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The Navajo Nation is the largest Indian tribe in the United States and occupies a reservation encompassing about 16 million acres in northeastern Arizona, southeastern Utah, and northwestern New Mexico, with three satellite reservations in western and west-central New Mexico. Coal mining is conducted on a commercial scale at four large surface mines: two are in northeastern Arizona and the other two are in western and northwestern New Mexico. The principal coal-bearing region in Arizona is the Black Mesa coal field, which underlies about 3,200 square miles in the northeastern part of the state. The Black Mesa coal field lies within the boundaries of the Navajo and Hopi Indian Reservations and makes a significant contribution to the economies of both tribes. Coal reserves in this field are estimated to be 352 million tons, less than one-tenth of one percent of U.S. coal reserves. This coal was first mined by prehistoric Indians as fuel for cooking and heating. It has been estimated that the Indians removed more than 100,000 tons between 1300 and 1600 A.D. The coal-bearing region of the Navajo Reservation in New Mexico is located in the San Juan Basin. The Navajo Mine, which supplies coal for the Four Comers Power Plant, is located in this area. The Office of Surface Mining's Albuquerque, New Mexico, Field Office is responsible for regulating mining and reclamation on Navajo land. The Navajo Abandoned Mine Land Program was approved May 16,1988.

Jan 1, 1992

It was decided to demonstrate the use of latex as a soil sealant by irrigating three small plots at Lanse coded 03A on the west of Brown's run, 3C on the east of the stream, and SB, an area on the spoil bank. Earlier plans to sprinkle latex on twenty acres of the side were abandoned because of easement difficulties. On the basis of results from laboratory soil columns and small plot experiments at Lanse, Naugatex J-3471 was the latex of choice. The irrigation system consisted of plastic pipe (two inch size) for the main line from the water meter to the irrigation area because of lower price and ease of handling. Because the municipal water supply was limited to 20 gpm, low-volume 14V-LA-TNT Rainbow sprinklers were specified with low angle 20° 3/32 inch nozzles to reduce wind effects. At the expected nozzle pressure of 50 psig, each sprinkler covers a 56 foot diameter area at a delivery rate of 1.64 gpm. Each of the three test areas was covered with twelve sprinklers, six used with water alone as a control area and six used with latex and subsequent water spraying. One-quarterturn inlet valves were positioned on the main lines leading to each area. Flow to the treated and untreated areas was set at 10 gpm by Dole flow control valves in the' main entry lateral to each of the six sprinklers. To avoid the use of dilution tanks, a Moyno pump operated by a 1/2 HP, 1140 rpm 5:1 reducing drive was used to deliver 1/2 gpm 48% latex for mixing with 10 gpm water. A 3500 watt gasoline generator provided power for pump and light. Latex was applied at night to take advantage of the lower evaporation and higher humidity. Neutron gauge access tubes were placed in orthogonal arrangement in the areas that were irrigated, with four tubes in the section that were sprinkled with latex and four tubes in the untreated irrigation section in each of the three locations. Each tube, five and one-half feet long and sealed at one end with a plug welded in place, was inserted five feet into the ground in a cored hole. In each irrigation area, the left section of six sprinklers spread the dilute latex while the right section sprinkled the same amount of water to serve as a control plot. As shown in Figure 7 each sprinkler has a 56-foot sprinkling diameter at 50 psig design water pressure. The sprinklers are 30 feet apart on the laterals and 35 feet apart between the laterals. The access tubes (coded AT on Figure 7) were placed in comparable spots on the two sections of each area, so that four soil moisture measurements on the two sections could be compared. Bouyoucos resistance blocks (BB) were also placed in areas 03A and 3C at one, two and three foot depths, but not on the spoil bank area.

Jan 1, 1972

By McHenry Mosier

The diamond-drill method was tried at the site selected for driving a 5- by 6 -foot air shaft at the east end of the Bellefonte mine . This method of sinking consisted , first , of drilling a 5-3/4 -inch-diameter cut hole from the surface , then drilling parallel 1.5 - inch ( EX ) holes around it at appropriate distances for a balanced blasting round . The cut hole was not charged with explosive but served as a relief chamber for the incipient expansion of the ground broken by. blasting the surrounding 1.5 - inch holes . The length of each hole was 236 feet , which was the finished depth of the shaft . The nature of the work in the air shaft was largely experimental , and many difficulties arising from the lack of special equipment and proper technique for precise alinement of the drill holes caused expensive delays before they could be overcome . Drilling the 5-3 /4 - inch hole , which was impeded by frequent core blocks in the core barrel , required 10 shifts . average advance per shift for an EX hole was about 150 feet . The The holes were charged and fired from the surface . The length of the section blasted in one cycle was 10 feet . Blasting progressed from the bottom of the shaft upward . The broken stone dropped into a chute , from which it was drawn into cars on the haulage level . The Bellefonte mine is an underground limestone mine.3/ The primary objective in undertaking the research on shaft sinking by diamond drilling was to perfect a method for driving connections between levels of the mine .

Oct 1, 1948

By F. S. Crawford

Mining companies are usually held responsible for accidents occurring to the public around their operations . Safeguards are provided by progressive mining companies to prevent the public from getting into danger from wandering unattended about the properties . At abandoned operations , where little or no supervision can be maintained , persons are often injured through the negligence of the operator in leaving the property in unsafe condition . After a mine is abandoned or a lease is given up by a mining company , the responsibility for the safety of the property rests sometimes with the State to which the property ownership reverts or with the owner of the land ; either is usually quite unwilling to spend any money to remedy unsafe conditions due to the carelessness of the mining company . The land owner or the State , or both , should see that the work of leaving safe conditions behind is done by the operator before he abandons the mine or gives up the lease . Some of the worst hazards to the public at abandoned mines are from old test pits , abandoned shafts or tunnels , open pits , caved ground , and bodies of water in abandoned open pits . Hazards exist from some of these sources during the active operation of a mining property as well . In the early stages of mining operations test shafts are often sunk from 50 to 100 feet deep through the more or less loose material overlying bed rock . When these shafts are left unprotected , they are a serious danger for many years to come . In the course of time they become largely concealed by brush and grass , and thus are a deadfall to anyone who might happen to step into them . This condition will hold true for any part of the country, but to a much less degree where it is not necessary to sink through a great amount of cover . Many progressive operating companies have adopted the practice of filling in test pits after securing the information desired . Some nonresident landowners and exploration companies , however , have left such pits open , and it is extremely difficult for the mine inspector to get into communication with the owner . Frequently operating companies voluntarily fill test pits on adjacent property not

Jul 1, 1930

By T. R. Jolley

An investigation was begun by the Bureau of Mines in 1948 to determine the preparation characteristics of American coals. Results of the investigation were planned to be reported on a county basis and, from the data developed, to determine: I. Which coals either as mined or after beneficiation by conventional preparation are chemically suitable under present standards for producing metallurgical coke. 2. Which coals require special and more intensive treatment in preparation than is now practiced. 3. Which coals cannot be considered potential sources of metallurgical coke, regardless of preparation by present-day methods. This report is one in the county series that describes the preparation characteristics of the coals of Bell County, Ky.

Jan 1, 1957

By John M. Devine

The oxygen bomb-is used extensively for determining sulphur in petroleum, although it has been suspected that the results obtained with this method are somewhat lower than the true varies. In this paper are compared the results given by the oxygen-bomb and Carius methods, the latter being generally accepted as the standard procedure for the determination of sulphur in organic compounds. The Carius determinations given hero were made in connection with a special study of this method.4 Only petroleum oils of the less-volatile type are covered in the present work. Summary and Conclusions. When used with oils containing 2.3 to 0.15 per cent of sulphur the oxygen-bomb method gives results that are consistenly if slightly lower than those obtained with the Carius method. This variation, for the oils used, amounts to 3.1 per cent (on tae basis of sulphur present) for the sample containing 2.47 per cent of sulphur, and gradually increases (as the percentage of sulphur in the oils decreases) to 7.3 per cent for the oil containing 0.15 per cent of sulphur. With the oils studied the presence of ammonium nitrate in the bomb does not influence the results materially.

Jan 1, 1928

By J. W. Walkiewicz

The Bureau of Mines investigated alloys containing mischmetal (MM), cobalt, copper, magnesium, and iron for use in permanent magnets in place of the scarce samarium-cobalt alloys. The magnetic properties of MM-Co, MM-Co-Cu, MM-Co-Cu-Mg, and MM-Co-Cu-Fe-Mg alloys were evaluated. Magnets were fabricated by powder metallurgy consisting of arc-melting the metals, crushing and grinding the resultant alloys, alining and compacting the powder, and sintering the compacts. Magnetic values of MHc = 4.7 kOe, BHc = 4.07 kOe, Br = 6.81 kG, and (BH)max = 10.0 MGOe were obtained with a MM-Co alloy containing 36±0.5 wt-pct MM. By substituting copper and magnesium for part of the cobalt, the values for MHc and BHc were increased to 29.0 kOe and 5.89 kOe, respectively. This intrinsic coercivity value exceeded that of Sm-Co alloy obtained in our laboratory. The addition of iron to MM-Co-Cu-Mg alloys resulted in an increase of Br to 6.90 kG and (BH)max to 10.3 MGOe. Copper and magnesium contents of these alloys were optimized to obtain maximum values of MHc.

Jan 1, 1981

By J. H. Jansky, P. W. Jeran, G. L. Hassett, M. C. Irani

This Bureau of Mines report tabulates methane emissions from U.S. bituminous coal mines with daily emission rates of at least 100,000 cfd according to States, counties, and coalbeds. Most of the methane emitted is from mines in the Appalachian States. Total daily emissionin 1975 was 216.3 MMcfd, compared with 227.0 MMcfd in 1971 and 214.5 MMcfd in 1973. Thirty counties in nine States had daily methane emissions of 1 MMcfd or more, These emitted 93 percent of the total methane. The highest emissions were from Monongalia County, W. Va,, which accounted for 18.9 percent of the total. This was followed by Marion County, W. Va., 9.6 percent; Buchanan County, Va., 8.5 percent; Jefferson County, Ala., 6.1 percent; and Washington County, Pa., 5.9 percent. The Pittsburgh coalbed emitted 45.7 percent of the total methane, followed by the Pocahontas No. 3 coalbed with 12.9 percent. The report lists 196 mines, of which 60 had daily methane emissions of 1 MMcfd or more. Included in this compilation are the identification and thickness of the coalbed, methane emission, coal production and ventilation rates, number of drifts, shafts, and slopes used for ventilation purposes, shaft depth, age of mine, and gas-to-coal ratio.

Jan 1, 1975

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

5.1 Incorporation with the Master Design Simulator Figure 1 diagrammatically illustrates the interaction between the various sub-systems that go to make up the total Master Design Simulator. As a total system, all sub-systems will be scheduled as needed by the controlling program. The data require¬ments for each will be made available either by providing common storage in core or by outputting data to external storage devices such as magnetic tape, disk, drum, etc. The incorporation of the Geology and Reserves sub-system into the total system involves the establishment of an interface through which data may be readily dissemminated in either direction. This interface will be provided by a geological data library in which all information available from the sub-system is stored and can be readily accessed by the main program and/or other sub-systems. In Figure 7, the information flow between the Geology and Reserves sub¬system and other sub-systems is shown. 5.2 Independent Function Geology and Reserves sub-system is capable of functioning as an independent unit or in direct communication with another sub¬system. This is possible by either continued use of the data library or by redirecting output to other peripheral devices separate from the total system libraries. (Figure 8).

Jan 1, 1974

By David F. Smith

The American Society for Testing Materials4 has recommended that the melting temperature of potassium chromate be used-for indicating the temperature to which a coal sample should be heated in the standard method of analysis to determine its volatile matter content. The specified temperature is 950± 20° C., the melting point of potassium chromate being noted as 940° C. Le Chatelier5 originally found the melting point to be 975°; later6 he found 9400. According to Shemtschushny7 the melting point is 984°; he also found a transition at 679°. Groschuff8 found the melting point to be 971° and the transition temperature 679°. Amadori9 reports a melting point of 9.78° and a transition point of 666°. The International Critical Tables10 take the melting point as 975°, which is estimated to be correct to two significant figures. In view of the apparent uncertainty existing, it has been though worthwhile to make a careful redetermination of this datum. It is of interest not only to determine the melting point of the pure salt, but also of the commercial product, in order to test its suitability for use as a temperature indicator. The work was suggested by A. C. Fieldner, chief engineer, experiment stations division, U. S. Bureau of Mines.

Jan 1, 1929

By Charles L. Kimbell

For 1991, the aggregate value of world international export trade in mineral commodities was roughly estimated at $685 billion in current 1991 dollars. The estimated level was about 6 % higher than that for 1990, but still more than 3 % below the record high set in 1980, and the year was the fifth in a row in which the value of exported minerals advanced. Table 8 provides a data series for the estimated value of world export trade in all mineral commodities from 1979 through 1991, as well as a breakdown of this total into fuel and non fuel commodities. The data shown are based on the historically more traditional and realistic (from the viewpoint of data availability) practice of reporting trade values for the year before the year of review. Despite the considerable increase in the dollar value of mineral commodity exports, their share of total export merchandise trade declined to 19.1 % in 1990 from 19.7% in 1989. At 19.1 %, this share was very near the recent year record low of 19.0% reached in 1988. Comparable 1991 figures were not available in time for inclusion here, but it is expected that the mineral commodity share of the total will be little altered from that of 1990.

Jan 1, 1993

Accurately measure time of detonation and explosive detonation velocity in a blasthole for analysis of explosive performance and delay intervals. Approach A blasthole-monitoring fiber optic probe was developed consisting of filer optic cable to carry the fight and a sensor to detect and convert the light signals into electrical signals. Events can then be recorded on an oscilloscope, tape recorder, digital computer, or start-and-stop timer. This method is simpler to use and more immune to electronic noise than existing resistance techniques. The fiber optic material,, are inexpensive and easy to obtain, and the recording devices, such as storage oscilloscopes, arc available or easy to obtain. The probe provides the mining industry, explosive manufacturers, and those studying blasting, with a practical alternative to complex and expensive coaxial cable reflectometry, resistor strings, or resistance wire methods for determining explosive and detonator performance. How It Works The fiber optic cable is a plastic light guide consisting of sixteen 0.01-inch-diameter optic strands having an overall outside diameter of 0.087 inch. The receiver module incorporates a photodiode to detect the signal and an integrated circuit receiver for signal amplification into one package about 1/2 by 5/8 by 7/8 inch.

Jan 1, 1989

By C. M. Frost

Crude shale oil produced during in situ retorting by the Bureau of Mines at Rock Springs, Wyo., was refined by three schemes. In the first, the crude oil was fractionated to raw naphtha and 400° F+ residuum. The residuum was hydrogenated over cobalt molybdate catalyst at 815° F and 1,100 psig. The blended raw and 160°-400° F hydrogenated naphthas were hydrodenitrified and then catalytically reformed. The 400° F+ hydrogenated oil was catalytically cracked. In the second scheme, the crude was hydrogenated over nickel-tungsten catalyst at 800° F and 1,500 psig. The hydrogenated fractions were processed by methods used in the first scheme. The third scheme resembled the second but, prior to hydrogenation, the crude was "hydrostabilized" over cobalt molybdate at 500° F and 500 psig. Yields of reformate and cat-cracked gasoline as volume-percents of the crude, and research-method octane numbers (with 3 ml TEL) for the three methods were as follows: First method, 32 pet reformate, 90 octane; 19.5 pct cracked gasoline, 99.2 octane. Second method, 29 pet reformate, 81.1 octane; 21.3 pct cracked gasoline, 99.6 octane. Third method, 28.2 pet reformate, 86 octane; 28.7 pct cracked gasoline, 95.9 octane.

Jan 1, 1974

By G. B. Atkinson

The Bureau of Mines, in cooperation with McDermitt Mine, investigated a method for recovering mercury metal from mercury sulfide flotation concentrate by a hydrometallurgical technique. This preliminary report shows the feasibility of leaching mercury with a cupric chloride solution and electrowinning mercury from the solution. Proposed chemical reactions are shown in equations 1 through 3: 2CuC12 + HgS + 2CuCl + HgC12 + S°; (1) electrolysis HgC12 ?Hg + C12; (2) 2CuC1 + C12 ? 2CuC12. (3) A bench-scale leaching-electrowinning apparatus capable of treating 2.36 kg of sulfide concentrate in 11 days was operated continuously. An acid chloride solution containing, in grams per liter, 70 Ca, 60 Cu, 30 Fe, and 80 Hg was used as both leachant and electrolyte; 99.5 pct of the mercury was solubilized from the sulfide concentrate. Electrolysis of the leaching solution recovered mercury metal at an average current efficiency of 32 pct and an average energy consumption of 1.2 kwhr/lb Hg. Results from a preliminary economic evaluation indicated that a plant producing 31,500 flasks of mercury per year would have a capital cost of $3.7 million and an operating cost of $62 per flask of mercury.

Jan 1, 1983

By C. B. Manula, R. L. Sanford, R. A. Rivell

The large amount of computer code required for this simulation model dictates that reasonable limitations be placed on some of the input variables. The program permits the following maximum sizes: 1. Ten mine car stations 2. Twenty track intersections 3. Three hundred mine cars 4. Eight motors 5. One hundred profiles 6. Twelve intersections between stations 7. Forty-five profile changes between stations 8. Thirty station pairs (routes) These may be changed by the user providing the particular computer in use has available core storage. Major Array Descriptions This section is a detailed description of four of the major arrays used in the rail haulage simulator for data storage and logical control. Other arrays are used; however, they may be considered sub¬servative to the major ones. These major variables are explained on the following pages. Motor Control Array (XXXDEF). The first 99 locations in this array are arranged according to and are developed by data input. Unit conversions are made at input time which include: 1. Tons to pounds. 2. Rolling resistance in pounds/ton to percent/100. 3. Miles per hour per second (mphps) to feet per min. per minute. 4. Miles per hour to feet per minute. The following information is stored for each motor at the array

Jan 1, 1974

By L. D. Gehi, A. W. Khair, M. Ahmad

"Bit wear presents problems to mining operations because it affects both productivity and economics. Most bit research deals with energy requirements and energy dissipation mechanics for ideal fragmentation during the cutting process.Mechanical cutting leads to dust generation. Past studies involving dust generation due to bit wear concluded that dust concentrations increased considerably after a 7% bit weight loss: leading researchers recommend that the bits be changed at that point.Bit wear is affected by three factors: machine operating parameters in-situ mining conditions and rock properties. Operating parameters such as the depth of cut tool material spray position and the drum's rotating velocity are very important and have a definite impact on bit wear.During the mining process the amount of rock in coal and the strength of that rock is important as abrasion and adhesion take affect. Diffusive wear or deformation becomes dominant only when the temperatures exceed critical. Temperatures rise as a function of drum velocity rock properties and the cutting tool itself.Miner bit surveyIn this study, 15 mines were surveyed, and of those six responded for nine seams. The survey reflected a wide range of information about coal production and bit consumption from continuous miner sections. Also used bits were collected to get first-hand information to analyze wear."

Nov 1, 1995

By L. Prosser, A. Iannacchione, G. Esterhuizen

The potential for roof falls in underground mines remains a clear and present danger for mine workers. An investigation of ground conditions in nearly 50% of the nation’s underground stone mines found that the state of roof stability is primarily determined in a limited and subjective manner. These large opening mines, with roof heights typically of 7 m or more, make physical observation difficult. Although some mines use monitoring techniques to gain additional information on roof stability, this practice is usually short-term and localized to address ground conditions in a particular section or part of the mine. A methodology to assess the risk for a roof fall is proposed in a preliminary fashion based on engineering judgment acquired from extensive underground stone mine experience and examination of related literature. The proposed method utilizes an observational technique to identify the risk of roof falls in three categories. Case study scenarios provide a realistic picture of model implementation. Providing the mine level decision maker with an accurate assessment tool to ascertain the level of risk related to ground conditions is expected to reduce mine worker injuries and fatalities. Moreover, the presences of danger can be overcome with a clear picture of quantified ground conditions.

By Christopher Mark

Since it was first introduced, the Coal Mine Roof Rating (CMRR) has been widely accepted as a tool for geologic characterization and mine planning. This paper discusses the application of the CMRR to another practical ground control problem. Extended cuts (cuts greater than 6 m (20 ft) in length) are commonly used with remote control continuous miners. Extended cuts can greatly increase productivity, but they have been associated with a number of fatal roof fall accidents. When extended cuts are attempted in weak roof, the roof may collapse before it can be bolted, causing hazardous conditions. Until now, it has not been possible to predict where conditions may not be suitable for extended cuts. In this study, data on the CMRR and extended cut experience were collected at 36 mines in 7 states. It was found that when the CMRR was greater than 55, deep cuts were routine in nearly every case. When the CMRR was less than 37, extended cuts were almost never taken. Between 38 and 55, extended cuts were feasible sometimes but not others. The data also shows that extended cuts are less likely to be stable if either the entry span or the depth of cover increases.

By W. Ayvazoglou

A large number of organizations interested in the search for mineral deposits have adopted the magnetic method of geophysical prospecting, as it has proved to be the most simple, convenient, and rapid, and at the same time the least expensive in comparison with other methods. According to the last statistical data, about two millions square kilo¬ meters were explored in the United States by different geophysical methods. Ninety-six per cent of this area was prospected for oil and 4 per cent for minerals. Detailed exploration by means of the magnetometer has been carried out over about 85 per cent of this surface. The great advantages offered by this method resulted in the fact that the magnetometer was tried out in most places in which geophysical methods were applied. The main purpose of such e:rploration is to obtain the maximum knowledge valuable to the geologist; therefore good geological interpretation of the results of measurements is important. The possibility of drawing practical conclusions from a map representing the results of the study of the terrane depends greatly on the accuracy with which the measurements are made. The purpose of this article is to show the different sources of error by which the work of an operator may be affected and to discuss the means by which the carrying out of a magnetic survey may be improved.

Nov 1, 1931

By W. P. Yant, G. W. Jones

During the testing of porcelain insulators with a " 60 - cycle flash - over apparatus " , gases are liberated having an odor greatly resembling that of ozone . Since the two main constituents of normal air are nitrogen and oxygen , and high voltages (approximately 90,000 volts ) are used during the tests , the conditions are favorable for the production of ozone or oxides of nitrogen , and perhaps both simultaneously . The presence of these gases is only noticeable during the testing period . They soon disappear after the test is over . No fatalities have ever resulted from these gases at the plant visited nor are there any records of any of the men who worked on the tests being confined to their homes from this cause . New employees sometimes complain of a depressed feeling and shortness of breath , which is most noticeable after leaving work and getting into fresh air , but they som become accustomed to it and no ill effects are noticed . In view of previous work*** which the Bureau of Mines has done on oxides of nitrogen in connection with mine gases and problems in gas analysis , tests were made at the plant of the Westinghouse High Voltage Insulator Company at Derry , Pa . to determine the kind and amount of gases liberated .

Jun 1, 1923