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To evaluate the potential for dust reduction, during blasthole drilling, of water or foam applied directly into the drill hole. Approach The Bureau of Mines tested a through-the-drill-steel water and foam injection technique in two noncoal mines to evaluate the method's potential for greater dust reduction. The tests were performed by drilling a series of blastholes for an entire face (a 90 minute time-period). For the first 45 minutes of drilling, foam or water was applied through the drill steel. Blowing ventilation was supplied to the entry about 50 feet from the face. The dust concentration measurements were made by a Real-Time Aerosol Monitor (RAM-I), and the data were read out on a strip chart recorder. The evaluation test results showed that the water or foam injection technique, when used on localized dust sources, can be an extremely effective dust control method. Small amounts of water used in this method have also allowed it flexibility to be used in water-soluble ores such as salt, gypsum, anhydrite, and trona. How The System Works Water or foam is mixed with compressed air (approximately 100-120 psi) with suitable metering systems and forced through the drill steel into the blast hole at the point of dust generation. Because the water or foam is applied at the source, minimal quantities of liquid are needed (0.2 to 0.6 gpm). These low flowrates are sufficient to just moisten the dust emerging from the blast hole, thus eliminating a major source of airborne dust. The wet drilling procedure also eliminates a second source of dust. With dry drilling, the dust from the blast hole falls to the floor. Since this can be a 15-ft drop or more, the impact can send the dust billowing up from the floor, adding to the airborne dust. With wet drilling, the moistened dust merely dribbles down the face.

Jan 1, 1982

By H. P. Greenwald, G. E. McElrey

"Part IV.- The Resistance Caused by Right-Angle Bends. IntroductionThis paper deals with the fourth phase of an extensive research on coal-mine ventilation factors. A comprehensive report of the entire work is in course of preparation, but abstracts of the various chapters are being prepared in order to make the major results public as rapidly as possible. Parts I, II and III were issued as Serials Nos. 2621 (July 1924), 2647 (October 1904) and 2671 (March 1925) and dealt with the resistance of unobstructed entries, the resistance due to mine cars and, the resistance of timbered entries respectively.Description of the Tests With the last cutthrough between the two butt entries of the test section in the Experimental Mine open, the air flow was deflected, around two right-angle bends in the same direction, the butt entries being 50 feet center to center and approximately 6 by 9 feet in cross section. Each bend had a dead end opening off of it; that at the first bend was in the direction of flow and that at the second bond was at right angles to the flow. For a primary test condition the inner corner of each bend was trimmed square and the dead ends were closed off with board brattices set flush with the line of the cut-through.The two corners of the first band were, then modified both independently and in conjunction. Various modifications of both bends in conjunction were also tested with rather surprising results as to the effect of conditions at the first bend, upon the pressure losses at the second. To confirm the theory developed to fit these results a series of tests were made to give a direct comparison of the -pressure losses as between two bends with the air flow deflected in the some direction at each and with the air flow deflected in opposite directions. The former is designated ""double"" and the latter a ""reversed"" right-angle bend. This series of tests was conducted, however, on the bends at one of the intermediate cutthroughs at which the flow could be deflected us either direction at the second, bend. Here also the average cross-section was 6 by 9 feet and the bends were 50 feet apart."

Jan 1, 1928

By Raja V. Rarnani, Robert L. Frantz

"Scientific, engineering, and biomedical research teams synergystically developed major topical research areas in the Respirable Dust Center. Descriptions of their interactions and resulting accomplishments demonstrate how cooperative research within and among these study groups benefits the overall respirable dust research and practical application needs of the mineral industries.The development of a state-of-the-art small animal inhalation exposure facility suitable for studies of fresh and aged coal mine dust, a major goal of this research group, will contribute toward the GTCRD goal of advancing the understanding of causative mechanisms for the occupation-related lung impairments and the airway diseases of mine and mill workers. Such understanding includes improved definition of the role of various aspects of the lung environment (e.g., surfactant and the mucociliary system) in the disposition and toxicity of respirable dust, comprehensive characterization of the role of the pulmonary alveolar macrophage (PAM) and its secretory products on the development of dust-related lung disorders, and an expanded search for methods to reduce the toxicity of respirable dust both before and after it enters the lung. The National Academy of Sciences has recommended, in Measurement and Control of Respirable Dust in Mines, that respirable dust research should include study of the behavioral and constitutional factors affecting individual susceptibility to workers' pneumoconiosis occupational disease and the development of methods for detecting the potential onset of pneumoconiosis earlier and more conclusively than can be done using X-ray methods. These recommendations have guided the research group in its goals of establishing and disseminating a scientific protocol for standardizing the inhalation testing for all research objectives within the Generic Center which involve small animal exposures and implementing the infrastructure necessary to make lung tissues obtained during inhalation studies available to all researchers within the Generic Center."

Jan 1, 1995

By Michael J. Sapko, Samuel P. Harteis, Eric S. Weiss

The Pittsburgh Research Laboratory (PRL) of the National Institute for Occupational Safety and Health (NIOSH) recently conducted full-scale explosion experiments; these experiments evaluated the strength characteristics of various seal designs used for safely isolating worked-out areas in underground coal mines. Large-scale explosion tests were conducted within the multiple entry section of PRL’s Lake Lynn Experimental Mine (LLEM) employing the only currently accepted test method endorsed by the Mine Safety and Health Administration (MSHA) for seal design in American mines. These explosion tests are labor-intensive, expensive to conduct, and can interfere with other critical underground safety and health research programs conducted by NIOSH. Therefore, the PRL has developed an alternative seal evaluation method, based on a hydrostatic pressure loading concept, that can facilitate the in-situ testing of seals in an operating mine. Two chambers within LLEM and one within the Safety Research Coal Mine (SRCM) were used for hydrostatic pressure loading various seal designs. The results from tests in these chambers compare favorably with those from the large-scale explosion tests in the multiple entries. Preliminary size-scaling relationships for predicting the strength of standard seal designs as a function of entry size is also presented. In addition to testing seal designs at the required 20 psi static pressure level, the new facilities also allow for the determination of the seal’s ultimate failure pressure. This new approach shows promise as an alternative evaluation method for improving mine sealing technologies which, in turn, will enhance the safety of the underground personnel.

By Arthur L. Murray

"Putting safety over , " using the accepted meaning of the word " Safety " as the elimination of personal injury , resolves itself into the prevention or disassociation of a combination of two distinct circumstances . Opportunity for accidents is afforded by conditions that are likely to cause injuries , but such injuries are effected only when the human body is placed in a position to be injured . As will be discussed later , in mining , as in most other industries , the degree of safety experienced depends upon the disassociation of these two circumstances , first , by making working conditions as free from hazards as possible , thus affording the minimum opportunity for accidents to occur , and secondly , the training of employees to recognize dangerous conditions and adopt measures to protect themselves and their fellow employees against injury . The responsibility for preventing accidents rests upon every person employed in an industry from the directing head down through the individual workers , and the safety of that industry will be proportionate to the extent that each member appreciates and accepts his individual responsibility . Hence , safety becomes decidedly a personal equation . Whether the personal equation will be so applied as to result in the greatest freedom from accidents , however , depends largely upon the initiative of those directing the industry . Putting safety over in a small mine presents the personal equation in a more intensified manner than the same endeavor in a large mine . Large mines , owing to their size and financial standing, can usually main ta in and support an effective safety organization with one or more directors who devote all their time to the supervision of safety work . Most small mines with a limited number of employees and limited capital can not afford and are not justified in supporting a paid safety organization , and must depend upon the close cooperation of their working force in maintaining safe working conditions . On the other hand , if proper supervision is exercised the small mine has in some respects advantages not afforded by the large mine . The limited workings in a small mine should afford less opportunities for dangerous conditions to arise and permits more frequent inspection of the entire mine . The small working force permits more latitude in the selection of employees known or believed to be either safe workers or amenable to safety educational work. Likewise , a small working force facilitates the carrying on of safety educational work and allows a closer personal contact between the superintendent and his employees .

Mar 1, 1923

By Charles E. Munroe

"It has been recently learned that certain foreign manufacturers are offering far importation into this country detonators at greatly reduced prices, some of these detonators being of such low grade (No.4), that their use would greatly increase the dangers in blasting.Detonators and electric detonators are used in initiating the explosion of the various high explosives that are now so extensively employed in a great variety of mining operations. Originally a large number of different grades (or ""strengths"") of detonators and electric detonators were marketed, the grade number indicating the weight of the charge of fulminating material contained in the copper case of the detonator. As late as 1914, a report of the Bureau of Mines (Bulletin 80, A primer an explosives for metal miners) shows that grades Nos. 3, 4, 5, 6, 7, and 8, containing from 0.54 to 2.00 grams each, were being offered for use. However, the writer bad observed as early as 1896, when the more insensitive types of explosives like nitro-substitution explosives began to be introduced for use in mining and other blasting operations, that use of a high-grade detonator was essential to insure certainty of discharge."

Mar 1, 1921

By D. S. Kingery

The practice of using doors to control main ventilating currents in bituminous-coal mines is common throughout the industry. When doors are erected in pairs to form air locks, they are usually accepted as satisfactory by both State and Federal coal-mine-inspection agencies. A review of the contributing factors in recent methane ignitions showed that interrupted ventilation (often caused by doors latched open, destroyed, or damaged) was the principal cause of methane accumulations that subsequently were ignited.

Jan 1, 1958

By Richard L. Lewis

This is a discussion of the application of the zonal harmonic series to problems involving dipole sources in the vicinity of spherical boundaries. The application of a formal zonal-harmonic series solution to directly obtain numerical results involves two distinct problems: namely, the convergence of the series solution and the accurate calculation of the special functions involved in the series solution. Some interesting techniques will be introduced which help to improve the convergence properties of the series solution.

Jan 1, 1973

By Fred N. Kissell

Tracer gas decay curves were experimentally obtained in a full-scale model coal mine working face area. The concentration of tracer gas was found to decay in exponential fashion but with a change in slope. This change in slope is attributed to eddy zones in the face area. A simple theory was developed that allows one to calculate the approximate amount of gas in the eddy zones and thus estimate the size of the eddy zones.

Jan 1, 1974

By Lon D. Santis

This paper presents data and theorem to calculate the risk of experiencing undesirable lightning related events while blasting or while engaged in other lightning sensitive activities. The paper provides an overview of lightning hazards in blasting operations and a brief history of lightning related blasting accidents. Lightning continues to be the primary cause of premature initiations of explosives in mining, on average, over two such incidents are reported to the Mine Safety and Health Administration (MSHA) each year. The energy contained in lightning discharges and how this energy sets off explosives in mining is presented. Three categories of lightning warning methods: public media, lightning detectors, and atmospheric electrostatic field measurement are discussed. The remainder of the paper presents recently acquired data on the spatial characteristics of lightning development. This data is used to estimate the risk assumed by blasters under certain conditions. Risk estimations are made for electric and nonelectric blasting if persons are evacuated when cloud-to-ground (c-g) lightning approaches within a certain distance. For example, about 0.6% of all c-g lightning strikes have no other c-g lightning strikes within 25 miles in the previous 30 minutes. The paper shows that a typical blaster using nonelectric initiation in a high lightning density area with a 100% accurate lightning detector set to alarm at 25 miles would have a lightning caused unplanned detonation every 14,000 years

By C. Travis Anderson

 The general conclusion of earlier investigators was that the action of sulfur dioxide on manganese oxides of Cuyuna ores was not sufficiently pre- ferential to be attractive as the basis for a commercial process. Bradley4/ and his coworkers discarded it for leaching with boiling ammonium sulfate solution, and Sweet5/ used a baking operation to render the iron insoluble. These results were confirmed by the early tests in the present investigation, which were made on brown ores (Sagamore), using acid additions, because at that time the Technologic Committee on Manganese held the opinion that these ores constituted the major available reserve in the Cuyuna district. However, when it appeared that the black ores, or the formation in which they occur, were a possible source of a large tonnage of manganese, the use of sulfur dioxide and sulfuric acid leaching was again reviewed, and a satis- factorily preferential action on the manganese was found. It was originally thought that the use of sulfuric acid ahead of, or at least with, the sulfur dioxide simplified the process, particularly by preventing formation of dithionates. Removal of the phosphorus from the solution by neutralization with lime or fresh ore was suggested. The tests recorded in this report indicate that it is not practicable to use enough acid to prevent formation of dithionate and that the value of excess sulfuric acid is doubtful, particularly in regard to satisfactory manganese:phosphorus and manganese:iron ratios.

Jul 1, 1942

By J. A. Zimrner, R. J. Tirnko

The roles of main mine fans in underground mines are to induce airflow and continuously re- move hazardous gases and dust. While most larger mines use multiple fans to accomplish these tasks, many smaller mines employ only a single fan. This paper concentrates on those mines having only one fan to provide ventilation needs. If this fan should cease to function, it is likely that methane concentrations will in- crease in some areas of the mine. As a safety precaution, Federal Regulations require that personnel must be- gin evacuating the mine within fifteen minutes after fan stoppage. Powered haulage can be used to transport workers to the mine portal as long as travel is through areas where hazardous levels of methane are not expected. To determine where methane will accumulate along haulageways during a fan stoppage, air samples must be taken at specific areas and analyzed for hazardous gases. During this study, procedures were developed for measuring methane levels along the haulegeways that are used to exit a mine following stoppage of the main mine fan. Methane levels were measured in four different mines at selected underground locations during four planned fan stoppages. In these mines the measurements taken only at the specified locations showed that, if workers had used electrically powered haulage to exit the mine following a fan stoppage, the methane accumulated along haulageways during fan stoppages would not have significantly increased the risk of an ignition.

By D. K. Ingram, G. L. Mowrey, T. R. Rehak

This paper summarizes fly rock and blast area security fatalities from 1989 to 1999 and examines the causative factors. Coal and nonmetal mining used about 43 billion pounds of explosives and blasting agents between1989 to 1999. A majority of this consumption was used at surface mines. Accident data indicates that fly rock and lack of blast area security were the primary causes of blasting related injuries in surface mining. Fatal injuries due to lack of blast area security were attributed to: failure to clear blast area; failure to follow instructions; inadequate guarding; inadequate blasting shelter; and unsafe location. Seven fatalities due to fly rock and lack of blast area security were reported in surface coal mines and six in nonmetal mines. Out of these, two fatalities (one each in coal and nonmetal mines) occurred outside the mine property. In the coal mining sector, two fatalities were attributed to fly rock, and five to lack of blast area security. In the nonmetal mining sector, three fatalities were attributed to fly rock, and three to lack of blast area security. Preventive measures include: ensuring that all personnel have evacuated the blast area during shot firing; using adequate blasting shelters for employees whose presence is required in the blast area; controlling and monitoring all entrances to the blast area; ensuring that the blast is properly designed, drilled, and loaded; and emphasizing education and training to enhance skill levels for implementation of engineering control techniques.

By R. V. Ramani, V. T. Burgos, J. A. McClay

Cost Systems Cost systems are organized methods of collecting, allocating, and accumulating resource utility expressed in terms of money. There are two major spheres of cost accumulation: financial accounting and cost accounting. Financial accounting deals primarily with historical costs and provides a means of measuring and recording cost events which have al¬ready taken place. While there are no legal requirements to control the methods of recording and reporting cost data in the United States, the procedures in use are guided by professional organizations such as the American Institute of Certified Public Accountants. The final pro¬ducts of financial accounting practices are the income statement and balance sheet which describe the financial status of the organization under study. Cost or managerial accounting is an expansion of financial ac¬counting procedures, providing more detailed cost breakdowns on present operations with one objective being the projection of present costs into the future. The projection process tends to limit the accuracy of the information; however, to management and project justification per¬sonnel, data on future operations tends to be of greater value. The basic principle behind cost accounting is the cost cycle. As products are developed from raw materials, the value of the work-in¬process increases due to the cost input of the various fabricating ele¬ments. Thus, from the basic input elements of facilities, labor, and raw materials, costs are added to each unit of production as a function of the work performed on that unit.

Jan 1, 1975

Design, fabricate, and test a machine capable of operating in coal seams 48-inches and thicker that can reduce the time, expense, and injury potential inherent in extending and retracting mine-wide section belt conveyors, while eliminating the need to use additional machinery or equipment in accomplishing such moves. The Problem Many underground coal mines use a belt haulage system to move cut coal from the face to the loading area. This system usually consists of a main line belt conveyor fed by section belt conveyors. A shuttle car at the face carries the coal cut by a continuous mining machine to a feeder breaker that breaks the large lumps of coal into smaller, transport-able pieces and unloads them onto the tail piece of the section conveyor. This tail piece is skid-mounted so it can be readily moved as needed. The conveyor is kept close to the face during mining in order to shorten the haul distance of the shuttle cars. Section belts can be lengthened by adding belting, wire rope, and support hard-ware. During pillar retreat mining, these section belts are shortened to keep pace with pillar extraction operations. However, these belt moves, either the extensions or retractions, usually require six to eight workers and three to eight hours to complete. In addition, a 100-foot belt move requires workers to handle up to 4,000 lbs. of material, not including the tail section. Strain and sprain injuries can be incurred during this heavily labor-intensive task. The time required to move the section belt can also create losses in the mine's production. To help solve this problem, the Bureau of Mines

Jan 1, 1984

By B. J. Scheiner

As part of its mission to reduce pollution associated with mineral production, the Bureau of Mines is developing a dewatering technique for phosphatic clay waste that uses a flocculant, polyethylene oxide (PEO). This flocculant forms strong, stable flocs, which can be partially dewatered on a static screen and dewatered further on a rotary screen. Using this technique, a field test unit (FTU) was operated at a nominal 100 gal/min. Consolidated phosphatic clay material containing 20 percent solids was produced when feed slurries of 3 percent solids were treated with 0.69 pound of PEO per ton of feed solids. Pit and column tests indicated that the PEO-treated material continued to dewater; compacted products containing 40 percent solids were obtained after 45 days.

Jan 1, 1981

By Everett A. Sondreal

Material and energy balances for stockpiled lignite were calculated by computer simulation techniques to determine the occurrence of spontaneous heating. Data on properties of lignite which affect spontaneous heating were required. In some cases, best values from reported data were utilized; in other cases, new data were developed. These include specific studies on the rate and conditions that influence oxidation, heat of oxidation, heat of wetting, specific heat, thermal conductivity, and permeability of stockpiles to air. These data are presented and discussed in detail for possible application to other studies. Results of computer simulations confirm past field experience on the storage of lignite containing as-mined moisture, and predict a greater risk of heat occurrence in stockpiles of dried lignite where experience is lacking. Successful storage was predicted for dried lignite of 20 pct moisture content, if it were cooled after drying and compacted to an interparticle void fraction of 9 pct (bulk density of 67 lb/cu ft) compared to 10 pet void fraction (72 lb/cu ft) for as-mined lignite.

Jan 1, 1974

By H. W. Gillett

This bulletin is issued by the Bureau of Mines as a contribution to the increase of safety and efficiency in the preparation and utilization of the mineral resources in the United States. Notable among the drains on these resources are the demands made by the nonferrous- metal industries, and in these the losses occasioned in melting brass and bronze are worthy of particular attention. This statement will be better appreciated by a consideration of the size of the brass industry.

Jan 1, 1914

By Isaac A. Zlochower

A number of ignition incidents involving mining and farm equipment have illustrated the surprising danger of drilling, cutting, or otherwise penetrating sealed compartments or frames that are filled with metal scrap for ballast. The presence of chemically-reactive metals (or electrochemically-active pairs), such as zinc in galvanized steel, can produce hydrogen from moisture in the steel scrap, and the resulting pressure, due to the internally-generated hydrogen, can produce an easily-ignited gas jet when the sealed component is penetrated.

By Powell. Alfred R.

"An investigation has been made of the applicability to a variety of coals, other than those from Illinois, of the Powell and Parr method of analysis for the forms of sulphur in coal.A complete study was made of the sulphur in the following coals:- Upper Freeport, Pa.; Pittsburgh seam, Pa.; Pocahontas No. 3, W, Va..; Letcher Co., Ky.; Morgan Co., Tenn.; and Cherokee Co., Kan.The method of analysis gave excellent results when applied to these coals. In the determination of pyrite, the iron-sulphur ratio checked in every case with the theoretical. After the extraction of the pyrite and sulphates, the remaining sulphur was proved to be organic in nature.This work indicated that sulphur exists in coal in three typical forms: - pyrite or marcasite, sulphates, and organic sulphur. A study of the changes which these forms undergo during coking has been made on a variety of coals, and the following five classes of reactions established.First. Complete decomposition of the pyrite and marcasite to ferrous sulphide, pyrrhotite, and hydrogen sulphide. This reaction begins at 300°C., is complete at 600 °C., and generally reaches its maximum between 400 and 500°C."

Mar 1, 1920