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By Harry W. Campbell

In 1986, at the request of the U.S. Bureau of Land Management, the U.S. Bureau of Mines studied parts of the 19,747-acre Lower John Day (OR-005-006) and 7,538-acre Thirtymile (OR-005-001) Wilderness Study Areas in order to evaluate their identified mineral resources. The study areas are located in Gilliam and Sherman Counties, OR, about 13 mi west of Condon. The area is underlain by the Miocene-age Columbia River Basalt Group. The Thirtymile study area is inferred to have 790,000 yd3 of subeconomic gold resources averaging 93.8 mg/yd3 gold ($1.21 yd3 at a gold price of $400 per oz) in two river bars. The Lower John Day study area has five river bars with gold occurrences totaling 1.0 million yd3 and averaging 13.5 mg/yd3 ($0.17 yd3) gold. A preliminary mining feasibility estimate suggests gold could be profitably mined in the study areas from placer deposits containing at least $3.83 yd3 in gold content. Stone, sand, and gravel occur in the study areas, but more accessible deposits are closer to regional markets. No mines or prospects are in or near the study areas; however, one lode mining claim, now abandoned, was located in the Thirtymile study area. Oil and gas leases cover both study areas.

Jan 1, 1987

By R. J. Matetic

Researchers associated with the National Institute for Occupational Safety and Health (NIOSH), Pittsburgh Research Laboratory (PRL), are investigating injuries (fatal and nonfatal) attributed to water well drillers during well drilling activities. Injury data, obtained from various sources, suggest that many water well drillers are experiencing unnecessarily high insurance premiums and very avoidable citations or violations from the direct result of accidents, injuries, and fatalities occurring throughout the water well drilling industry. Expensive insurance premiums and unnecessary citations are the direct result of poor health and safety practices. Visits to numerous water well drilling sites were conducted by NIOSH personnel to characterize and identify the injury data collected and to provide inexpensive strategies to prevent, minimize, and eliminate safety and health issues associated with water well drilling activities.

Jan 1, 2001

By W. L. Anderson

Standard tellurium analytical procedures were modified to obtain the greater sensitivity needed to support a survey of tellurium resources by the Bureau of Mines. Resulting volumetric and spectrophotometric methods for determining tellurium in all types of tellurium-containing materials are described. Sample decomposition and preliminary separations of interfering selenium, gold, and silver are the same in both procedures. The volumetric method is applicable to samples containing a minimum of 10 parts per million of tellurium, In this method the tellurium is reduced to elemental form, and the colloidal suspension is titrated with 0.01 normal iodine solution. The weight of sample for analysis, from 0.25 to 10 grams, should be selected so that the sample contains 0.1 to 5 milligrams of tellurium. To insure accuracy, duplicate samples should be run and the results averaged, particularly on samples of low tellurium content: The limit of determination is 0,001 percent tellurium, which is equivalent to 0.1 milligram of tellurium when using a 10-gram sample. The spectrophotometric method is particularly applicable to low-tenor materials containing from 0.5 to 50 parts per million of tellurium. Up to 10 gram samples are required for making accurate determinations in the low-concentration range. The sample used should contain between 5 and 100 micrograms of tellurium. The turbidity of a colloidal suspension of tellurium prepared from such a sample is determined in a spectrophotometer .t at a. wave-length of 440 millimicrons, and this value is compared with those of standard samples containing known amounts of tellurium.

Jan 1, 1963

By E. G. Davis

The flash reduction of fine iron ore in a dilute-phase system was investigated at 500° to 900° C within a heated stainless steel tubular coil. Alabama brown iron ore and Mesabi semitaconite ore were fed into the coil concurrently with reducing gases of hydrogen and reformed natural gas, so that the velocity of the gas at the reaction temperature was sufficient to entrain and pneumatically transport the solids through the coil. Reduction is very rapid with retention time of the ore in the coil varying from 4 to 15 seconds. Variables affecting reduction are temperature, ore feed rate, and gas flow or velocity. Results of the studies show that a partially reduced product (10-to 25-percent reduction) can be made which, on subsequent magnetic concentration, yields a concentrate containing 64 to 70 percent iron with a 90-to 94-percent iron recovery. Maximum reduction of 77 percent was obtained when hydrogen was used in a four-stage reduction test to yield a product containing 61 percent metallic iron.

Jan 1, 1972

By R. P. Maloney

The economic viability of four depleted mines near Kennicott, Alaska, was determined by the Bureau of Mines to assess their profitability if they were developed today and thus estimate cost data for a medium-sized mine, mill, and town complex. This information should help the State and Federal Governments appraise the potential development of similar-sized mining operations in Alaska. Modern conventional mining and milling techniques were assumed, and I capital and operating costs were derived using flowsheets and standard costing methods. A total capital investment of $37.6 million was necessary to install a 1,000-ton-per-day shrinkage stope mine, flotation-type mill, and support facilities; capital requirements were divided about evenly among the three components. Operating costs, including labor, supplies, power, water, and ' indirect and fixed costs, totaled $42.79 per ton of ore; the mining accounted for nearly one-half the cost. To obtain a 12-percent discounted cash flow rate of return on investment, the analysis showed that the ore could be mined and milled, and the concentrate could be trucked via public roads to a dock facility in Valdez, Alaska, barged to Tacoma, Wash., and smelted and refined for 34.7 cents per pound for copper and 104.0 cents per ounce for silver. The average May 1972 price was 52.6 cents for copper and 158.3 cents for silver. The same set of data was used to examine the economic viability of the mine with the cost of constructing a private road from McCarthy, Alaska, To Valdez, Alaska. The 148-mile road added $53.3 million to the original capital costs and raised the required price of copper to 57.4 cents per pound and the required price of silver to 172.0 cents per ounce. The additional coat of building, maintaining, and paying taxes on a private road from McCarthy, Alaska, to Valdez, Alaska, would make this venture uneconomical.

Jan 1, 1973

By M. J. Ferrante

The Bureau of Mines recovered rhenium powder from scrap tungsten alloy containing approximately 25 percent rhenium. Oxidation of the scrap at 950° C separated volatile rhenium heptoxide from tungsten trioxide. Minor amounts of tungsten, iron, and nickel were removed by selective precipitation from a rhenium solution from which was recovered crystalline ammonium perrhenate. Experimentation showed that hydrogen reduction of minus 325-mesh ammonium perrhenate at 1,000° C produced rhenium powder averaging 2.1 microns in size (Fisher Sub-Sieve method). The powder contained <50 ppm of total metallic impurities and 1,070 ppm of total interstitial impurities. Sintered compacts had a density of 95 percent of theoretical. Approximately 96 percent of the rhenium contained in the scrap alloy was recovered.

Jan 1, 1969

By L. L. Oden

The purpose of this research was to develop a method to remove copper and tin from the magnetic fraction of incinerated municipal refuse, as recovered by the Bureau of Mines pilot plant in College Park, Md. Copper, present in the scrap in a metallic form bonded to ferrous objects, was amenable to leaching in cupric ammonium carbonate solution. Copper levels in small 5-to 10-lb lots of treated scrap were reduced from the initial level, 0.5 to 0.6 pct, to 0.1 to 0.2 pct. Larger lots of scrap (100 lb) were treated less effectively because of poor contact of solution and scrap. Oxidation in the air or in aerated aqueous solution was applied to liberate tin from the dilute solid solution present on the surface of incinerated tinplate. Shredding of air-oxidized scrap or caustic leaching of scrap oxidized in solution reduced tin levels from the initial 0.25-pct level to about 0.1 pct.

Jan 1, 1973

By A. B. Hooker, L. C. IlsLey

"Field inspections and data relative to the use of rubber—sheathed trailing cables in mines show that much cable is ruined or its probable life shortened by being overheated in service.The length of cable required to make connection from a source of supply to a machine at the face workings may be 100 to 500 feet. As a rule, the cable is carried on a reel and unreeled as required; and thus in most cases some of the cable is not removed from the reel. The overheating generally occurs in that part of the cable which is left on the reel.Factors that determine the heating of trailing cables are:1. The size of the conductors in the cable.2. The maximum energy to be supplied through the cable, based upon the horsepower of the motor and the degree to which the motor is loaded.3. The voltage of the supply and the variation of the voltage at the receiving4. The time length of current flow and whether continuous or intermittent.5. The kind of current (direct or alternating).6. The amount of cable left on the reel.7. The type of cable used.8. The kind of reel used.In connection with an investigation of the general performance of rubber—sheathed trailing cables, the Bureau of Mines has made a limited study of some of these factors, including the following points:"

Jul 1, 1931

By E. M. SEELEY, F. S. LOTT

Expansion in total capacity of the natural-gasoline industry.to produce light products continued in 1944 and 1945. The average rate of gain, however, was reduced to about 6 percent annually from over 9 percent in the preceding 4-year period. New facilities at cycle plants again accounted for most of the increase in productive capacity. Between January 1, 1944, and January 1, 1946, the number of cycle plants increased from 32 to 37 and their total capacity from 3,278 thousand gallons daily to 4,848 thousand (48 percent). Capacity of all other plants in the industry increased in the same period from 12,735 thousand gallons daily to 13,080 thousand, or only 3 percent. The capacity of the average plant increased to 28,600 gallons a day from 24,400 gallons on January 1, 1944, continuing the established trend. The average capacity of cycle plants increased to 131,000 gallons a day from 102,400 gallons, and the average capacity of all other natural-gasoline plants increased to 22,200 gallons a day from 20,400 gallons. Although the number of plants increased in only three States in the 2-year period (Texas, Okla- homa, and Arkansas), the total capacity of plants increased in eight States. A further decline of 29 reduced the total number of plants to 626 on January 1, 1946. Compared with 10 years earlier, gasoline plants were more numerous in only 4 States: Texas (48 more), New Mexico (7 more), Arkansas (2 more), and Louisiana (1 more). The largest declines, in order, were in Oklahoma, Illinois, Pennsylvania, West Virginia, and California. The number of plants with fractionating equipment for separating the liquefied gases from heavier fractions was 202 on January 1, 1946 and 182 2 years earlier. The principal changes in number and capacity of shut-down plants were an increase in Texas and a decrease in California. Shut-down natural-gaso- line plants are usually among the smallest units in their vicinity.

Apr 1, 1947

By E. L. Litchfield

Descriptions are given of the techniques and apparatus used by the Bureau of Mines for determinations of flat plate ignition quenching distance and minimum spark ignition energy. The descriptions include discussions of reaction vessel materials, shapes, and sizes and discussions of electrode configurations. Spherical, hemispherical, or flat plate electrodes are recommended. The preferred metal in electrodes and electrode supports is stainless steel; the flat plate electrodes incorporate flanges of low electrical conductivity which are most frequently formed from glass. Various aspects of the electrical energy supply system are discussed and suitable arrangements of components are indicated. The concept of a thermal relaxation time is introduced as a basis for the choice of a minimum interspark time during the testing of a gas mixture. The following indicate the varied conditions under which these tech were utilized: Mixture pressures between 10 mm Hg and 45 psig, initial temperatures between -78° and +198° C, and gaseous mixtures representing wide ranges of cL.. D reactivity and corrosiveness. Minimum ignition energies were determined which varied from about 10-7 joule to nearly 103 joule.

Jan 1, 1967

By A. F. Jolly

In an effort to better predict the useful service life of friction rock stabilizer mine roof bolts in coal mine environments, the Bureau of Mines has evaluated the corrosion resistance of stabilizer steels in air-saturated and deaerated waters from seven Eastern and Midwestern underground coal mines. Accelerated electrochemical corrosion tests were used to estimate corrosion rates for the two high-strength, low-alloy (HSLA) steels used to fabricate Split Set stabilizers and for galvanized steel. Long-term static-immersion weight-loss tests were also conducted with the HSLA steels. Corrosion rates developed from the weight-loss tests (steady-state air dissolution) were roughly comparable to rates determined from electro-chemical testing in aerated waters. Although the highest rates occurred in waters with the highest chloride contents, rates in the other waters were low (<2.0 mpy) and exhibited little correlation to the widely varied chloride contents of the waters. The generally low corrosion rates are attributed in part to the tendency to deposit a protective CaC03 film, reflecting susceptibility to carbonate precipitation as indicated by positive Langelier (saturation) index values for the waters. With one exception, corrosion rates for the galvanized steel were (2.0 mpy. Rates with aerated waters were higher than those with deaerated waters. Pitting tendencies of the steels were also estimated.

Jan 1, 1987

By William K. Tolley

The U.S. Bureau of Mines (USBM) is investigating hydrometallurgical processing techniques to recycle metals from semiconductors and other advanced materials. Cadmium and tellurium were recovered from mixed CdTe/CdS scrap produced in the manufacture of thin-film photovoltaic devices. Selective cadmium dissolution was achieved using oxygen-pressure leaching in dilute sulfuric acid. Leaching the scrap for 90 min at 110° C in 2.2NH2S04 under 400 psig 02 yielded 97% Cd extraction; however, cadmium content of the residue ranged between 4% and 7%. Soluble iron was added to the lixiviant to catalyze oxidation of the CdS component. Tellurium and sulfur remained in the leach residue primarily in the elemental form. The iron and tellurium were removed from the cadmium-rich leach liquor by adjusting the pH to 5.3. The cadmium was recovered as cadmium sulfate crystals by evaporating the solution. Alternative leaching and purification schemes are discussed.

Jan 1, 2010

By A. F. Colombo, D. W. Frommer

Bench- scale tests were made to determine the applicability of flotation as a means of upgrading finished and semifinished magnetic concentrates prepared from reductively roasted iron ores . Both anionic and cationic flotation methods were used to remove siliceous gangue from the magnetic concentrates and produce upgraded products . Seven magnetic concentrates , containing from 6.7 to 13.7 percent SiO2 , were improved by the flotation treatment , yielding final concentrates containing less than 6.0 percent Si02 . Cationic flotation methods were studied on four of the seven samples ; of this group only one sample yielded better results by cationic flotation than by the anionic method .

Jan 1, 1964

By James S. Browning

Domestic production of lithium minerals, source of the lightest metallic element, has increased from an average of 1,327 short tons of concentrate in the 1935-39 period to approximately 37,830 short tons in 1954. This rather phenomenal growth in production and use reflects growing industrial interest in the metal and its salts. This information circular has been prepared to summarize for the casual technical reader the reserve, mining, processing, and consumption situation pertaining to the mineral spodumene - the principal source of lithium. A relatively complete bibliography is included for those who may need more detailed information.

Jan 1, 1958

By Frank W. Sansom

This paper is one of a series of articles dealing with milling methods and costs at metal mines in the United States, written in accordance with an outline prepared by engineers of the United States Bureau of Mines and published by the bureau. The Netta mine and concentrator is the largest of the plants of the Eagle-Picher Lead Co, It is located in Ottawa County, in the extreme northeast corner of Oklahoma, and in what is known as the Tri-State zinc and lead mining district, embracing the northeast corner of Oklahoma, the southwest corner of Missouri, and the southeast corner of Kansas. A majority of the production comes from the Oklahoma section, with Kansas second, and Missouri third.

Jan 1, 1930

By R. J. Bielicki

A sorption apparatus for measuring the equilibrium sorption isotherm and the rate of diffusion of methane from fine-sized coal was designed and constructed. Preliminary work that used Pittsburgh and Pocahontas No. 3 coal shows that the diffusion coefficient varies with pressure and that the fracture spacing in these two coals is substantially different.

Jan 1, 1972

By Thomas L. Chapman

The San Manuel is a deposit of low-grade disseminated copper ore. It was brought to the attention of the government in October 1942 through an application to the Reconstruction Finance Corporation for a development loan, supported by a report by Henry W. Nichols. Nichols' sampling of two outcrops containing copper silicate and his geologic map indicated that a minable body of copper ore might be developed. The Federal Geological Survey was requested to examine the property. The Survey mapped the geology in March 1943 and cut a few check samples, which substantially confirmed Nichols' sampling of the outcrops and a few shallow workings. The geologists of the Survey recommended some test drilling by the Bureau of Mines. The property was examined by the author and Wm. D. Hughes, associate engineer of the Bureau of Mines on March 22 and 26, 1943, and a preliminary drilling campaign was recommended. Five shallow churn-drill holes totaling 1,480 feet were put down under the supervision of Chas. A. Kumke, one of the Bureau's engineers, from November 23, 1943, to January 13, 1944. This work indicated that the deposit extended deeper than the 350-foot depth attained in the deepest hole and might have very considerable areal extent. Additional funds were allotted, and the drilling contract was extended twice to a final maximum of 17,000 feet with 2,000 feet maximum depth of hole. Drilling was continued by the Bureau of Mines until February 2, 1945, when 15,844 feet of hole had been drilled. The Magna Copper Co. took options on this and adjoining property in September 1944 and commenced drilling with two machines. It also engaged the two contract drills used by the Bureau after it stopped drilling. It exercised its options in August 1945 and organized the San Manuel Copper Corp. to operate the property. That company is continuing to explore the ore body with six churn drills. Thomas Lyon took options early in 1945 on holdings adjoining the San Manuel property on the east and is operating two churn drills. Exploration up to the completion of the last holes started by the Bureau of Mines indicated a block of ore 3,100 feet long, 400 to 800 feet wide, and 500 to 900 feet thick, and averaging about 0.8 percent copper. Subsequent drilling showed that the ore body extends considerably beyond this block. The axis of the ore body strikes about N. 61° E. Copper minerals are disseminated in hydrothermally altered quartz monzonite and monzonite porphyry. Ore outcrops in two small areas over the northeast part of the block. Elsewhere the copper-bearing rocks are capped by Gila conglomerate, which thickens toward the west and south. Drilling showed that the ore is in a thick, tabular body with a steep dip to the south. The ore body lies near the surface in the northeastern part of the block. The south side of the block is much farther below the surface than the north side. There the upper part of the copper-bearing rocks is lightly mineralized, so that the hanging wall of minable ore lies deeper than the base of the conglomerate cap, which thickens southward.

Aug 1, 1947

By David P. Conover

The U.S. Bureau of Mines (USBM) has developed a computer-assisted Ground Control Management System (GCMS) for near realtime evaluation of underground coal mine ground conditions. The GCMS combines existing mine-monitoring system and sensor technology with automated computer analysis techniques to enhance the acquisition, analysis, and display of geostructural data. The GCMS, which combines mine structure data with other information, such as geologic conditions and mine layout, has provided information on mining-induced stress transfer and pressure buildup associated with several ground control events. The GCMS collects and analyzes shield-leg pressure, ground pressure, and strata movement data to evaluate the structural behavior of an active Western U.S. Iongwall coal mine. Field data are transmitted over a dedicated phone line to the USBM Denver Research Center (DRC) mine monitoring laboratory, 240 km (150 miles) from the mine site. Real-time collection, processing, and analysis of mine structure and shield pressure information has proved to be an effective method for near-face ground conditions while mining progresses. The GCMS provides the mine operator with information to rapidly identify and manage potentially hazardous ground conditions while mining is in progress.

Jan 1, 1994

By John Durkin

The Bureau of Mines has conducted field studies in coal mines through- out the United States to determine the effectiveness of electromagnetic techniques in locating miners trapped underground following a mine accident. Data from these tests have been used to generate models of expected signal and noise distributions as found above mines throughout the coalfields. These distributions have aided in placing the expected performance of a through-the-earth electromagnetic communications technique into a probabilistic framework. Results show that at a 10-pct false alarm rate, the expected probability of detecting a miner&apos;s signal from a depth of 1,000 ft is 54 pct; at 500 ft it is 95 pct. These depths exceed the actual depths of 90 pct and 50 pct, respectively, of U.S. coal mines. Sensitivity studies have shown that at a depth of 1,000 ft, for every 3 db of improvement in signal-to-noise ratio, the probability of detection increases 6 to 8 pct.

Jan 1, 1982

By John P. Harmon

Vehicles traveling on untreated dirt and gravel roads at open-pit mines disseminate high concentrations of dust (fig. 1). Excessive dust creates safety hazards by reducing visibility and may endanger the health of the men who breathe it. If the dust contains free silica, it can cause silicosis. In addition, maintenance costs of mechanical equipment are increased if equipment is operated in extremely dusty places. Dust on haulage roads in hot, arid country can be controlled if the surface of the road is sprinkled frequently and adequately with water. Often, when roads are sprinkled with water to control the dust, hazards are caused through reduced traction and skidding because of the mud that is formed on the surface of the road. Reducing the amount of water used will help to reduce the hazard of skidding.

Jan 1, 1957