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By Louis C. Brichta

The Capital Mine is one of, the more important zinc mines in Lawrence County, Mo. Owing to the urgent need for additional zinc during the war and the available facilities for the treatment of zinc-lead ores at the property, the Bureau of Mines examined the mine during July and August 1942 and selected it for further investigation by churn drilling, which was accomplished in the fall of 1943 and spring of 1944. At the conclusion of the drilling on the Capital holdings, the Greenland property, immediately south, was investigated by drilling two additional holes. Results of the investigation are presented in this report.

Jan 1, 1948

By I. Gaballah

Cadmium is toxic to plant, animals and humans. The council directive of the European Communities No 8315 13lEEC has fixed an upper limit of 200 ppb for the release of cadmium in industrial effluents. The conventional method of neutralization/prccipitation ( NIP ) is inefficient to reach this limit. The objective of this study is to test the efficiency of the trees modified bark to remove cadmium cations from industrial effluents after the conventional treatment by NIP. Laboratory scale tests with synthetic solutions show that the cadmium removal depends upon pH and initial concentration. About 99.9 %, 99 % and 96 % of the metal are removed from synthetic solutions containing 10, 100 and 1000 ppm respectively. Pilot plant trials ( 1 m3 ) with an industrial effluent bearing about 500 ppb of cadmium lead to the almost complete elimination of the cadmium ( final concentration is lower than the analytical detection limits < 20 ppb ).

Jan 1, 1994

By B. Fazan, O. D. Sherby, J. E. Dorn

McLean&apos;s technique was employed to determine the effect of temperature on the contribution of grain boundary shearing to the total creep strain in pure aluminum over the range of 610° to 747°K. The contribution of grain boundary shearing at the same stress was found to depend upon a temperature-compensated time, where the activation energy AH was practically identical with that for total high temperature creep. SEVERAL outstanding investigations by McLeanl-&apos; have demonstrated that the total creep strain in polycrystalline aluminum at 473°K arises from microscopically detectable slip, subgrain tilting, and grain boundary shearing. Whereas microscopically detectable slip arises from dislocations that move out of the grains, subgrain tilting is due to entrapment of dislocations of like sign in subgrain bound-arles. Consequently, creep in polycrystalline aggregates appears to be the resultant of two mechanisms, namely, migration of dislocations and grain boundary shearing. This suggests that an appropriate law for creep should contain two additive terms ?1 = ?d + ?yb where the total creep strain, ?1, depends on the sums of the creep strains arising from motion of dislocations, ?d, and from grain boundary shearing, ?yb. But extensive investigations5 over wide ranges of alloys, grain sizes, temperatures, and stresses have shown that the total creep strain is quite accurately dependent on the single term functional relationship ?1 = f (?), s = constant where s is the stress (constant),* ? = te-?H/RT, t is the time under stress, AH is the activation energy, R is gas constant, and T is absolute temperature. Since Eq. 2 was found to be valid over those ranges where appreciable grain boundary shearing is observed, it appears probable that the activation energy for grain boundary shearing might be identical with the activation energy for migration of dislocations. Thus, under a given stress, it could be thought that where fd and fgb represent the appropriate functions for each mechanism of creep. The primary objective of this investigation was to obtain a direct check on the accuracy of this hypothesis. Experimental Procedure and Results Rolled aluminum sheet (99.987 pct Al), 0.100 in. thick, was used in this investigation. Creep specimens having a 0.25 in. wide, 3 in. long gage section were so machined that their axes were in the rolling direction. All specimens were annealed for 9 hr at 894°K before testing to produce a mean grain diameter of about 0.1 in. One face of each specimen was polished electrolytically in 10 pct HBF4. In order to permit accurate measurements of grain boundary shearing, the polished surface was lightly scribed with a ruling machine accurate to ±0.0001 in. to give a uniform grid of lines which were 0.01 in. apart. Although it was thought that surface working might modify the creep behavior in the vicinity of the scribed grid lines, this technique nonetheless was employed because such local differences in creep properties should not modify seriously the fractions of the total creep strain due to grain boundary shearing or to migration of dislocations. All creep tests were conducted under constant load with an initial stress of 250 psi. The specimen tem-

Jan 1, 1955

By Mark D. Tabak

Measurments of the transient photoconductivity in fillns of amorphous selenium with blocking- contacts haue been used in studying the transport properties. The results shozu that the transport of free carviers cannot be explained in terms of a model using bulk trapping. Measurments of the photocurrent-applied field characteristic as a unction of sample thickness do not pvoduce the scaling lau,s approps-iate lo a range-limited model. THE investigation described here was undertaken to explore the mechanisms of transport in amorphous selenium films. These films are used in xerography, where a photoconductor is charged electrostatically and then exposed to light to form a latent image. The sensitivity that one can expect when operating in this mode is directly related to the transport properties of the photoconductor material. The study of photoconductivity at high fields in insulating materials is often complicated by contact effects. Injection from the contacts can easily mask the fundamental transport processes that are being studied. A technique is described in this paper that allows a study of photoconductivity and transport processes in amorphous selenium films without the complicating effects introduced by contacts. Accordingly, the samples are prepared using insulating films between the sample and the metal electrodes, where the electrodes are capacitively coupled to the surfaces of the sample under study. A transient measurement is performed with pulsed light, and the photocurrent measured near the beginning of the pulse before the field is distorted by polarization at the blocking elec- trodes. A similar method has been described by Manyl using pulsed electric fields in studying the high field transport in insulating cadmium sulfide. EXPERIMENTAL It is possible to eliminate undesirable contact effects and maintain uniform high fields across the sample by using blocking electrodes and measuring the current-field characteristic under transient conditions. The sample configuration is shown schematically in Fig. 1. Each of the gold electrodes is separated from the amorphous selenium film by a layer of insulating plastic (Formvar) approximately 2000~ thick. The electrode on the front surface is semi-transparent to allow illumination of the sample. The whole configuration is mounted on a KEL-F substrate, which is a transparent insulating plastic having a thermal expansion coefficient that allows low-temperature measurements without cracking the selenium film. Fig. 2 shows a block diagram of the measurement apparatus. Light was obtained from a 450-w Xenon arc lamp, the intensity of which was controlled by stainless-steel screen filters and monitored by a pho-tomultiplier which was calibrated periodically against a thermopile. A camera shutter operated at 1/200 sec provided the light pulse and a synchronized signal to the oscilloscope. The photocurrent transient is ob-served with a high-speed picoammeter (response time of 1 msec) the output of which is displayed on an oscilloscope. The applied voltage is constant throughout the transient measurement. The selenium films were prepared by evaporation of four 9&apos;s pure selenium shot supplied by Canadian Copper Refiners. Spectrographic analysis showed about 1 ppm of Te and Fe, 5 ppm of C1, and 0.05 ppm

Jan 1, 1968

By Chas Kemsley

NO. 6 shaft at Dome Mines, Limited, is an internal shaft going down vertically from the sixteenth, or 2,000-ft., level to the twenty-ninth, or 4,000-ft., level, with loading station and sump below this. It is located approximately 4,600 feet east of No. 3 shaft, the latter being Dome&apos;s main shaft through which ail ore, men, and material are handled. The decision to sink this shaft was made in May, 1935, and work was commenced almost immediately. To reach the site of the new shaft, 3,100 feet of cross-cutting, and 1,400 feet of slashing to widen existing cross-cuts, was necessary. On account of the long haul and the large amount of material to be handled, trolley haulage was decided upon and installed as soon as possible to facilitate the removal of waste rock from the crosscut and other necessary excavations. Figure 1 shows a cross-section of the cross cut and the manner in which the trolley wire is supported and guarded. For hanging each individual cross-piece, two holes are drilled into the roof and filled with snug-fitting wooden plugs. Into these, %-in. ragged hanger bolts arc driven by means of a stoper. The bolts are of various lengths to allow for irregularities in the face of the rock, and they have sufficient thread to permit several inches of adjustment for proper alignment of the cross-pieces. These hangers are spaced at 8-ft. intervals except at the curves, where 5-ft. spacing is used. A 2 in. by 6 in. cedar plank is fastened to the cross-pieces directly over the centre of the track. To this the trolley-wire supports are bolted, and 1 in. by 10 in. cedar boards are nailed on each side to act as guards. Two power-supply cables, one telephone cable, and one .Y2-in. conduit lighting circuit are also supported on the trolley-wire hangers, as shown in Figure 1. A 20-in. ventilating pipe and a 10-in. air line are placed as shown, thus acting as a guard to prevent workmen from walking on the track.

Jan 1, 1938

Pillar design is the first line of defense against rock falls—the greatest single safety hazard faced by underground coal miners in the United States and abroad. To help advance the state of the art in this fundamental mining science, the National Institute for Occupational Safety and Health organized the Second International Workshop on Coal Pillar Mechanics and Design. The workshop was held in Vail, CO, on June 6,1999, in association with the 37th U.S. Rock Mechanics Symposium. The proceedings include 15 papers from leading ground control specialists in the United States, Canada, Australia, the United Kingdom, and the Republic of South Africa. The papers address the entire range of issues associated with coal pillars and have a decidedly practical flavor. Topics include numerical modeling, empirical design formulas based on case histories, field measurements, and post failure mechanics.

Jan 5, 1999

By Albert F. S. Seboukens

A 5;6 MVA pyrometallurgical demonstration-scale pilot plant was commissioned at Mintek, South Africa, in 1994, using lead blast-furnace slag and coke as raw materials. The plant included a.d.c. power. supply, two plasma-arc furnaces&apos; and an ISP lead-splash condenser. Slags containing typically 12 per cent zinc oxide and 3 per cent lead oxide are produced as a waste during the smelting of lead in a blast furnace. These slags are a potential: resource for the recovery of lead and zinc. During 1994, about 600 t of lead blast-furnace, slag was proceced through the pilot plant using the Enviroplas process. Zinc and lead oxides were reduced to their respective metals, volatilized, and collected in an. ISP lead splash condenser. Prime Western grade zinc and slags disposable in accordance with current US legislation, were produced. Zinc condensation efficiencies of about 80 per cent were achieved.

Jan 1, 1995

By Ben V. Petersen

Surveying is to tunneling, as a steering wheel is to a car—somewhat essential if you want to end up in the right place. The principles of tunnel surveying have remained unchanged since the first tunnels were constructed: To mathematically connect the selected start and end points of a tunnel and physically translate that alignment onto the ground. The alternative is just dead-reckoning and leaves the probably less than satisfying end results to chance. Today’s laser-guided tunnels often have to thread the needle through a maze of soil types, utilities, structural features and other underground considerations. Surveying makes it possible.

Jan 1, 2005

By Wenjun Ju, Xiaodong Sun, Jianwei Zheng, Bingding Yi, Zhaorui Ma, Pengfei Jiang, Lingfu Zhu, Yangfa Zheng

The presence of hard roof poses a significant threat to the safety of underground coal mining. In this study, a mechanical model is established for underground stopes to analyze the overburden structure with the advancing of the mining face. It is indicated that the load-bearing arch structure (LBAS) is formed above the stope for roofs with regular hardness, whereas the reversed-basin bearing structure (RBBS) is formed for hard roofs. Compared with the LBAS, the RBBS can induce an additional pressure at the foot of the bearing structure, and this effect becomes less as the distance between the hard roof and stope increases. The ultra-deep borehole hydro-fracturing technique (UHT) is proposed to weaken the hard roof for pressure release, and the associated equipment and procedures are also presented. The UHT is based on horizontal drilling along the axial direction of the roadway in the roof, and then the receding segmented hydraulic fracturing technique is used to achieve the purpose of fracturing (weakening) of the surrounding rocks. Field tests are carried out in 3022 roadway of Daliangwan coal mine in Yulin to assess the developed technique, equipment, and fracturing procedures. To evaluate the fracturing effect, deep borehole stress-meters are installed at different depths of the fractured and un-fractured zones of the roadway, along with hydraulic support pressure sensors placed at the mining face. The results show that compared with the un-fractured zone, the pressure on the wall and pillar sides is lower than that in the fractured zone, and the associated weighting duration and weighting area ratio of the fractured are reduced by 53.23% and 33.59%, respectively. This indicates that the presented UHT can provide a new solution for the prevention and control of mining disasters induced by hard roof failures in underground coal mining.

Nov 2, 2020

By Douglas R. Piteau, Dennis C. Martin

Instability of rock slopes may occur by failure along pre- existing structural discontinuities, by failure through intact material or by failure along a surface formed partly along discontinuities and partly through intact material. Although certain fundamental failure modes are recognized, the mechanisms of slope failure are varied and complex. Such mechanisms are governed by the engineering geology conditions of the rock mass, which are almost always unique to a particular site. An understanding of failure mechanisms requires a knowledge of the physical, mechanical and strength properties of the intact material and discontinuities which make up the rock mass, as well as the structural geology and hydrogeology. These engineering geology parameters also must be evaluated with respect to the slope geometry to determine failures which are kinematically possible. Only after obtaining a reasonable appreciation of the possible failure modes can a rational mechanical stability analysis be carried out. Both two-dimensional and three-dimensional failure mechanisms should be considered in assessing the design of rock slopes. Simple analyses methods are used initially to identify those possible failures which could control slope stability. More complex and detailed analyses usually are required for a few critical failure modes which could control stability, or where a more complex type of failure mechanism is envisaged. Complex failure mechanisms are usually identified when assessing previous slope failures. More detailed information concerning slope geometry and engineering geology parameters must be acquired where analyses must be more rigorous.

Jan 1, 1983

By John D. Sullivan

In leaching copper ores knowledge of the rate of dissolution of the copper minerals and of the impurities is desirable. With knowledge of these two factors and data as to what minerals are present in the ore the metallurgist has an excellent start on planning a leach¬ing cycle to fit that ore. The rate of dissolution is rather difficult to predict, and the time required to leach a given ore may differ markedly from that required by another ore. The whole study as related to mixed oxidized and sulphide ores is one of relative rates of dissolution of the different copper minerals and the impurities in sulphuric acid and in ferric sulphate. The aim of leaching is to choose a cycle and strength of solvents to ex¬tract a maximum amount of copper and a minimum of impurities. The consumption of reagents per unit of copper dissolved is a function of the strength of the reagents used and the relative rates of attack on copper minerals and impurities. An opportunity may exist of ob¬taining the kind of electrolyte wanted for electrolytic precipitation and yet of saving on the consumption of reagents by modifying the strength of ferric sulphate and of sulphuric acid during the cycle to fit the time needed for dissolving the sulphide and oxidized minerals present. Former papers have considered the rate of dissolution of azurite, malachite, tenorite, chrysocolla, cuprite, chalcocite, boruite, and covellite in common leaching solvents. A paper soon to be written will consider some of the less important copper minerals. The present paper considers the rate of dissolution of copper from ores of the Southwest in sulphuric acid of varying strength and the ratio of sulphuric acid consumed to copper dis¬solved. In this study two ores containing the copper principally in an oxidized condition were used. A study of other ores, especially those containing high percentages of sulphide copper minerals, would be desirable, but since consumption of acid has paramount interest,

Jan 1, 1931

By John D. Sullivan

In leaching copper ores knowledge of the rate of dissolution of the copper minerals and of the impurities is desirable. With knowledge of these two factors and data as to what minerals are present in the ore the metallurgist has an excellent start on planning a leaching cycle to fit that ore. The rate of dissolution is rather difficult to predict, and the time required to leach a given ore may differ markedly from that required by another ore. The whole study as related to mixed oxidized and sulphide ores is one of relative rates of dissolution of the different copper minerals and the impurities in sulphuric aoid and in ferric sulphate. The aim of leaching is to choose a cycle and strength of solvents to extract a maximum amount of copper and a minimum of impurities. The consumption of reagents per unit of copper dissolved is a function of the strength of the reagents used and the relative rates of attack on copper minerals and impurities. An opportunity may exist of obtaining the kind of electrolyte wanted for electrolytic precipitation and yet of saving on the consumption of reagents by modifying the strength of ferric sulphate and of sulphuric acid during the cycle to fit the time needed for dissolving the sulphide and oxidized minerals present.

Jan 1, 1931

By Alan R. Buehler

A mineral survey of the 20,792-acre Sand Butte WSA (Wilderness study Area), ID-57-8, was conducted by U.S. Bureau of Mines personnel in 1985, at the request of the Bureau of Land Management. The WSA is in Lincoln County, south-central Idaho, about 15 miles south-southeast of Carey. There are no mines, prospects, or claims within the WSA. Basaltic lava is the only rock type exposed within the WSA. The only mineral resource is a few tons of a platy variety of lava (slab pahoehoe) which is suitable for use as decorative stone. Similar slab pahoehoe, produced from other lava fields in southern Idaho, retails for approximately $125/ton. The deposit in the WSA is too small to be mined commercially but may have value to consumers who gather their own decorative stone. INTRODUCTION A mineral survey of the Sand Butte WSA (Wilderness Study Area) 1/ was conducted in June 1985 by the USBM (U.S. Bureau of Mines) and the USGS (U.S. Geological Survey) at the request of the BLM (U.S. Bureau of Land Management). The USBM examines individual mines, prospects, claims, and mineralized zones, and evaluates identified mineral and energy resources. The USGS evaluates potential for undiscovered resources based on areal geological, geochemical, and geophysical surveys. Results of the investigations will he used to help determine the suitability of the WSA for inclusion into the National Wilderness Preservation System. The immediate goal of this and other USBM mineral surveys is to provide data to the President, Congress, government agencies, and the public for land-use decisions. The long-term objective is to ensure the Nation has an adequate and dependable supply of minerals at a reasonable cost.

Jan 1, 1986

By J J. Robinson, B N. Noller, D M. Franks, G McIlveen, D R. Mulligan, A T. Soro

Fiji is a developing nation that relies on economic opportunities and recently experienced increased interest in mining and mineral exploration that could potentially benefit the country. In order to understand the lack of, or failure of, past rehabilitation activity, this study applied risk-based approaches to identify the significance of activities or actions applicable to rehabilitation of post-mined land in Fiji. This involved the use of qualitative, semi-quantitative and quantitative risk methodologies, which allow assessment and evaluation of mechanisms of mining administration and practices, as well as rehabilitation methods and approaches that have been applied successfully to post-mined land in other countries with similar environmental and climatic conditions and social and regulatory environments to Fiji. The semi-quantitative risk-based technique followed an earlier approach applied for mine closure by Laurence (2006) and provided a systematic way to assess and evaluate rehabilitation issues that were identified on-site by assessors, who were able to estimate through expert judgement the likelihood and consequences based on conditions, situations observed and circumstances on-site. It demonstrated that experienced professionals could reproducibly score likelihood and consequence of selected items relevant to mine site rehabilitation, despite the wide-ranging issues identified during post mined land rehabilitation. The quantitative risk-based approach was adopted as a method to determine environmental bonds for mining projects using the assessment of consequences and likelihood results from the semi-quantitative approach, then incorporating unit cost of materials or activities for local conditions required to fully rehabilitate each issue at net present value. Application to both an abandoned mine site and a newly opened mine site in Fiji showed that this approach could provide a risk-based estimate for rehabilitation cost and demonstrated potential applicable in the Fijian context.CITATION:Soro, A T, Mulligan, D R, McIlveen, G, Franks, D M, Robinson, J J and Noller, B N, ?2014. Risk-based Rehabilitation Strategies for Post-mined Land in Fiji, in Proceedings Life-of-Mine 2014 , pp 165–180 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By Richard Hamburger

Calculating machines permit the use of the more rapid cotangent and semigraphic solutions of plane triangulation. The results of these methods are as accurate as those of other methods. Simple adjustment of the net, and ease of calculation encourage more complete field work. Errors in field work or in computation are easily localized.

Jan 1, 1950

By Arthur L. Crawford, R. E. Head, Fred E. Thackwell, A. Lee Christensen

"The data in this report are the result of a detailed, microscopic study of composite samples of mill products collected over a 14 day period. The samples were collected and assembled under the direction of W. R. Lindsay, superintendent of the Anyox plant, and are fairly representative of the heads, middlings, tailings, and products peculiar to the scheme of concentration practiced at Anyox in 1930.The work was done in the Cooperative Microscopic Laboratory at Salt Lake City, Utah. This laboratory was equipped by the mine and mill operators of the Salt Lake Valley and functioned under a cooperative agreement between the United States Bureau of Mines the University of Utah. This is the fourth of a series of papers 6/ dealing with a microscopic study of mill products."

Oct 1, 1935

By W. M. Merritts, Charles Findlay, Stephen Krickovic

Methane emission rates were studied in a Pittsburgh coalbed mine in northern West Virginia. A fullface boring-type continuous miner, equipped with a methane monitor, was used in the development of a set of eight main headings, one side of which was near old workings; the other side abutted virgin coal. Two air splits ventilated the section. Air volumes and methane percentages were recorded, and time studies of the miner were made during five consecutive operating days, during which a complete cycle of mining the headings and one line of related breakthroughs was accomplished. Methane emission rates increased significantly as mining progressed from the side near old mine workings toward the virgin area, and generally increased with coal extract ion. Mining in the heading immediately adjacent to virgin coal was interrupted quite frequently, owing to excessive methane concentrations at the face, despite adequate air volumes and acceptable methane concentrations in the immediate return airway.

Aug 1, 1970

By Donald E. Archibald, Gerald A. Allcott

"Noranda&apos;s Heath Steele Mine is located 35 miles north of Newcastle in northern New Brunswick. Blasthole open stoping with partial delayed filling is the major mining method in use. The selection of stope and pillar dimensions to ensure mine stability was originally subjective. This paper describes the progress made in attempting to rationalize design.One major feature is the fracture pattern. The main orebody now being mined, the B zone, has undergone five periods of folding and each period of folding developed a family of fracture planes which are axial planar to that particular fold pattern.The other major features considered are the calculated increase in stress due to mining and depth and its relationship to measured strain. Experience has shown that ground problems affecting production have usually originated in rib pillars. Field observations of ground deterioration and multi-wire extensometer measurements of strain are tabulated together chronologically for each pillar.A factor representing stress and a common plane of weakness are identified. Graphical analysis is used to determine the effect of the plane of weakness on a typical stress/strain curve and provides a base for the design of pillars."

Jan 1, 1981

By J. Douglas Causey

In 1986 and 1987, at the request of the Bureau of Land Management, the U.S. Bureau of Mines studied 73,010 acres of the 76,480-acre Lower Owyhee Canyon Wilderness Study Area (OR-003-110) in order to evaluate its identified mineral resources. The study area is in Malheur County, OR, about 8 miles north of Rome. There are approximately 7.7 million tons of subeconomic zeolite (heulandite and/or clinoptilolite) resources with ammonia ion exchange capacities as high as 1.75 milliequivalents per gram in the Chalk Basin-Lambert Rocks area. The bed is 25 ft thick with less than 100 feet of overburden. This resource probably will not be mined in the foreseeable future.

Jan 1, 1989

By Vernon R. Miller

The Bureau of Mines developed and evaluated a detector-sorting device to separate the copper-bearing portion from the barren portion of native copper ores. Small-scale sorting tests were run on minus 1-inch plus 1/2-inch size materials from the following Michigan samples: Winona, Kingston, and Centennial ores, and "poor" rock from the Champion mine dump. Results showed that 60 to 85 pct of the copper was recovered in 9 to 19 wt-pct of the sorted fraction. About 90 pct of the copper recovered was obtained in the first pass through the sorter. For the Kingston and Centennial ores, combinations of sorter concentrates and fines generated during crushing, resulted in recoveries of 93 and 94 pct, respectively, of the total copper in the samples. The combined products constituted about 66 wt-pct of the total sample and allowed the balance to be discarded with only a 6 to 7 pct copper loss.

Jan 1, 1974