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By D. F. Zhang, X. Q. Jiang, N. Tahreen, D. L. Chen, F. S. Pan, D. Y. Li

Alloying with rare-earth elements has been proven to be an important tool for magnesium to achieve superior properties via texture randomization. The role of yttrium in texture weakening and the link between texture and formability are fairly well known. Since annealing is a vital part of metal processing which could significantly contribute to ductility, it is important to ascertain the impact of annealing in terms of texture change. Also, an understanding about texture development during pre- and post-deformation annealing is essential to improve the industrial deformation processing as texture is closely related to formability. While there is independent information in the literature regarding the effect of rare-earth addition and heat treatment on the texture formation, no systematic studies have been reported on the collective impact of yttrium addition, annealing and deformation on texture evolution. The present study was, therefore, aimed to identify the inter-relationship among texture evolution, yttrium content, heat treatment and deformation behavior in an extruded Mg-Zn-Mn-Y alloy. Annealing at 450°C for 1 hour and uniaxial compression tests at a certain amount of deformation were performed. The deformed and annealed samples were then subjected to annealing and compression test, respectively, followed by texture measurements. It was observed that yttrium addition effectively weakened the crystallographic texture in all conditions, which was evident from the changes in the microstructures and textures.

Jan 1, 2015

By Antonio Zechini, Giuseppe Iorio, Massimo Grisolia

"The new C Line of the Rome Metro runs northwest to southeast across the city center, with 30 new stations and a total length of more than 15.5 mi (25 km). Due to the presence of a dense urban environment, the selected solution for the construction of the stations consists of a cutand- cover excavation, which is retained by multipropped diaphragm walls to minimize settlement and prevent damage to the nearby structures. The bottom of the excavation of the underground stations are about 82 ft to 98 ft (25 m to 30 m) below ground surface, and about 66 ft to 82 ft (20 m to 25 m) below the groundwater table. To ensure dry conditions and to prevent hydraulic base heave instability, deep well systems were designed and used to lower groundwater level and piezometric head during excavation. The dewatering field tests showed how the local stratigraphy strongly affects the dewatering efficiency. Finite element modeling (FEM) of steady state filtration were used to compare the soil permeability with Lefranc in-situ test results and to verify the efficiency of the grouted plugs. The geological sequence along the new C Line consists of a very heterogeneously layered deposit. The base deposit of stiff overconsolidated clay (“APL,” Pliocene age) is overlain by fluvio-palustrine very dense sandy gravels (“SG,” Pleistocene age), which are then overlain by medium stiff clayey silts or dense sandy silts (“AR” or “ST,” Paleotevere units). These deposits are overlain by covered by pyroclastic volcanic soils, tuff, silty sand and sandy silt (Middle to Upper Pleistocene) deriving from the Colli Albani apparatus. A layer of made ground (“R”) of varying thickness covers the stratigraphic sequence and the natural soil profile everywhere.The local hydrogeological framework is very complex and characterized by a double groundwater system. The upper main aquifer is mainly represented by the “pozzolane sequences” (“PR/PN”), while the lower aquifer consists of a deposit of sandy gravel (“SG,” Pleistocene), which is underlain by marine claystone bedrock (“APL,” Pliocene). The pyroclastic deposits (e.g., pozzolane, lithoid and pseudolithoid tuff and clayely tuff) show large variations in permeability due to porosity, compaction processes, sealing and secondary fracturing. Permeability values for each soil type were determined using Lefranc tests at variable head and using pumping tests. The Lefranc tests indicated a wide range of permeability."

Jan 1, 2017

By D. H. Whiting, A. J. Sinclair

"The San Antonio gold mine, in the Archean Rice Lake Greenstone Belt, was in continuous production from 1932 to 1968, during which period the mine yielded 1.35 million ounces (41 985 kg) of gold and 195,000 ounces (6065 kg) of silver from 4.88 million tons (4.44 Mt) of ore for average grades of 0.28 ounces per ton (9.0 g/t) gold and 0.04 ounces per ton (1.3 g/t) silver. After the 1968 closure, several publications appeared implying that gold grade of ore decreased with increasing depth.Recent observations by the authors show that the relatively low grades in the middle levels of the mine are controlled geologically. However, the reduced average are grade in the lower levels of the mine was due to inadequate mining control that led to a high level of dilution. The similarity of geological factors such as vein styles within structural domains, variations in mining control, recent ore muck sampling results and lithogeo chemical profiles all indicate the similarity of ore in both upper and lower levels.Strict mining control should yield improved grades of gold for future production from the lower levels at San Antonio, in contrast with the production grades of the late 19605that represented extensively diluted ore. High-quality production records are an essential component in re-evaluating a past producer, particularly where workings are accessible. Poorly founded conclusions which cannot be checked by access to workings can have a negative impact on the evaluation procedure."

Jan 1, 1990

By H. F. Feldmann

This work was done to provide thermodynamic, electrical, physical, and compositional data on the seeded combustion plasmas of seven coals for the design of MHD generators. The systems were assumed to be at equilibrium, and the design data were generated by solving for the composition of both gas and liquid phases with a digital computer. The coals reported in this study include a western subbituminous, a Pennsylvania anthracite, and five representative bituminous coals. The data include dewpoint temperatures of both slag and seed. A Mollier diagram for each coal is given to allow graphical generator calculations.

Jan 1, 1970

By A. C. Richardson

The mine of the Porter Coal Co. is located at the town of Porter in the northwestern part of Jefferson County, Ala., about 20 miles west of Birmingham. The operations in this mine are confined to the middle and bottom benches of the Mary Lee bed. The coal from the Porter mine is handpicked and washed before being placed on the market. The handpicked lump coal and the washed egg and nut coal are sold chiefly as domestic and steam-raising fuel. The coal finer than 1 inch is washed in 2-compartment jigs and utilized in the manufacture of metallurgical coke. The coal company has found that, when coal ranging in size from 1 to 0 inch is washed in jigs, a considerable amount of high-ash material finer than 3/16 inch remains in the washed coal. This paper gives the results of an investigation which was conducted for the purpose of determining the washing characteristics of the through 3/16-inch raw coal and the extent to which the ash might be reduced by means of coal-washing tables. This investigation has been made in connection with an extensive study of the coal-washing problems now being conducted by the United States Bureau of Mines and the University of Alabama, with the cooperation of the coal-mining companies of the State.

Jan 1, 1933

By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

By John L. Ham, Robert M. Parke

The melting point of molybdenum is 2625° + 50°C. Heretofore the metal has been considered too refractory to be melted in commercial quantities; hence, it has been formed into rod, wire, and sheet by the methods of powder metallurgy, wherein the temperatures do not exceed 90 per cent of the melting temperature. The manipulation of liquid metals at temperatures of about 2000°C. and above presents some difficult problems in heat transfer, in purification, and in container materials. For metals melting at such high temperatures, it is generally regarded as expedient to circumvent the more common melting and casting process by resorting to powder metallurgy. This circumvention is not accomplished without some penalties, the more notable of which are the relatively high cost of the powder metallurgy method and particularly the size limitations of its product. In an effort to overcome the inherent disadvantages of the powder metallurgy method, especially in respect to the limit of size, a project was begun in 1943 at the Climax Molybdenum Co. laboratory to develop a method of melting and casting molybdenum capable of producing larger bodies of molybdenum than were then obtainable. Later, from Feb. I, 1944, to Aug. 31, 1945, the project was under the supervision of the National Research Committee, Division One. In peacetime, too, there are needs for molybdenurn, which make consideration of another process desirable; for example, large parts for heat engines generally, the omnipresent demand for metal of higher quality and lower cost, and the development of molybdenum-base alloys. This paper describes methods for melting and casting molybdenum, with details of the special process needed to make molybdenum castings ductile, and reports some of the properties of what may be regarded as a new, or at least modified, form of the metal. First Process for Melting and Casting Molybdenum The maintenance of a temperature greater than 2625°C. in a 'pace large enough to melt several pounds of molybdenum may involve the loss of heat at a Very considerable rate. For melting molybdenum, therefore, the source of heat must be of high potential and of fairly high power. Evidently, also, the source of heat must not impair the purity of the product. The electric arc operating in vacuum appeared to meet these requirements and was accordingly the first choice as a source of energy. A number of arc devices for melting molybdenum were examined, and several were tried. In the first experiment, a 20-kw., 60-cycle-per-second alternating-current arc was operated between two horizontal and opposed electrodes of molybdenum- In this arrangement the falling droplets of molybdenum solidified below the arc in an irregular mass. This experiment did not produce a useful casting, but the data obtained indicated that

Jan 1, 1947

By John L. Ham, Robert M. Parke

The melting point of molybdenum is 2625° + 50°C. Heretofore the metal has been considered too refractory to be melted in commercial quantities; hence, it has been formed into rod, wire, and sheet by the methods of powder metallurgy, wherein the temperatures do not exceed 90 per cent of the melting temperature. The manipulation of liquid metals at temperatures of about 2000°C. and above presents some difficult problems in heat transfer, in purification, and in container materials. For metals melting at such high temperatures, it is generally regarded as expedient to circumvent the more common melting and casting process by resorting to powder metallurgy. This circumvention is not accomplished without some penalties, the more notable of which are the relatively high cost of the powder metallurgy method and particularly the size limitations of its product. In an effort to overcome the inherent disadvantages of the powder metallurgy method, especially in respect to the limit of size, a project was begun in 1943 at the Climax Molybdenum Co. laboratory to develop a method of melting and casting molybdenum capable of producing larger bodies of molybdenum than were then obtainable. Later, from Feb. I, 1944, to Aug. 31, 1945, the project was under the supervision of the National Research Committee, Division One. In peacetime, too, there are needs for molybdenurn, which make consideration of another process desirable; for example, large parts for heat engines generally, the omnipresent demand for metal of higher quality and lower cost, and the development of molybdenum-base alloys. This paper describes methods for melting and casting molybdenum, with details of the special process needed to make molybdenum castings ductile, and reports some of the properties of what may be regarded as a new, or at least modified, form of the metal. First Process for Melting and Casting Molybdenum The maintenance of a temperature greater than 2625°C. in a 'pace large enough to melt several pounds of molybdenum may involve the loss of heat at a Very considerable rate. For melting molybdenum, therefore, the source of heat must be of high potential and of fairly high power. Evidently, also, the source of heat must not impair the purity of the product. The electric arc operating in vacuum appeared to meet these requirements and was accordingly the first choice as a source of energy. A number of arc devices for melting molybdenum were examined, and several were tried. In the first experiment, a 20-kw., 60-cycle-per-second alternating-current arc was operated between two horizontal and opposed electrodes of molybdenum- In this arrangement the falling droplets of molybdenum solidified below the arc in an irregular mass. This experiment did not produce a useful casting, but the data obtained indicated that

Jan 1, 1947

By John R. Benham

The Blue Joint Mtn. RARE II area may contain barite resources. One claim, the Woods Creek Barite No. 1, contains two veins of barite of drilling mud quality. The quantity of barite present could not be determined due to limited surface exposures. Sections 1, 2, 11 and 12 of T. 2 S., R. 23 W. in the study area are classified by the U.S. Geological Survey as prospectively valuable for geothermal energy resources. These sections border Blue Joint Creek Hot Springs, which are outside the study area.

Jan 1, 1981

By Dengfu Chen, Mujun Long, Youguang Ma, Xin Xie, Xing Zhang, Leilei Zhang

"Processing parameters of a submerged entry nozzle (SEN) affect significantly on the patterns of fluid flow and temperature distribution in beam blank continuous casting mold with double SEN s. Effect of the geometry structure and immersion depth of double SEN s on the fluid flow in beam blank continuous casting mold was studied by water modeling. Results showed that, the optimal nozzle inner diameter was 40mm, and immersion depth was 70-90mm. The mold surface fluctuation, slag distribution, impact depth and removing rate of inclusion particles were discussed under different processing parameters of double SENs. As the results indicated, slag distribution was more non-homogeneous with a higher casting speed. The results provided basic data for the design and optimization of SENs used in beam blank continuous casting mold.IntroductionNear-net-shape casting has the distinct advantages of less rolling procedures, higher metal recovery and lower energy cost. Therefore, in the recent decades, one of the recent trends in the steel making process is moving toward near net shape casting. in which beam blank casting is a typical technology. The schematic section of beam blank is shown in Fig 1."

Jan 1, 2014

By AIME AIME

OCTOBER will be western month for the Institute. With meetings at Spokane, Tulsa, Los Angeles, and San Francisco, and with a large number of American Institute of Mining Engineers members and their families in the party sailing out through the Golden Gate to Japan on the afternoon of the tenth, the major activities of the Institute will, for this month at least, be west of the Mississippi. Despite cancellations due to necessity, most of the officers of the Institute will be present and taking part in some one or more of the Western meetings. Prof. Robert S. Richards, our senior past-president, will attend both the Spokane and San Francisco meetings and then go to Japan. President Bradley will preside at the San Francisco meeting and then head the party going to Tokyo. At Tulsa and Los Angeles, the principal officers of the Petroleum Division will be present and active and at all the meetings, directors, ex-directors and -present and past officers will be on the job and keeping things moving.

Jan 1, 1929

By T. Price, H. A. Holl, A. P. Greenough, A

UdIN, Shaler, and ulff' first used wires for the determination of the surface energy of a solid metal. A gage length was marked by tying knots in the wires, which were then suspended in a cylinder of the same metal, heated in an inert atmosphere for the appropriate time, cooled to room temperature, and finally removed from the furnace so that the lengths of the wires could be measured. With iron wires, difficulties associated with deformation during the phase changeszp3 made it necessary to measure the gage lengths at the tem-

Jan 1, 1964

By Carson Sharp Beltch

The determination of aqueous cyanide is accomplished by either detection in the sample as collected or by removal of the cyanide of the collected sample, followed by detection. The direct determination of the cyanide content of a solution can only be used in the simplest of systems, since many interferences can result in errors. Thus the most widely used method for total cyanide determination involve the reflux distillation of the sample using heat, acid and vacuum or air flow. The basic theory underlying this analytical approach is that by reducing the pH, the quality of molecular hydrogen cyanide will be in- creased and can be removed from the solution by elevating the temperature. The hydrogen cyanide is then collected in a caustic solution and analyzed. The general approach of removing the cyanide from the sample by distillation represented a major advance in this area of analytical chemistry. It enabled the determination of cyanide in many additional types of samples. When it became apparent that the tightly bonded cyanide complexes were not being detected by this approach, many methods fox complex ion destruction were proposed. These methods usually involved low pH adjustment with oxidizing mineral acids, coupled with the catalytic decomposition of the tightly hound complexes. All off the complexes except the strongest could he broken by these methods. The catalytic reflux distillation step has been found to have some negative aspects, since many materials that can be distilled with hydrogen cyanide will interfere with the color development step and result in a false positive reading.

Jan 1, 1981

By Martin G. Casto

Mixer-settler extractors, among the oldest forms of extraction equipment, have been used since the beginning of this century. Over the past years, a very large number of mixer-settler devices have been designed and operated. Both batch and continuous flow mixer-settlers have been used. The continuous flow mixer-settler extractor offers a great number of advantages over other extraction equipment. The main advantage of the continuous flow mixer-settler is that it is a compact device. A large number of stages can be assembled in a relatively small area, and more stages can be added fairly easily. Many improvements have been made in the design of mixer-settlers. These improvements occurred in almost every phase of mixer-settler' operation; interface control, mixer design, intrastage pumping, interstage flow of the two liquid phases, and stage construction. It is believed that Barstow1 used one of the first continuous mixer-settler extractors. This mixer-settler had concurrent flow within each stage and countercurrent flow overall. The flow for both phases took place by gravity. The Barstow device and other early mixer-settlers had no separate interface control for each stage, therefore, "tilting" of the stages was necessary. In 1934 Holley and Mott2 designed a mixer-settler in which the heavy phase was recirculated within each stage by the use of a valve controlled central line instead of by a pump. Gordon and Zeigler3 in 1939 were among the first to use an adjustable valve to control the interface levels in a mixer-settler, making tilting of the extractor unnecessary. In 1952 Coplan et al. 4 used a modification of the Holley-Mott idea to build a highly improved mixer-settler with flat common walls, and with the

Jan 1, 1971

By L. L. Chasko

A mine fire is one of the most challenging safety issues facing a mine operator and can occur at any location underground. The National Institute for Occupational Safety and Health (NIOSH) and the Twentymile Coal Mine, Colorado, conducted in-mine experiments to determine the capability of high-expansion foam for addressing underground fires. The tests were conducted in sloping entries with high-capacity diesel-powered foam generators. The following results were observed: a well-designed, maintained, and properly operated high-expansion foam generator can propagate a foam plug hundreds of feet in steep upward sloping (20 pct) multiple entries against a ventilation pressure; stoppings and partitions designed to contain the foam plug in upward sloping entries must be substantially constructed; and the predicted quantity of foam concentrate can be significantly less than the actual amount used due to foam losses, such as bubble breakage on dry surfaces. This paper describes the production of high-expansion firefighting foam and discusses the in-mine experimental tests and results. DISCLAIMER: The findings and conclusions in this report are those of the author(s) and do not necessarily represent the views of the National Institute for Occupational Safety and Health.

Jan 1, 2009

By John Parianos

"Seafloor mining provides significant potential economic benefits for many Pacific nations that may be "land poor" but "coastline rich," meaning they have the potential to utilize and profit from the mineral wealth within their Exclusive Economic Zones that extend 200 nautical miles (360 km) out from their coastlines. The Independent State of Papua New Guinea (the State) recognizes this tremendous opportunity and is taking a leadership role in this emerging industry by becoming a 15% partner in Solwara 1 (with an option to increase their investment to 30%).The Solwara 1 Project has been gaining momentum since Nautilus Minerals and the State signed an Agreement on April 24. This new agreement with the State in which they have staked a minimum 15% interest as a meaningful partner, underscores the State’s continued and widespread support for the Project. Signing of this Agreement was a historical occasion and a milestone achievement for Nautilus Minerals and the people of PNG.The machinery being built for the Solwara 1 Project is taking shape. The Seafloor Production Tools are now over 90 percent complete. Assembly of the biggest tool, the Bulk Cutter, was completed in April this year and assembly work on the Collecting Machine and the Auxiliary Cutter commenced in May and June this year respectively. The Riser and Lifting system, which will be used to transfer the mined material from the seafloor to the surface within an enclosed system, is over 50 percent complete. The critical path for the Solwara Project is to secure a suitable vessel arrangement. Nautilus aims to have this completed by the end of this year."

Sep 1, 2014

By Christian Wawrzyniak

An analysis method for modeling compensation grouting based on the finite element method is presented. By this method the design parameters of compensation grouting, grouting pressure and grouting volume, can be evaluated. Also the subsidence of a building due to tunnel driving as well as the lifting of the building by compensation grouting can be simulated. The stress changes in the tunnel lining and in the foundation of the building are computed as well. The application of the analysis method is demonstrated on two examples for compensation grouting in combination with construction of tunnels.

By J. W. Jefferies

"Plant availability up 13%, production output up 40%, tons per employee up 15% and labor turnover down 52%. One mine manager states these were direct spin offs from the work being done to qualify for a safety award in the program known as the ""IMSRS 11 over a three-year period of time. ""Of Course"", stated this manager, ""these were in addition to an 87% reduction in our disabling injusry frequency rate.”Officially known as the International Mine Safety Rating System or the Five Star Mine Safety Rating, a new approach to mine safety is being used with enthusiasm by a significant number of mines in Canada, the U.S., Australia, Chile, Indonesia, the Philippines and South Africa. Significant results are coming into the International Loos Control Institute from over 125 mines from these countries.The program is built around a mine safety audit system that gives positive recognition to a mine at 5-step intervals along the way, (a star with each step attained). The Institute has licensed auditors in locations in Canada and the U.S. available to audit.It is believed the success of this program is the result of the premise upon which the program is based. The results being achieved are due to the concept that a modern, well-managed mine safety program maintains the basic foundation of a complete management system. When properly implemented, it reinforces those factors common to production, quality and safety. (Attitude - skill - knowledge - fitness - structure - machine - equipment - materials and processes). The IMSRS contains elements that reinforce all these factors. It uniquely treats safety as part of the ""right way"" to do the job and not just the ""safe way"" as a distinctly separate entity. Mines using this program now recognize that many of the older safety techniques were not necessarily the right way or the most efficient way to do things. It appears that this new approach to mine safety is a new avenue for the attainment of management's real goals for being in business, and at the same time achieves optimum support for the important target of safety and health for the mine worker."

Jan 1, 1983

By J. Douglas Causey

In 1987, at the request of the Bureau of Land Management, the U.S. Bureau of Mines studied 13,374 acres of the 14,994-acre Ireteba Peaks Wilderness Study Area (NV-050-438) in order to evaluate its identified mineral resources. The study area is located in Clark County, NV, about 4 miles south of Nelson, NV. No identified resources were found within the Ireteba Peaks study area, although there are identified subeconomic resources just outside of the study area boundary. Most of the area is underlain by Tertiary granitic rocks, which apparently postdate mineralization. The area is virtually surrounded by mining districts with mines developed in older rock units. On the north side, quartz monzonite, volcanic tuff, and tuffaceous sediments in the Eldorado mining district host deposits of gold, silver, copper, zinc, and mercury. On the south side, Precambrian gneiss, intrusive andesite porphyry, and hornfels in contact with Quartz monzonite in the Searchlight mining district host gold, silver, copper, and lead deposits. The Newberry mining district to the southeast had copper, gold, and silver mineralization in Precambrian gneiss and andesite.

Jan 1, 1988