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By John H. Soule

"Although known since 1902, the Capitan iron deposits have not been mined, mainly because of the low grade of the ore. Only a small amount of exploratory work has been done since the discovery. C. H. Johnson, district engineer of the Bureau of Mines in New Mexico in 1942-45, visited the property; as a result, preliminary sampling of the property was done in 1943.3/ Owing to the results of this preliminary work, an extensive development program by wagon drilling was planned and executed early in 1944.4/ Equipment available at that time was limited in the depths that could be reached; and, as geologic studies indicated the possibility that other limestone beds below the surface might have been replaced by iron minerals, a supplementary development program by churn drilling to reach greater depths was prosecuted in late 1947 and early 1948, with the author as project engineer.The following sections note the location and accessibility of the deposits and mention the physical features, climate, labor and living conditions. Property and ownership are outlined, the history and production of the property stated briefly, and the ore deposits are described. The work done by the Bureau of Mines is explained and supported with maps, sections, and assay logs.From March 1, 1944, to June 1, 1944, 179 wagon-drill holes totaling 3,490 feet and 4 test pits totaling 75 feet were completed, and from November 1947 to February 1948 seven churn-drill holes totaling 2,488 feet were drilled and sampled. Metallurgical tests are summarized and referenced.With respect to this report, special acknowledgment is due to J. H. Hedges, chief, Tuscon, Ariz., Branch, Mining Division, and to W. R. Storms, engineer in-charge of the Silver City, N. Mex., field office. Acknowledgment also is due to J. Bruce Clemmer, chief, Tuscon Branch, Metallurgical Division, for analyses done in the last phase of this project, and to Dr. Vincent Kelley of the Federal Geological Survey for his cooperation during this work.The Capitan iron deposits are in secs. 10, 11, 13, 14, and 15, T. 8 S., R. 14 E., central Lincoln County, N. Mex., as shown on figures 1 and 2. The only known large deposit upon which the Bureau did development work lies around the common corner of secs. 10, 11, 14, and 15, as shown in figures 2 and 3, and just southwest of and in the foothills of the Capitan Mountains. This range of mountains trends east and west. The elevation of the highest peak is 10,205 feet.The deposits are reached by traveling on State Highway 48 from Capitan north for 4 miles, to a point just inside the Lincoln National Forest fence and thence due east over a poor dirt road for 1.2 miles. Just over a small wooden bridge, the left-fork is taken, and the poor road is followed to the deposits, a further distance of 1.1 miles (see fig. 1). This road follows the fence to and through the Davis ranch. All the roads from Capitan to the deposits are dirt roads but are impassable only for short periods during and after heavy rain or snow."

Aug 1, 1949

By D. D. Bolstad

This report describes the construction and use of a direct-reading azimuth protractor (DRAP) by the Bureau of Mines for measuring fracture orientations in mines, tunnels, or other excavations having a magnetic environment. The DRAP allows a rapid and simultaneous measurement of dip and azimuth of dip of a fracture plane. The instrument is rugged, lightweight, and easily constructible from standard hardware items. The DRAP can measure dip to within ±0.5° and azimuth of dip to within ±5° of corresponding values obtained with a Brunton compass.

Jan 1, 1974

The Bureau of Mines has conducted considerable research to improve life support technology for underground mining applications. This proceedings volume presents several new developments that may help increase the chances of mine workers in surviving underground disasters. Several papers address the performance of present self-contained self-rescuers (SCSR's) and provide proposed guidelines for the design and testing of a second-generation SCSR. Improvements In the safety and effectiveness of mine rescue and recovery operations are described, including the design of a low-profile rescue breathing apparatus and a rescue team helmet.

Jan 1, 1987

By Robert E. Johnson

THE MILDEST recession of the postwar period ended in the first months of 1961, and the economy climbed steadily during the remainder of the year. The Federal Reserve Board (FRB) Index of Industrial Production climbed from a low point of 102 (1957=100) in February to 115 in December, a rise of 12.5 percent. But economic activity for the year as a whole was only 1.5 percent higher than the 1960 level, as measured by real gross national product (GNP). Although activity increased throughout 1961, much of it was merely regaining ground lost in the decline during the second half of 1960 and the first months of 1961. The FRB index did not re ain the July 1960 high of 110 until June of 1961. Mineral fuels did not fully share in the economic recovery. The (FRB) crude oil and natural gas index rose only 5 percent between February and December. Although the index of monthly coal production increased 11.5 percent during the year, the February level was so low that total production for the year was the lowest since 1954. Total production of energy measured in Btu's was up slightly (1 percent) during the ear. Coal production was down; production of crude oil and natural gas was up. Calculated consumption paralleled production-total up slightly, coal lower, and crude oil and natural gas higher. The most interesting product developments were in coke and residual fuel oil. Sales of coke showed a sharper drop than steel

Jan 1, 1962

By C. Lorimer Colburn, Spencer P. Howell

"Numerous accidents are still being reported that have resulted from the delayed explosions in drill holes. Some of these accidents have boon due to retarded burning of the fuse, causing the charge to explode after the miners had returned to the face, but the majority have been caused by the minors drilling or picking into unexploded explosives, either in unexploded bolas or in broken rock. Although the industry has been cautioned by the Bureau of Mines, manufacturers of explosives, and other agencies regarding the care that is necessary in the handling of explosives, yet reports of these accidents are still being received and many of the accidents are fatal, They occur in spite of the best efforts of the mine officials and men, but many of those that have been reported could have been avoided if the following precautions had been observed:Misfires should not be approached even for the purpose of inspection until three hours have elapsed if fuse was used, and ten minutes if electric detonators or electric igniters were used. These are minimum recommendations. It would be preferable to allow a longer time, for fuses have been known to hang fire for twelve hours."

Aug 1, 1920

By D. J. Parker

Owing to its flexibility, convenience, economy, ands availability, electric power is peculiarly adapted to the needs of the coal -mining industry. Because of such obvious advantages, the application of electricity has in recent years become more and more extensive and is now a most important factor in the Nation's coal prediction. With the introduction of this type of power into coal mines certain new, specific, and well-recognized hazards resulted. The question arises, therefore: Has the industry met these hazards squarely and. effectively? With respect to some individual companies, the answer happily is in the affirmative; with respect to the industry as a whole, there is much room for improvement.

Jan 1, 1933

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

APPENDIX G TABLE IX MINE A 1 General Total Employees 3000 Maximum No, of Men Underground/Shift 800 Shifts/Day 3 Type of Ore Copper Daily Production (TPD) 25,000 Mining Method Room and Pillar Vein Dip/Thickness l0°/l5 ft Number of Levels/Max Depth 1/0 to 1800 ft 2 Haulage, Conveyances and Travelway Main Haulage Rubber Tire/10 miles belts Man-Ways - Typical Route of Travel to most Remote Sites Rubber Tired Vehicles Hoists - Capacity/Speed/Depth Depth under construction 3 Ventilation 2,500,000 cfm - 5 surface fans 4 Compressed Air N/A cfm 10" and 12" lines 5 Incendiary Storage 2 day powder 1000 gal fuel oil 6 Communications 1 complete system - all shops, dinner areas, and feeder/breaker stations 2 partial systems 7 Escapeway- Systems Man Check System Section boss log and punch cards Alarm System Mine phones with paging system Posting of Escape Route Maps No Escape Signs Every intersection, fluorescent paint Evacuation Drills 2/yr/shift Refuge Chambers - 1 cinder block - No longer required

Jan 1, 1976

By B. J. Scheiner

The Bureau of Mines investigated an electrolytic technique for oxidizing ore slurried with brine to determine its effectiveness in extracting mercury from ores. Optimization experiments conducted on a 2.5-pound-mercury-per-ton ore resulted in 95 percent mercury extraction with operating conditions of 30° C, a current density of 0.5 ampere per square inch, and a treatment time of 3.5 hours. Corresponding power consumption amounted to 38 kilowatt-hours per ton of ore treated. Amenability experiments conducted on 62 different mercury ores resulted in 90 to 99 percent mercury dissolution from ores containing 0.6 to 20 pounds of mercury per ton of ore, Treatment time ranged from 1 to 7 hours in 4-to 20-percent sodium chloride solutions, and power consumption was in the range of 10 to 50 kilowatt-hours per ton of ore. Pilot mill experiments in a 100-to 200-pound-per-hour extraction plant have been in progress to quantify power and reagent requirements. These data will be reported in a subsequent publication.

Jan 1, 1973

By Spencer P. Howell, J. F. Xrawshaw

"A recent premature explosion in a quarry that resulted in the death of nine men and the injury of several others, was directly traceable to a spark from a steam-shovel working under the face directly in front of the hole being loaded. This accident and similar accidents on record in the Bureau of Mines indicates the necessity for pointing out the dangers in quarries of explosives being ignited by sparks emitted from steam shovels and locomotives, and making recommendations for greater safety.Many premature explosions of which the following are examples, have resulted from sparks, both from steam shovels and passing locomotives. (A) The premature explosion occurred during August, 1920 at a limestone quarry, in a 100-foot drill hole, loaded with 1000 pounds of high-grade nitroglycerin and gelatin dynamites. This hole had been loaded with 500 pounds of 75 per cent strength gelatin dynamite, 300 pounds of 60 per cent straight nitroglycerin dynamite, and 200 pounds of 50 per cent straight nitro¬glycerin dynamite; but no steaming had been placed in the hole. The powderman loading the hole saw the spark enter the hole and was able to run a short distance before the explosion took place, thus escaping serious injury. The explosion of this first hole was transmitted to nearby hole similarly charged. The fall of rocks from these two holes buried the steam shovel and some of the gang of men working around it.(B) In September, 1909, in the Canal Zone, a steam shovel uncovered a misfired charge of dynamite. While the dynamite was being picked up the steam shovel continued working nearby. A spark from the shovel set fire to some loose dynamite spilled on the ground from the troken cartridges. This fire was transmitted to a hole which was being loaded close by and caused a premature explosion. One man was injured."

Nov 1, 1920

By F. E. Block

This investigation was made as part of a larger program to develop basically new methods for preparing yttrium metal in high-purity form. It was performed under a cooperative agreement between the U.S. Atomic Energy Commission (AEC) and the Federal Bureau of Mines. Laboratory-scale tests were made to devise methods for preparing high-purity, anhydrous yttrium trichloride, for reducing the yttrium trichloride with sodium or lithium, and for consolidating the resulting yttrium sponge into ingots.

Jan 1, 1960

By Raymond B. Ladoo

Talc mining ·in the state of New York was first started about 1876, but important production did not start until 1880, when about 4,000 tons were shipped, In 1883 the production was about 6,000 tons valued at $75,000, or an average of $12.50 per ton, From this date the production gradually inc~eased to a peak of 93,236 tons in 1916, valued at $961,510 or about $10,30 per ton, In 1918 the production declined to 71,167 tons valued at $902,100, or about $12,70 per ton. For ma.ey years New York ,-Jas the largest producer in the United states, but in 1917 and 1918 its output was surpassed by Vennont in tonnage thou[;h not in value. In June, 1920, there were three companies producing talc in the Gouverneur district, st. La,1Tence Co'Ul'lty, and one near Natural Bridge.

Oct 1, 1920

By Robert J. Anderson

"Introductory Statement:In melting the nonferrous metals and alloys for foundry and rolling-mill work, even with the best practice, there is always a loss, due principally to oxidation. Melting losses are a serious source of financial loss in brass and bronze practice, but they are especially serious in the case of aluminum and its light alloys because of the relatively high price of aluminum. The indications are that the melting losses in many aluminum foundries of the United States are at times high, and in some plants (especially where certain furnaces are used) they are entirely too high when compared with the average practice. The fact that the oxidation losses in melting aluminum and its light alloys are high is well known, but it is doubtful whether the factors governing the most suitable methods of melting for a minimum of dross are equally well known. The metal losses involved in melting are worthy of especial attention now because of the rapidly growing importance of the aluminum industry.Factors in the operation of furnaces which affect metal and fuel losses:The essential requirements in the commercial melting of any metal or alloy are that the melting losses should be low, the melting should be rapid, and the melting cost should be low. Certainly it may not be possible to obtain these requirements in any furnace selected at random, and it follows that the design and the operation of furnaces for melting different metals are subjects that warrant the most extensive study and tests. As a general rule, in the industrial operation of furnaces, run on aluminum and aluminum alloys, an effort is made to keep the dross and oxidation losses as low as possible. In attempting to hold dross losses low, numerous factors involved in metal melting may be overlooked, or at least disregarded, and the net practical result may be that the furnace is exceedingly wasteful of fuel. In the case of aluminum and its light alloys, the effect of increasing the temperature of melting, at least in an atmosphere containing free oxygen and nitrogen, is to increase the net melting loss."

Apr 1, 1921

By R. Barany

The heats of formation at 298.15° K of calcium tungstate, calcium molybdate, and magnesium molybdate were determined by solution calorimetry. For the heats of formation from the elements, the values obtained were calcium tungstate, -402.4 ± 0.5 kcal/mole, calcium molybdate, -369.5 ± 0.3 kcal/mole, -334.8 ± 0.2 kcal/mole. For the heats of formation from the oxides J the values obtained were calcium tungstate, -49.2 ± 0.4 kcal/mole, calcium molybdate, -39.6 ± 0.1 kcal/mole, and magnesium molybdate, -12.9 [ ] 0.1 kcal/mole.

Jan 1, 1962

By T. O. Llewellyn

Laboratory batch and semicontinuous studies showed the feasibility of using methylene bromide and trichloroethylene in heavy liquid cyclone concentration of a Nevada brucite ore. Cyclone tests of minus 20- plus 200-mesh material at specific gravities of 2.42 and 2.44 recovered 71.6 to 74.9 percent of the MgO at grades of 95.5 and 95.4, respectively. Neither increasing the feed pulp density from 5 to 25 percent solids, nor varying feed pressures from 25 to 60 psig had any apparent effect on the heavy liquid separation.

Jan 1, 1971

A working understanding of present coal mine communications equipment and its usage is an essential prerequisite to the determination of future communications systems needs. As a means of gaining this understanding, a field survey program was planned and implemented. The plan formulated called for obtaining both qualitative and quantitative data from three main areas of interest: the types, physical condition, and performance characteristics of equipment presently used in the sample of mines to be visited; the relationship present communications has to safety and productivity; and the practical, future communications needs as de- rived through interviews with present mine management. In compliance with the original plan, the type and physical condition of equipment in use was found through at-mine interviews of personnel responsible for installation and maintenance. Further evaluation was made through the review of equipment instruction books and maintenance manuals and through in-mine inspection of communications equipment and signal distribution networks. The performance characteristics were determined through listening to conversations and noise levels and by talk-testing from randomly selected stations throughout each mine. System operation was evaluated by asking communication system users their opinions and by noting any difficulties personnel had using the systems. Plans in the second area of interest--that of relating communications to mining operations-- were to observe mining operations and usage of communications; to gain insight through interviews with mine management, from mine superintendents to section foreman; and to look for correlation between communications, production, and safety. Lastly the communication systems needs and requirements were determined from first hand interviews of mine personnel, from the analysis that lead to the individual reports on the eight mines, and from an overview of all of the mines studied. The latter was done to find commonality between the needs of the individual mines such that system requirements having the greatest universal value could be derived, thus lowering per-mine cost of equipment through economics of scale. Armed with tape recorders, backgrounds in communications technology, and with an interview questionnaire that was improved as experience was gained, engineers visited the eight mines to obtain the raw data from which the needs and requirements could be developed through analysis. During the first part of the program two days were spent at each mine so that interviews and tours could be completed, and so that full eight hour shift phone usage recordings could be obtained for statistical analysis. One day visits were sufficient for those trips where tape recordings were not made. During the later trips less attention was given to the statistical nature of communications, and more to how communications relates to the total mining operation.

Nov 15, 1974

By A. Wahab Khair, N. P. Reddy

"Abstract. This paper attempts to characterize coal breakage due to the action of a continuous miner. Coal blocks with an approximate dimension of 45.7 cm x 38.1 cm x 15.2 cm (18 in. x 15 in. x 6 in.) were first subjected to confining pressures equivalent to in-situ conditions. Then the blocks were cut along the face/butt cleat direction by the cutting head of a specially designed unique automated rotary coal cutting simulator (ARCCS), simulating the action of the continuous miner in underground coal mining. During the tests under a particular set of in-sit and operating parameters a number of other para¬meters such as 1) penetration of bit into coal, 2) penetration resistance (thrust and cutting pressures), 3) rotating velocity of cutting head, and 4) acoustic emission activity in the coal block were monitored. After each cutting cycle the fractured surfaces were photographed and a velocity survey was conducted by using a sonic technique. At the end of each experiment, the coal fragments were collected in bags using a vacuum cleaner. These fragments were then sieved for size distribution analysis.INTRODUCTIONThe Federal Coal Mine Health and Safety Act of 1969 was enacted to emphasize to coal operators that the airborne respirable dust should not exceed 2 mg/m3, thus reducing the incidence of coal workers pneumoconiosis.Dust control techniques such as conventional water sprays and dust collectors are only partially effective and require additional equipment expenditures. A more authentic approach would be to reduce respirable dust at the source the continuous mining machine cutting head, where the fragmentation process occurs. Improving the fragmentation process by characterizing the coal breakage will not only reduce respirable dust at the face, but it will also decrease the amount of respirable dust that is liberated during the secondary handling.Fragmentation of coal can be expressed as a function of three major groups of parameters; namely, 1) coal properties, 2) in-situ conditions and 3) operating parameters. This paper deals with the characterization of coal breakage along both face/butt cleat direction as a function of machine operating parameters such as rate of advance, angle of attack, bit configuration, bit lacing (spacing and pattern) and bit speed at various equivalent in-situ pressures."

Mar 1, 1989

By R. Hogg, T. F. Dumm, L. G. Austin, M. Yekeler

"Dedicated to Professor Kurt Leschonski on the occasion of his 60th birthdayAbstractThe kinetics of batch grinding quartz from a feed of 600 by 425 Jim to a product of 80% less than 10 µm have been determined using screening and laser diffractometer sizing for size analysis. The specific rates of breakage decreased by a factor of about three when the material became less than about 100 µm in size, but the primary breakage distribution function also changed to give proponionately more fine material, so that the grinding efficiency expressed as the development of surface area (B.E:T.) per unit of energy input did not decrease. Analysis of the shape of the particles in the 2S x 38 µm size range showed that particles of this size produced by roll crushing or by 8 minutes of grinding of a 42S x 600 µm feed were not different but at long grinding times the particles were rounded. This suggests that the breakage mechanism changes to give more chipping and abrasion and less disintegrative fracture. As the material approached the ultrafine size range it adhered to the mill case and there was no further size reduction. However, a technique for striking the mill case to dislodge the particles was successful in allowing further grinding to 400Jo by weight less than 2 µm.1 IntroductionIt has long been known that it is difficult to accomplish ultrafme grinding in a tumbling ball mill. It has been reported [12, 13] that there is no further size reduction after some hours of grinding cement clinker in laboratory ball mills. Austin and Bagga (1] and Shah and Austin [16] have given an analysis of the kinetics of dry grinding in a laboratory ball mill, introducing the concept of a slowing-down factor K, which is the ratio of the specific rates of breakage at some long time of grinding to the normal specific rates of breakage. They showed that this factor decreased as fine material accumulated in the mill charge. Cottaar and co-workers [8-10] showed that the movement of solid in a dry laboratory ball mill was influenced by the fineness of the solid and by the presence of air, and that evacuation of air from the mill caused particle sedimentation and low rates of breakage.The concept of the slowing-down factor has also been used in wet grinding [17, 18]. At very .high slurry concentrations it appears to be correlated with a steady decrease of mill power caused by the build up of a layer of deposit sticking to the mill case. This reduces the effective mill diameter and, eventually, the balls start to adhere in the layer and centrifuge. A decrease of mill power always indicates less lifting and tumbling of the media and, hence, less breakage. However, slowing-down of breakage rates was also observed at lower slurry concentrations where there was no change in mill power."

Jan 1, 1990

By Barry Beith

This report details a study which was designed to assess the effect of retroreflective material on detection and form recognition of workers in the mining environment. Further, the study attempted to determine an optimal configuration of retroreflective material with respect to amount and placement on the body to facilitate detection and form recognition. The study involved a one-fifth scale simulation of the perceptual task found in the underground low seam coal mine. Four different configurations of reflective material were presented in various body positions, common to low coal miners, at three different locations in the visual field, i.e., 10, 25, and 45 degrees from focal line-of-sight. All reflective configurations significantly improve detection and form recognition over configuration (cap only) worn by miners. As the amount of reflective material increased, accuracy also tended to increase. Based on the configurations investigated by this study, the Armband configuration (i.e., refectorized armbands, belt, self-rescuer and battery pack) was determined to be optimum. It is recommended that an actual field evaluation be made using the Armband configuration to establish if simulator results are generalizable to the underground mining environment.

Jan 1, 1980

By William A. Beck

The K.T. Dome Mine in Owen and Henry Counties, Ky., has been operated intermittently under several different owners since the early nineteen hundreds. No record of total production is available, but it is said that during 1943 and 1944, while being operated by the Twin Valley Mining Co., 30 cars of zinc concentrates and 10 cars of lead concentrates were shipped. The vein can be traced for about 10 miles along the strike from Lockport, Henry County, to Big Twin Creek in Owen County. It pinches and swells both vertically and horizontally and is reported to average 20 inches in width. In order to develop additional reserves on the property, the Bureau of Mines drilled four holes on the south extension of the vein. Drilling disclosed that the vein dips to the west on the south side of the Kentucky River and is continuous along the strike and in depth. However, the extent of mineralization in the vein varies widely.

Jan 1, 1949

By D. J. Bauer, V. E. Shaw

Bastnasite concentrate was processed to recover total rare - earth content and obtain high - purity ceric nitrate . Reaction of the concentrate with 45 percent caustic soda for 1 hour at 180 ° C in stainless steel vessels resulted in 99 - percent conversion to rare - earth hydroxides . A drying step followed by solution in nitric acid produced a tetravalent cerium fraction amenable to extraction with TBP . Single- stage separatory funnel experiments resulted in 90 -percent recovery of ceric nitrate at a purity of 99.9 percent . Fractional extraction in pressure - bubbler columns resulted in 98.5 - percent recovery of 99.5 -percent - pure ceric nitrate .

Jan 1, 1964