those due to all other causes $14,106,200. Directly and indirectly, the fires resulted in the deaths of nearly 150 persons and were responsible for almost as many more being permanently disabled. For the past three years the Bureau of Mines has been conducting investigations to determine the nature and the specific causes of these fires, with a view to suggesting means whereby they may be suc- cessfully combatted and, in some places, perhaps eliminated entirely. As the subject of oil and gas fires is necessarily broad, no attempt is made in this bulletin to treat it exhaustively; rather the bulletin aims to point out what has been done by operators in the past, and to describe various fire-prevention methods and fire-fighting appa- ratus which are being used or adopted by many of the larger oil companies. These methods and apparatus, it is believed, if uni- versally employed by operators, will largely decrease the present enormous annual losses. In the discussion of oil-well fires, the methods used in extinguishing several of the larger and more perplexing ones are described. As the right procedure for extinguishing any oil-well fire must be de- termined largely by local conditions, it is not probable that all of the methods used at some one fire, as described, will be applicable to a future fire. However, the operator, from these descriptions, may at least gain suggestions that will greatly assist him in fighting similar fires.

Jan 1, 1919

By Richard Mainiero, James H. Rowland

For many years there have been small scale tests available for evaluating the toxic fumes production by cap-sensitive explosives (DOT Class 1.1), but these could not be used with blasting agents due to the large charge sizes and heavy confinement required for proper detonation. Considering the extensive use of blasting agent in construction and mining, there is a need to determine the quantities of toxic fumes generated by blasting agents. At the International Society of Explosive Engineers Twenty Third Annual Conference on Explosives and Blasting Technique in 1997, the authors reported on a facility for detonating large (4.54 kg), confined blasting agent charges in a controlled volume that had been constructed at the National Institute for Occupational Safety and Health’s Pittsburgh Research Lab’s Experimental Mine. Since 1997, this facility has been used to collect data on toxic fumes produced by the detonation of various ammonium nitrate/fuel oil (ANFO) mixtures and several cap-sensitive explosives. ANFO composition ranging from 1 to 10 percent (pct) fuel oil have been studied. As expected from previous studies, with an increase in fuel oil content the carbon monoxide production increases, while nitric oxide and nitrogen dioxide production decrease. The detonation velocity varies from 3,000 to 4,000 m/sec for the 1to 10 pct range of fuel oil content, suggesting that ANFO mixes with improper fuel oil content may appear to detonate properly, while their fume production differs significantly from optimum. The study also considers such factors as degree of confinement, water contamination, and aluminum content on blasting agent fume production. Results indicate that water contamination of the ANFO has little effect on carbon monoxide production, but causes significant increase in nitric oxide and nitrogen dioxide production. Decreasing confinement from Schedule 80 steel pipe to 0.4-mm thick sheet metal also has little effect on carbon monoxide production, but significantly increases nitric oxide and nitrogen dioxide production. Adding5 and 10 pct aluminum to the ANFO had no clear effect on carbon monoxide, nitric oxide, or nitrogen dioxide production.

By Linda L. Chasko, Larry D. Stowinsky, Ronald S. Conti

Mine operators often rely on mine rescue teams to save lives during an underground emergency such as an underground fire, explosion or roof fall. It is extremely important that team members are provided with adequate exploration equipment and that training simulations are conducted in a realistic manner. A series of mine rescue training exercises was developed, conducted and evaluated by the National Institute for Occupational Safety and Health (NIOSH) in cooperation with the Pennsylvania Department of Environmental Protection, Bureau of Deep Mine Safety. The exercises were conducted at NIOSH’s Lake Lynn Laboratory during 1995 to 1998 and resulted in improved technology and training for mine rescue teams. For example, existing technologies were identified to help rescue teams during exploration. These included chemical light sticks to identify team members and mark crosscuts and mine materials, strobe lights for mapping out escape ways, and a helmet mounted thermal imaging system to see through smoke. A hands-free communication system showed potential for enhanced communications between team members and the fresh air base. A new high-visibility team lifeline that allows for flexibility of movement between team members was developed. A positive-pressure inflatable escape device (IED) was used to isolate the “hazardous” environment from fresh air and allow rescue team members to traverse through. An inflatable feed-tube partition that can rapidly block large openings such as underground passageways and simultaneously provide a feed-tube for high-expansion foam generators was also deployed for several simulations. During the simulations, mine rescue team members donned self-contained breathing apparatuses (SCBA) and traversed more than 305-m of mine passageways filled with nontoxic smoke. Visibility ranged from 0.3 to 0.9-m. In one area of the mine, a kerosene heater was used to simulate a fire and reduce the oxygen concentration. In another area, a 1.5 pct methane zone was established. The team members searched for "victims,” mapped the passageways, built temporary stoppings and ventilated the smoke-filled entries. The teams indicated that the simulations were challenging and very informative.

By G. R. Hopkins

"Summary An apparent decline in fuel efficiency at refineries, which first became came evident in 1930 and continued through 1933, was checked in 1933 when the average amount of heat needed to refine a barrel of crude oil amounted to 660,000 B.t.u. This compares with an average requirement of 692,000 B.t.u. In 1932 and with a low of 639,000 B.t.u. in 1929. (See fig. 1.) The improvement in fuel economy in 1933 probably was closely related to the increase of 5 per cent in crude runs to stills. As much of the fuel burned at refineries is utilized for purposes other than under crude-oil stills and in relatively stable amounts, it is evident that a material increase in crude runs will usually reflected in a decrease in average fuel requirements. Another factor which has influences fuel economy at refineries in the past is the trend in cracking, a refining process requiring more than the average expenditure of heat. In 1933 production of gasoline by cracking increased slightly, but no appreciable rise in heat requirements resulted as an equivalent amount of fuel was saved in improved technique, particularly in the field of heat exchange.Since the beginning of the refining industry, the transition in fuels has been from solid to liquid to gaseous; in other words, the leadership was first held by coal, then passed to oil, and is now held by gas. In 1933, 54.7 percent of the total heat units generated came from gas, 39.0 percent from oil (including acid sludge), and 6.3 percent from coal and coke. The principal change in the relative importance of the various refinery fuels in recent years has been the rapid growth in consumption of refinery or still gas at the expense of all the other common fuels. Still gas owes its ascendency to its cheapness and availability, and probably it would have become the primary refinery fuel years ago if recovery methods had been perfected. The relative importance of coal showed a small increase in 1933, the first gain recorded since these surveys were inaugurated in 1925. The increase in use of coal was undoubtedly due to increased operations at plants using it,"

Feb 1, 1935

By David C. Byrne

PURPOSE: Evaluate practical technological advances in hearing protective devices for use in different mining environments. RESEARCH SUMMARY: Previous NIOSH research shows that miners develop hearing loss much more quickly than nonoccupationally noise-exposed individuals. Also, miners experience a greater severity of hearing loss than would be expected for nonoccupational noise-exposed persons of the same age and sex. As described in the National Occupational Research Agenda, problems created by occupational hearing loss include (1) the reduced quality of life due to social isolation and unrelenting tinnitus, (2) impaired communication with family members, the public, and coworkers,(3) a diminished ability to monitor the work environment (e.g., warning signals, equipment sounds, etc.), (4) lost productivity and increased accidents resulting from impaired communication and isolation, and (5) expenses for workers’ compensation and hearing aids. Although the elimination of hazardous noise is the ideal long-term solution to the noise problem, miners will likely need to rely on personal hearing protection in certain situations for quite some time. However, it is well known that noise attenuation provided by hearing protectors as they are worn during everyday use in the workplace is usually worse than when tested under laboratory conditions. Additionally, workers may refuse to wear their hearing protection for any number of reasons. Therefore, this project was initiated to evaluate practical technological advances in hearing protectors, with an emphasis toward optimizing the balance between effectiveness, comfort, overprotection, and communication ability. Currently, the necessary in-house laboratory facilities and research techniques are being finalized. Individual experiments are planned that involve analysis of new hearing protector technology under both laboratory and field conditions. The laboratory components are to consist of con-trolled studies using human subjects and/or an anthropometrically correct artificial test fixture to evaluate the adequacy/effectiveness of various hearing protective devices. Field studies are planned to determine the actual usefulness and practicality of the hearing protectors as used in diferent mining environments. Recommendations can then be offered to the mining community regarding correct earplug/earmuff usage that maximizes recognition of hazard/warning sounds while adequately protecting the miners’ hearing.

Jan 12, 2000

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

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

Mar 1, 1989

By Jeffery M. Listak, Ernest A. Curth

"It took only 9 years from the first appearance of roof shields on the U.S. longwall mining scene to the present predominance of shield faces. An apparent consequence is the welcome downward trend in accidents related to failures in longwall ground control. The report addresses load prediction, the effects of shield geometry on support loads, factors contributing to ground control, and related techniques. The Bureau of Mines took an active part in the evolution of longwall mining, including developing one of the first shield faces in 1975, the first lemniscatetype shields in 1976, the Mine Roof Simulator completed in 1980, the first shields in steeply pitching coalbeds in 1981, and the first multilift working of a thick coalbed in 1982. The information offered in this state-of-the-art report will assist in establishing criteria for roof support selection. INTRODUCTIONIn terms of roof support technology, the introduction of shields in the seventies and eighties was a major step forward. Table 1 indicates the predominance of shield faces on the U.S. longwall mining scene and its apparent consequence, a welcome downward trend of ground-control related accidents.The rapid progress of longwall technology began in 1975 and is highlighted by a development leading from the caliper shield to the modern lemniscate type. Shields have been gaining preference over chocks since their advent to longwall mining in 1975. Safety and productivity factors favoring shields over chocks are--1. A sheltered working space requiring minimum cleanup work.2. Structural stability that allows advancing without delay even with brushing roof contact and, by controlling all lateral loads, removing the requirement for cumbersome leg restraint and restoration devices.3. Wide hydraulic range of mining height.The development of roof shields in the Federal Republic of Germany preceded and paralleled their rapid adoption by U.S. miners. In 1983, 87 pct of the total production in Germany came from shield faces (1).2 In Great Britain, the National Coal Board introduced the first shield support system in 1977 to initiate Advanced Mining Technology (ATM) and Heavy Duty Mechanization (HDM) programs; as of mid-1982, 44 shield faces were operating, or 7.5 pct of a total of 581 faces CJ_)."

Jan 1, 1985

By John J. Sammarco

Computer-related accidents have caused injuries and fatalities in numerous applications. Normal Accident Theory (NAT) explains that these accidents are inevitable because of system complexity. Complex systems, such as computer-based systems, are highly interconnected, highly interactive, and tightly coupled. We do not have a scientific methodology to identify and quantify these complexities; specifically, NAT has not been operationalized for computer-based systems. Our research addressed this by operationalizing NAT for the system requirements of safety-related computer systems. It was theorized that there are two types of system complexity: external and internal. External complexity was characterized by three variables: system predictability, observability, and usability -the dependent variables; internal complexity was characterized by modeling system requirements with Software Cost Reduction dependency graphs, then quantifying model attributes using 15 graphtheoretical metrics the independent variables. Dependent variable data were obtained by having 32 subjects run simulations of our research test vehicle: the light control system (LCS). The LCS simulation tests used a cross-over design. Subject perceptions of these simulations were obtained by using a questionnaire. Canonical correlation analysis and structure correlations were used to test hypotheses 1 to 3 - the dependent variables predictability, observability, and usability do not correlate with the NAT complexity metrics. Five of 15 metrics proposed for NAT complexity correlated with the dependent data. These 5 metrics had structure correlations exceeding 0.25, standard errors < 0.10, and a 95% confidence interval. Therefore, the null hypotheses were rejected. A Wilcoxon signed ranks test was used to test hypotheses 4 to 6 - increasing NAT complexity increases system predictability, observability, and usability. The results showed that the dependent variables decreased as complexity increased. Therefore, null hypotheses 4 to 6 were rejected. Lastly, this work is a step forward to operationalize NAT for safety-related computer systems; however, limitations exist. Opportunities addressing these limitations and advancing NAT were identified. Lastly, the major contribution of this work is fundamental to scientific research - to gain knowledge through the discovery of relationship between the variables of interest. Specifically, NAT has been advanced by defining and quantifying complexity measures, and showing their inverse relationship to system predictability, observability, and usability.

By Thomas M. Barczak

The safety of mine workers depends on the proper installation of roof supports to prevent the ground from collapsing into the working areas of an underground mine. As new support systems are developed, they need to be properly evaluated to make sure that they are capable of providing adequate roof support before they are first used in a mine. In addition to making certain that the supports meet basic safety criteria, the limitations of the support need to be fully defined in order to avoid improper application of a particular support design. The National Institute for Occupational Safety and Health (NIOSH) operates a world-class facility called the Safety Structures Testing Laboratory. This laboratory contains a unique load frame, the Mine Roof Simulator, which is capable of simulating the ground behavior in underground mines for conducting full-scale evaluations of roof support systems. Safety performance testing protocols using the unique Mine Roof Simulator have been developed for both standing roof support systems and longwall shield supports. The purpose of this paper is to describe these test procedures. The protocol for standing roof supports incorporates seven test series:( 1) uniform loading baseline tests, (2) height evaluations, (3) asymmetric loading, (4) biaxial loading, (5) load rate studies, (6) active loading determination, and (7) static loading evaluations. For longwall shields, a four-series test program that accurately simulates in-service conditions on a longwall face is proposed. This test program consists of (1) transfer of horizontal load to the caving shield-lemniscate assembly (zero-friction test),(2) point loading of shield joints due to lateral movement or rotation of the canopy, (3) evaluation of leg socketand leg cylinder integrity, and (4) face-to-waste racking of the shield. In addition, an evaluation of the shield’s hydraulic components will be conducted prior to the performance testing. These protocols will provide state-of-the-art safety performance evaluations of emerging support technologies, as well as a means to assess the safety of both new and aging longwall shields. Hence, this effort will enhance the safety of mine workers by ensuring that critical support elements are properly designed and that aging supports are retired before their support capability is jeopardized.

Jan 10, 2000

By Maria I. De Rosa

Table 1 and figure 1 show the number of fires and fire injuries that occurred in underground metal/nonmetal and stone mines by state during 1990-2001. Table 1 also shows the injury risk rates, employees&apos; working hours, and lost workdays. Overall, 65 fires occurred in 20 states; these include 2 fires and no injuries for contractors. Six of the fires caused nine injuries. The yearly average was 5.4 fires and 0.75 injury. Forty-one fires with 2 injuries occurred in metal mines, 14 fires with 7 injuries occurred in nonmetal mines, and 10 fires with no injuries occurred in stone mines. The underground mine fires required 25 mine rescue team interventions and 30 mine/section evacuations. The Ewhr value was 260 x 10° hr (Irr = 0.007), and the LWD value was 83. Idaho had the most fires (eight fires and no injuries), followed by Louisiana (seven fires and two injuries), Michigan (six fires and six injuries), and Missouri (six fires and no injuries). Of these states, Michigan had the highest injury risk rate value (Irr = 0.146). Table 2, partly illustrated in figure 2, shows the number of fires, fire injuries, risk rates, employees&apos; working hours, and lost workdays by time period. The number of fires during the six time periods show an increase during the fourth period followed by a sharp decrease during the fifth period and a sharp increase in the last period. The number of fire injuries show an increase followed by a decrease during most of the periods, accompanied by a decline in employees&apos; working hours during most of the periods; an increase is seen during the third and fourth periods. The In values follow patterns similar to those shown by the injury values. Tables 3-8 show the number of fires by ignition source, method of detection and suppression, equipment involved, location, and burning material by time period. Figure 3 shows the major variables related to fires for 1990-2001. Table 9 shows the number of fire injuries per number of fires causing injuries and total fires by year, ignition source, equipment involved, and location.

Jan 1, 2004

Although the technology involved in removing material from below the earth&apos;s surface has a long history, communication systems in underground workings are relatively new to the industry. Communication equipment did not begin appearing in underground mines until the early 1900&apos;s. Figure 1-1 shows a miner using a Western Electric standard telephone set for underground mines in 1913. These early phones were essentially the same as those used aboveground, except that they were enclosed in cast-iron boxes as protection against moisture, acid fumes, and gases. In the 1950&apos;s, the Chesapeake and Potomac Telephone Co. of West Virginia introduced a telephone set for use in explosive atmospheres that was designed around the philosophy of explosion containment. To contain an explosion within the telephone set required a 39-pound casing, which greatly limited portability. [ ] Both of these telephone sets required "pipe" or conduit, not a very practical item for a 100-squaremile coal mine. A common thread prevalent in the first 40 years of design and development is that the primary effort was placed on the telephone set. In the early 19701s, as work began on the design of modern communication systems for use in underground mine environments, the following requirements were established: 1. Must meet intrinsic safety standards .--This applies not only to coal mines, but to other explosive environments. 2. Must be compatible with the environment.--The system must with- stand dust, moisture, and corrosive conditions. 3. Must be rugged in structure.-- The system must withstand maintenance by 10-inch pliers and 4-pound hammers, as well as impact from a falling piece of roof. 4. Must be size-flexible.--Whatever is built must be sized for the small operator as well as the large companies. 5. Must be a total system in design.--The system must be intrinsically safe, including cable, power supply, and station set. 6. Must work with present telephone system.--The system should be compatible with aboveground communications already in place; it should not be the single cause of change in aboveground communications. 7. Value-added pricing.--The system should be reasonably priced so that savings created by its introduction will more than offset the installation cost.

Jan 1, 1984