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By David A. Monaghan, Kenneth R. Maser, Adi R. Gurdar, D. Randolph Berry

APPENDIX E EQUIPMENT USED FOR FIELD TESTING OF MINE SEALING SYSTEM I. Drilling Operation - Equipment supplied by local drilling contractor 1. Mobile drill rig. 2. Flat bed truck, containing 20 lengths of drill pipe. 3. Auxiliary compressor. 4. Pickup truck for small parts and supply of diesel fuel. 5. Bulldozer. 6. Water supply, in this case the nearby stream. 7. Supply of casing, positioned next to the drill rig. 8. Ready mixed cement grout. 9. Drill bits, stabilizer units, cutting and welding equipment, other miscellaneous tools and supplies. II. Sonar Probing Operation 1. 25 foot tower with pulley mounted atop. - FMA 2. Electric winch with 100 feet of 1/4-inch cable, Model No. 100A, City Engineering Co. , Indianapolis, Indiana. 3. 400 feet of 1-1/2-inch-diameter aluminum pipe with alignment holes and coupling grooves. - FMA*

Jan 1, 1973

By Richard J. Mainiero, Eric S. Weiss

This project investigates the behavior of blast waves from the detonation of high explosives in an underground mine. A series of explosive tests was conducted in the underground and surface facilities at the Bureau of Mines' Lake Lynn Laboratory to evaluate the potentially dangerous effects of blast waves produced by open shooting, blown out holes, or accidental detonations of explosives. Shots of C4 and TNT were fired at the dead end of a 6 ft high by 20 ft wide entry, at the intersection of two entries, and on the surface. C4 and TNT were chosen for this research because they exhibit very consistent performance. Behavior of the blast wave was evaluated through recording of the pressure as a function of distance from the explosive detonation. For both explosives, C4 and TNT, peak pressure decay with travel down the entry was inversely proportional to distance to the 0.9 power. When comparing pressure measurements at 200 versus 100 feet from the detonation; at 200 feet the pressure would be 1/2 0.9 or 54 percent of that measured at 100 feet. By comparison, peak pressure for a surface shot exhibits a decay with distance to the -1.6 power, yielding a pressure at 200 feet 1/2 16 or 33 percent of that measured at 100 feet. This result confirms that the blast wave from detonation of an explosive in an underground mine maintains its force for a greater distance than would be the case on the surface. Table and figures follow the text.

By T. H. Jeffers

Significant amounts of cobalt, a strategic and critical metal, are present in readily accessible spent copper leach solutions. For example, recovery of cobalt at two major u.S. copper operations could produce about 1,300,000 lb Co annually, about 13 pct of domestic consumption. However, techniques such as solvent extraction and precipitation have not proven cost effective in separating and recovering the cobalt from these low-grade domestic sources. The Bureau of Mines has devised a procedure using a chelating ion-exchange resin to extract cobalt from a pH 3.0 copper leach solution containing 30 mg/L Co. Cyclic tests in 4-ft-high by 1-in-diam columns gave an average cobalt extraction of 95 pct when 65 bed volumes of solu-tion were processed at a flow rate of 4 gpm/ft2 of resin area. After an impurity scrub, eluates contained, in grams per liter, 0.4 Co, 0.2 Ni, <0.001 Cu, 1.3 Fe, 2.8 Zn, and 0.008 AI. Solvent extraction procedures to remove impurities, reject nickel, and concentrate cobalt produced a cobalt sulfate solution containing 80 g/L Co, 0.05 g/L Ni, and not more than 1 mg/L Cu, Fe, Zn, or AI. Based on data from commercial electro-winning operations, this solution appears suitable for production of metallic cobalt.

Jan 1, 1985

By Peter. PH. D. Cain

Smoky River Coal Ltd. mines low-volatile metallurgical coal by surface and underground methods in the foothills of the Rocky Mountains of Alberta, Canada. Current underground operations are confined to the 5B-4 Mine. Development of 5B-4 began in January 1998; production from depillaring sections commenced in July 1998. This paper describes the history of underground mining on the Smoky River property in terms of extraction methods and pillar design. The development of the present pillar design guidelines is discussed in this context. Recent work to prepare a number of case histories for back-analysis using the Analysis of Retreat Mining Pillar Stability (ARMPS) method is described, along with the modifications developed for calculating the ARMPS stability factor for retreat extraction of thick seams. The design criteria are described, as well as the geotechnical program implemented in order to verify its applicability.

Jan 5, 1999

By R. V. Ramani, P. M. T. White, D. Sutton

4.1 Introduction Using methods outlined in previous sections of this report, programs were written and/or obtained to: i) select best combinations of "a" and "k" for up to 8 different value parameters, ii) calculate available and unavailable ore tonnages and other statistics, and iii) produce contour maps of various parameter characteristics and available tonnages on a line printer. All the above programs are written in Fortran IV for execu¬tion on IBM 360/370 computers. 4.2 Calculation of Coal Reserves The Ore Reserve program suite consists of the following sections: (a) UNIT 1: This unit is a subroutine called by Unit 2 to determine values for "a" and "k" from the given data. (b) UNIT 2: These programs calculate ore reserves and are composed of a main program and several sub¬routines. In the following, their functions are abbreviated: i) MAIN - calls subroutines, reads run parameter variables, reads borehole data and interpolates grid node values for the area of investigation

Jan 1, 1974

By E. J. Kloos

The Bureau of Mines devised a new machine-test method for measuring carbon dioxide inspired by a wearer of self-contained breathing apparatus. The new method compares favorably in speed and precision with standard man-tests in which the apparatus is tested while actually worn, and is scheduled for use in future approval tests for facepiece carbon dioxide &apos;measurements. However, man-tests will still be used to measure the efficiency of the carbon dioxide sorbent materials in closed-circuit breathing apparatus.

Jan 1, 1966

By McHenry Mosier

Although local conditions at each mining property affect the percentage of increase in drilling speed for a given reduction in gage of bit, the results of drilling tests conducted at the Mount Weather testing adit show such a large advantage in favor of the smaller gages that they should be indicative of the possibilities at most mines. GENERAL DISCUSSION Although it is common knowledge among miners that drilling speed ill- creases with reduction in gage of bit, the large amount o this increase is not so universally realized, A recent Bureau of Mines4/ survey of drilling practices in 28 metal mines showed that the maximum gage of starters is 3 inches and that at 50 percent of the mines starters are 2-1/4 inches or larger, Because this size appears too large for conditions at most&apos; mines where abrasion is moderate and where powder distribution permits, the use of smaller bits presents a favorable opportunity for reducing drilling costs.

Jan 1, 1939

By P. Kovalchik, S. Peng, G. Cole, R. Matetic

Previous studies conducted by the National Institute for Occupational Safety and Health (NIOSH) have shown that approximately 90 percent of coal miners and 49 percent of metal/nonmetal miners had a hearing impairment by age 50 (Franks, 1997). Mine workers are exposed to additional noise levels underground due to the reflection of machine-generated noise that would otherwise dissipate in an aboveground setting. Determining the sound-absorption coefficients is essential in modeling the sound field in an underground mining environment. Sound absorption is a measure of the amount of acoustic energy that strikes a surface and is absorbed rather than being reflected. The acoustic environment that mining machines operate in is a critical factor affecting the sound pressure level exposure for mining machine operators. Impedance tube testing is an inappropriate method for determining underground absorption coefficients due to the brittle composition of materials such as coal and slate. Classic absorption coefficient estimation using T60 measurements will not work well in an underground environment because the theory assumes a finite room, a diffuse field and a relatively uniform absorption. None of these assumptions are true in an open-ended mine entry. This paper presents a method using a ray-tracing technique to determine absorption coefficients for underground mines. Absorption coefficients are determined and presented for octave bands from 63 Hz to 8 kHz. The absorption coefficients are essential for determining and predicting potential overexposure to machine operators in different mine environments.1

By H. H. Lowry

Abstracts of papers on low-temperature tar published since 1942 are included in this selected bibliography. Abstract journals -including Chemical Abstracts, Fuel Abstracts, and B.C.U.R.A. Monthly -have been consulted through 1957. Abstracts of papers on low-temperature carbonization are not included unless information is given on composition, economics, processing, or utilization of the tar; thus, papers reporting only yields of tar from assay tests, pilot plants, and commercial operations are excluded. Listing in the bibliography is alphabetical by authors; arrangement under each individual is chronological. Secondary authors are cross-indexed to the main entry. The subject index uses the main entries: Composition, economics, processing, and utilization with appropriate subheadings. A code letter following the abstract number indicates the country where the paper originated.

Jan 1, 1959

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 Julian Vigil

Mountain bumps have been a problem in the coal-mining industry since the 1920&apos;s. The problem became severe in the late 40&apos;s and early 50&apos;s with the introduction of mechanized mining. Fatalities and serious injuries resulted from mountain bumps during this period. Mountain bumps again caused fatalities in the 80 &apos;s. Interviews were conducted mostly with retired coal miners having experience in bump conditions with five different coal companies. These men held positions as president, general manager, roof-control specialist, mining engineer, inspector, superintendent, mine foreman, shift foreman, section foreman, miner operator, electrician, and general laborer. Together, their comments give a good idea of the things tried to mitigate mountain bumps in the Appalachian Coal Fields.

Jan 1, 1991

Identify abnormal low-frequency blast vibrations. Develop blast design procedures for such adverse vibration cases, including the selection of delay times based on ground conditions, vibration frequency, and structural response. The Problem Research by the Bureau of Mines and others has demonstrated how structural response and the potential for cracking and other damage depends on blast vibration frequency in addition to particle velocity amplitude. The Bureau has published recommended criteria for blasting near low-rise residential-type structures; criteria that specified frequency-dependent amplitudes. The typical vibration frequency from surface mine and quarry blasting is 20 hertz (Hz). Sites have been found, however, that regularly produce ground-roll-type waves with frequencies as low as 4 Hz. The most serious cases arc those where vibration frequencies fall in the natural frequency range for structures of 4 to 12 Hz. Such vibrations have a higher potential for causing, damage to residential structures, particularly at the low end of this frequency range, because of the resulting high strains these frequencies produce in structural materials.

Jan 1, 1990

By F. B. Tough

An inusual job in controlling a gas well was executed on No. 1 well, Inland Oil and Refining Co., situated on the Alkali Butte structure, Freemont County, Wyoming. The work was done by the company in cooperation with the . S. Bureau of Minos, and resulted in a fuel saving of $5,000 in one season and of a repair job estimated at $15,000, or a total of $21,000. The woll had an open flow of 6,000,000 cubic feet of gas per day with a rock pressure of 560 pounds per square inch. A string of 6-1/4 inch casing with a toothed shoe was set on the cap rock above the gas sand at 2181 feet. Then the well was shut in, the gas worked out around both the 6-1/4 inch and 6-1/4 inch strings, breaking through the ground surface at several points under the derrick. Casing troubles encountered while drilling made it unsafe to attempt Coving the 6-1/4 inch string. Moreover, it was important from the standpoint of good management to make the gas from this well immediately available for fuel in other operations and, in addition, to save the gas from wasting. An ordinary mudding and cementing job would have necessitated sitting down this well for probably two months out of a very short operating season. For these rousons it was determined to mud the well, if possible, between the 6 and 8 inch strings by pumping mud fluid through the connection from the gas clamps. The gas clamps used wore of ordinary construction with a side outlet, making them equivalent to a Braden-hoad. When the mud fluid was pumped through the gas clamp into the space between the two strings, it began to appear inside the 6-1/4 inch string. It apparently entered the well either around the toothed shoe or through a "line sawed" leak in the string and was blown out of the well as fast as it was pumped in. In order to stop the leaks, about a half a gunny sack of finely chopped bull ropa was fed with the mud between the casings as before. The hemp gathered around the valves of the mud pump to some extent, and it also stuck in the casing-head fittings a few times. An obstruction in the casing- head would be removed by permitting the woll to "blow back" for a few seconds, and one in the pump valves, by removing the valves when necessary. In two days' time sufficient hemp and mud had been pumped into these strings to kill external and internal pressure. The wall was shut in and allowed to stand over night. The next day, by opening and closing the master gate at intervals,

Mar 1, 1921

By J. Nielen van der Merwe

For the last 3 decades, coal pillars in the Republic of South Africa have been designed using the well-known strength formula of Salamon and Munro that was empirically derived after the Coalbrook disaster. The database was recently updated with the addition of failures that occurred after the initial analysis and the omission of failures that occurred in a known anomalous area. An alternative method of analysis was used to refine the constants in the formula. The outcome was a new formula that shows that the larger width-to-height ratio coal pillars are significantly stronger than previously believed, even though the material itself is represented by a reduced constant in the new formula. The formula predicts lower strength for the smaller pillars, explaining the failure of small pillars that were previously believed to have had high safety factors. Application of the new formula will result in improved coal reserve utilization for deeper workings and enhanced stability of shallow workings.

Jan 5, 1999

By Allen D. Look, Alan A. Sharp

"Churn drills for blast-hole drilling are important to the operation of most open-pit metal mines, but certain hazards are created through their use that cause fatal and nonfatal accidents.2/ The purpose of this circular is to present a code of safe practice prepared and followed at a large Arizona open-pit copper mine, where safety is recognized as a definite operating problem. Such a code, based upon a thorough study of available published information as well as practical experience gained from accidents and near accidents, is the result of considerable effort and monetary expenditure. It is hoped that other operations using churn drills may benefit by the code presented, either in comparison with existing rules or in the preparation of new operating instructions.The code presented is in use at the Morenci Branch of the Phelps Dodge Corn., Morenci, Ariz. Similar codes are used at the New Cornelia Branch, Ajo, Ariz., and in the waste-pit operation at the United Verde Branch, Jerome, Ariz.The authors gratefully acknowledge the cooperation of the Phelps Dodge Corp., whose personnel made: possible the collection of information included in this circular. Special acknowledgment is due Paul Simmons, safety engineer, and 3. U. Larsen, general drilling and blasting foreman, Morenci Branch, Morenci, Ariz.The importance of churn drills in Arizona mining is apparent when it is realized that at all the open-pit copper mines blast holes are drilled with this equipment. At Morenci, approximately 98 percent of the ground is broken by primary blasting of churn-drill holes.There has been a growing tendency in Arizona's copper-mining industry to change from underground to open-pit mining, with consequent increase in the use of churn-drill blast holes for primary blasting.3/Thirty-three churn drills are in use at Morenci for blast-hole drilling. The drills operate with 440-volt alternating current, and trailing drill cables, in 500-foot sections, connect with the portable drill transformer station.4/ Three Diesel-driven churn drills for drilling prospect holes also are in use. About 25 churn-drill shifts are operated currently on a three-shift basis. Ordinarily, a two-man crew is assigned to each drill; however, with prospect-drill crews a third man is added to handle the sampling.Benches are carried at a uniform height of 50 feet. Churn-drill holes are 9 or 12 inches in diameter and about 60 feet deep. The holes are 18 feet apart in hard waste and 21 feet in soft waste and ore, with toes from 11 to 13 yards at the shovel grade. Both the hole spacing and the toe have been increased gradually, until they are now 2 yards longer than in 1940. The average ground broken per hole is 3,500 tons, or 60 tons of ore per foot of hole drilled. The holes are drilled 10 feet below the bench to avoid leaving hard bottom, which requires secondary drilling and blasting.Starting with a 12-inch casing, the prospect holes average 1,200 to 1,500 feet in depth and terminate with 6-inch casing; the deepest hole drilled to date went to 1,735 feet.In addition, churn-drills are used at Morenci to drill blast holes in the silica pit, limestone quarry, and the tailings dam. Other irregular applications are the drilling of holes to set trolley roles and for concrete piling in building foundations."

Jan 1, 1950

By McHenry Mosier

Although local conditions at each mining property affect the percentage of increase in drilling speed for a given reduction in gage of bit, the results of drilling tests conducted at the Mount Weather testing adit show such a large advantage in favor of the smaller gages that they should be indicative of the possibilities at most mines. GENERAL DISCUSSION Although it is common knowledge among miners that drilling speed increases with reduction in gage of bit, the large amount o this increase is not so universally realized, A recent Bureau of Mines4/ survey of drilling practices in 28 metal mines showed that the maximum gage of starters is 3 inches and that at 50 percent of the mines starters are 2-1/4 inches or larger, Because this size appears too large for conditions at most mines where abrasion is moderate and where powder distribution permits, the use of smaller bits presents a favorable opportunity for reducing drilling costs.

Jan 1, 1939

By C. Melvin Lepper

To provide the U.S. mining and energy industries with information in the selection of seismic detectors for high-resolution applications, the Bureau of Mines evaluated 35 velocity detectors (geophones) and 9 accelerometers. It was found that velocity detectors with natural frequencies from 10 to 100 Hz were effective for seismic studies in the 5-to 600-Hz spectrum and that accelerometers were most effective in the 100-to 1,000-Hz range. Although the accelerometers described in this report performed about equally with the better velocity detectors tested, the accelerometers showed higher output voltage levels at higher frequencies, which is an advantage. It is important that a seismic detector and its intended purpose be matched properly to achieve the intended results in high-resolution applications. There are many factors that could affect proper matching, thus careful selection of the best detector for a particular application must be exercised.

Jan 1, 1981

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 W. A. Heitbrink, J. D. Ashley, R. E. Gresh, J. A. Organiscak, A. B. Cecala

Many different types of mobile equipment used in surface coal mining utilize enclosed cabs to protect equipment operators. The overburden removal process is extremely dusty and can cause excessive exposure to respirable dust, especially crystalline silica. After equipment is used for years, many components of the enclosure deteriorate and its effectiveness is greatly reduced. This report discusses a cooperative research study performed on an Ingersoll Rand DM45E surface drill, retrofitted with a new Sigma pressurization and filtration system. Respirable dust concentrations were substantially reduced from 0.64 mg/m3 during pre-testing to 0.05 mg/m3 during post-testing with the new system, representing a 92% reduction in dust levels in the drill cab. This new system appears to be a very well-built and sturdy device, well-suited for the mining industry.

By J. B. Gayle

Numerous investigators have studied the addition of inert4 substances to coking blends. Generally, the results have yielded trends applicable only to the particular coals and inerts studied. Recently, somewhat more general results were obtained for cokes produced from blends containing added inerts, using the sole-heated oven. The present investigation represents an extension of this work to include a wider range of inerts and to determine the applicability of the various conclusions to cokes produced with the Tuscaloosa oven. The results indicated that an increase in dry-coal charge density of 1 pound per cubic foot (p.c.f.) caused a corresponding change in the hardness (1/4-inch tumbler index) of Tuscaloosa oven cokes of 0.9 percentage points. This value compares favorably with the value of 0.7 determined previously for the sole-heated oven and Bureau of Mines-American Gas Association (BM-AGA) retort.

Jan 1, 1961