Search Documents

By Butcher NM, Walker GS

Argyle Diamond Mines Pty. Limited (ADM), manager and operator of the Argyle Diamond Mines Joint Venture (ADMJV) will recruit approximately 400 employees during 1985 to operate the new large-scale, open-cut diamond mine and treatment plant being developed at Argyle, in the East Kimberley Region of Western Australia. These employees will live in Perth or environs and will commute to work at Argyle via chartered jet aircraft. On site, they will be accommodated in self-contained single person's quarters, with central messing and recreational facilities. Rostering systems have not been finalised, but it is likely employees will rotate through a cycle of 14 days at site and seven-days on leave in Perth. ADM spent three years evaluating residen- tial options before deciding to commute its workforce from Perth. Another 70 senior staff will reside in Kununurra (the nearest established town, 250 km from Argyle) and commute to the mine daily by air.

Jan 1, 1984

The sixth annual conference of the Mineral Economics and Management Society (MEMS) was held Feb. 27 through March 1 in Boulder, CO. This year's event was attended by more than 100 academic, industrial, government and other professionals interested in mineral economics and materials management. With about 250 members worldwide, this not-for-profit society seeks to promote scientific and economic analysis of current relevant issues. This year's program was sponsored by Alcan, Bateman Engineering, Cyprus Amax Minerals, J.P Morgan, Knight Piesold LLC and Newmont Gold. The conference's technical program consisted of six sessions. These included "Financial tools for a changing world," "Access to minerals," "Comparing North and South America," "Environmental issues in today's world" and "Transporting minerals." The program was concluded with a panel discussion titled "National minerals policies in Asia." In addition, a student papers session was held, in which students were given the opportunity to compete for a $1,000 award. The following is a brief review of the technical pro¬gram.

Jan 1, 1997

By G. A. Savanick

Borehole (slurry) mining, where a tool incorporating a water-jet cutter and a downhole slurry pump mines phosphate through a borehole, was field tested in northeastern Florida as part of a US Bureau of Mines program to increase the domestic availability of critical minerals. About 1.5 kt (1700 st) of phosphate ore were mined at 23.6 t/h (26 stph) from a depth of 76 m (250 ft). The borehole was filled with water during mining for roof support and to assist slurry pumping. Underwater cutting was enhanced with an air-shrouded water jet. Mined cavities were backfilled to prevent subsidence after abandonment. Roof subsidence during mining was avoided only when the cavity was filled with water. Environmental monitoring before, during, and after mining included hydrological studies and measurements of water quality. Significant water quality changes occurred only in the zone being mined and were confined to the immediate area of the borehole.

Jan 1, 1986

The Charters Towers 1:250 000 Gravity Data Set represents a new initiative of the Department of Resource Industries (DRI, formerly Queensland Department of Mines) to provide high quality regional gravity data over areas of the State that have high potential for new mineral discoveries. The Department has adopted a strategy to upgrade geophysical data to complement the ongoing Regional Geological Mapping Program (RGMP) of this Department and in areas where the data would be of most benefit to the mineral exploration industry. Part of this strategy is to ensure that these data are released within 6 months of data collection.The 1989 Charters Towers survey was completed as a joint project with the Bureau of Mineral Resources, Geology and Geophysics (BMR), and is the first regional geophysical data acquisition project undertaken by the DRI. Prior to this survey, the only gravity data in this area dates from the B MR regional surveys carried out between 1960 and 1967.

Jan 1, 1990

By G. Robert Frey

INTRODUCTION This chapter on underground rotary drills briefly describes the equipment features, selection or applica¬tion criteria, safety features, and relative costs of both face and roof drills. After a brief description of the basic components, the application, safety, and cost con¬siderations will be explored. ROTARY DRILL COMPONENTS As illustrated in Fig. 1, an underground rotary drill consists of five basic components. These components are the energy system, chassis, boom, feed assembly, and operator's station. Energy System The energy system of a typical underground drill con¬sists of an electric motor, driving a hydraulic pump or pumps. Usually, the drill is equipped with a cable reel to store the trailing power cable from the portable power center in the mine. A simple control box is supplied to house electrical switchgear and protective devices. In addition to the pump, the hydraulic portion of the energy system includes the hydraulic cylinders, hydraulic tram motors, relief valves, control valves, etc., necessary to actuate the various machine functions. Chassis The chassis provides the frame or support structure for the other four basic components of the drill. Fre¬quently, it incorporates an oil reservoir for the hydraulic system. The chasis provides the support for the trans¬mission, which drives rubber tires or, in the case of some foreign machines, the crawler tracks. The steering link¬age and steering tires also may be considered as parts of the chassis. Boom The boom is the structure that supports the feed assembly; it is either hinged or pivoted on the front of the chassis. Hydraulic cylinders on the boom are used to lift or swing the boom relative to the chassis and to swivel or "roll" the feed assembly. Feed Assembly The feed assembly is made up of the support struc¬ture with an integral thrust mechanism, a rotation unit, centralizers to stabilize the auger, and a hose handling arrangement. In addition, most drills have an extension frame with a tilt device to aid in positioning the feed. Operator's Station The operator's station is the area from which the miner operates the drill. UNDERGROUND ROTARY DRILL APPLICATIONS Underground Face Drills Underground face drills are used in mining opera¬tions that utilize the conventional mining system. The face drill is used to drill holes in the exposed face of the material, and the face is then "shot" by means of explo¬sives or high pressure air. The loose material then is ready for loading and transporting from the face to the preparation plant. Typical specifications for under¬ground face drills are listed in Table 1. Certain parameters should be established as to where the drill is to be used and what is expected of it. To create an awareness of some of the considerations that too often are overlooked, some questions are listed that should be asked in conjunction with the selection or application of the proposed drills. Energy System: Can the drill be supplied with the proper voltage to match the available or proposed elec¬trical power system of the mine? Is the system 250, 460, 500, 550, or 950 v; dc or ac; 3-phase; 60 or 50 Hz? Is the drill equipped with an adequate electric motor to handle the proposed duty cycle? Typical motor power on a single-boom drill would range from 25 to 40 kW (35 to 50 hp); the nominal power for a dual-boom

Jan 1, 1982

By W. D. Robertson, R. E. Reed-Hill

DEFORMATION of magnesium crystals in a direction parallel to the basal plane has a special significance as a result of the preferred orientation characteristic of cold worked and recrystallized polycrystalline magnesium sheet and extruded rod in which the basal plane tends to lie in or near the plane of the sheet or the axis of a rod. The principal objective of the present study is a determination of the modes of plastic deformation in magnesium single crystals in this orientation. Evidence for nonbasal slip in magnesium has been presented by a number of authors. Schmid and Was-serman&apos; indicated that pyramidal slip on {1011} < 1120> or {1012} <1120> could occur above 225°C. Bakarian and Mathewson2 stablished that the slip system {1011} <1120> operates at high temperatures and that 225°C was a lower limit for pyramidal slip. The latter authors also showed that the critical resolved shear stress for {1011} slip was 400 g per sq mm at 330°C, while that for basal slip was 66 g per sq mm at the same temperature. Thus, at 330°C basal slip should be expected for all orientations except those in which the stress axis lies almost in the basal plane. More recently Burke and Hibbard3 oserved pyramidal slip at 25°C, when the stress axis was 6" to the basal plane and 14" to a <1120> slip direction. The critical resolved shear stress for pyramidal slip was 52 g per sq mm, which is only slightly greater than their value for basal slip (46 g per sq mm). Chadhuri, Chang, and Grant4 have observed non-basal slip in creep tests with polycrystalline specimens in the temperature range 250" to 350°C. The observed slip bands, which were extremely irregular in appearance, can be explained in terms of cooperative slip on {1011} and (1010) planes which have a common slip direction. Prismatic (1070) slip in polycrystalline magnesium was observed by Hauser, Landon, and Dorn.5 It was found to occur at low temperatures (78" and 195°K) and it was accompanied by cross slip in which the cooperating plane was the basal plane. Since prismatic slip occurred only at regions of high stress concentrations, the critical resolved shear stress could not be determined. Experimental Procedure The present work was performed on crystal specimens which were all cut from the same large single crystal. The original cylindrical crystal, 0.5 x 7 in., was grown in a Bridgman type furnace in which the temperature gradient was moved past the stationary crystal in the manner described by Jillson6 and Burke.&apos; High purity magnesium was furnished by the Dow Chemical Co. with the following spectro-graphic analysis in weight percent: 0.003 pct Al, <0.01 pct Ca, 0,002 pct Cu, <0.001 pct Fe, 0.008 pct Mn, <0.001 pct Ni, <0.002 pct Pb, <0.01 pct Si, <0.001 pct Sn, and <0.01 pct Zn. The large crystal was cut into four cylindrical pieces about 1.5 in. long. A flat surface parallel to the basal plane was made by hand grinding with lubricated metallographic paper on one side of each cylinder. In order to prevent bending in subsequent cutting operations, the crystals were supported by

Jan 1, 1958

The pellet plant, built by Dravo Corp, in Groveland, Mich., for the Hanna Mining Co., is now in full operation. Although using jasper, a type of iron ore with low iron content (approximately 28% Fe), the plant has a rated capacity of 1,250,000 tans of pellets per year. In its four years of existence, the nearby concentrating plant has seen its capacity increase from 700,000 tons to 1,500,000 tons a year as part of a major expansion program at Groveland. Utilizing the Dravo-Lurgi continuous grate process, the pellet plant produces %a-in. pellets of more than 60% iron content. The fully-hooded pelletizing grate, largest in the U.S., is 9 ft. 10% in. wide and 223 ft. 1% in. tong over 34 windboxes, with an effective hearth area of 2195 sq. ft. Drying, firing, and cooling functions are all carried out on this single piece of equipment.

Jan 4, 1964

By G. L. F. Powell

TURNBULL and Cech1 reported undercooling small droplets (approximately 50 mp in diam) of a large number of metals by approximately 0.18 TE (TEoK = the equilibrium freezing point). This was considered to be the undercooling for homogeneous nucleation in

Jan 1, 1968

The paper is a general review of practice in design of sub-level open stopes in Northern Rhodesian copper mines and gives examples of some interesting developments during the last five or six years. It discusses the effect of hangingwall subsidence and the methods used to combat rock pressure on working areas.INTRODUCTIONAll producing mines on the Northern Rhodesian Copperbelt have large ore reserves and are designed for high rates of production. The three mines (Mufulira, Rhokana Corporation, and Roan Antelope) where sub-level open stoping is used, hoist from 250,000 to over 400,000 short tons of ore a month so that any stoping method used must be adaptable to concentrated&apos; mining. Individqal stopes produce from 5,000 to over 25,000 short tons of ore a month.Structure af the ore bearing sediments is simple; faulting in ore is almost entirely absent and copper mineralisation is found in the same beds over great distances...

Jan 1, 1954

By Matthew Janes

Advantages: Clean samples Low sample disturbance Very low spoil Versatile Fast Penetrates all formations Resonant Drill 100 or 150 kW Self tuning to resonance Rotation to 100 rpm Increased depth in dense soil Increased productivity Reduce water use Reduce casing plugging

May 1, 2013

By Michael S. Miller

In 1984, at the request of Bureau of Land Management, the Bureau of Mines studied 8,500 acres of the 20,250-acre Seven Blackfoot Wilderness Study Area (WSA) consisting of all 3,950 acres of MT-024-657A and 4,550 acres of the 4,850 acre MT-024-657B in order to evaluate the mineral resources. Within this report these 8,500 acres are referred to as the study area. Located about 30 miles northwest of Jordan, MT, the study area is underlain by outcrops of Quaternary alluvium and Cretaceous to Paleocene sandstones, shales, clays, and coal. No mines or prospects were found in the study area. Except for some current interest in petroleum and natural gas, there has been no mineral exploration activity in the study area. The study area contains subeconomic clay, placer gold, and coal. Land within the study area is leased or under lease application for oil and gas.

Jan 1, 1986

By Dwight L. Sawyer

To provide technology that can help decrease the dependence of the U.S. aluminum industry on imported bauxite, the main ore of aluminum, the Bureau of Mines is investigating the extraction of aluminum from abundant domestic reserves of kaolinitic clay. Miniplant studies of the continuous leaching of calcined kaolin with boiling 25-pct HC1 acid resulted in a 93- to 99-pct extraction of the acid-soluble alumina in the calcined kaolin feed. Data are presented for an acid leaching study carried out in three or four continuously stirred tank reactors (CSTR) in series, feed rates of 100 to 250 lb/hr of calcined kaolin, and total nominal residence times of 0.7 to 2.3 hr. A material balance for a 176-lb/hr calcined kaolin feed rate in a circuit consisting of three reactors, four classifiers, and five thickeners is given. Data provide detailed compositional and physical information for calcined kaolin, process liquors, and acid-leached residue solids.

Jan 1, 1983

By M. Clair Smith

This report is one of a series being prepared by the Bureau of Mines on exploration, development, mining methods, and costs at various open-pit mines in the United States. These reports have been prepared in cooperation with various mining companies. A similar program, conducted from 1928 to 1939, was generally conceded to have contributed to improving the efficiency of mining operations. The purpose of these reports is to promote development and conservation of our mineral resources by disseminating technological information on mining methods and practices. This report describes exploration, development, equipment, and mining by open-pit methods of the remainder of an ore body that had been previously worked by underground methods. The Tripp pit of the Consolidated Coppermines Corp. is located at Kimberly, Nev., a small mining company town having an elevation of 7,250 feet, 9 miles by road west of Ely, Nev. (fig. 1). The pit extends over the area formerly mined by underground methods from the Morris and Brooks shafts.

Jan 1, 1962

By A. E. Raddatz

The Bureau of Mines is investigating methods for recovering tungsten from domestic resources. A calcine-quench-leach (CQL) technique was developed for the extraction of tungsten from the Golconda, Nev., hot springs ores. The deposit contains disseminated tungsten that cannot be beneficiated by conventional techniques and must be treated by hydro-metallurgical methods. The CQL procedure consists of calcining the ore at 900° C for 3 hours, quenching in a solution containing 1.5 parts of Na2C03 and 1.5 parts of NaOH per part of WO3 in the ore and at a solution-to-ore ratio of 3 to 1, leaching at 200° C and at 250 psig for 3 hours, and filtering and washing the residue. A series of closed-circuit batch tests showed that 92 pct of the tungsten was extracted from a sample containing 2.32 wt-pct W03. The pregnant solution contained 11 g/l W03.

Jan 1, 1982

By Ingemar Hermansson, Thomas Bjerendal, Mattias Grävare, Carl-John Grävare

"AbstractHigh strain dynamic pile tests (HSDPT) on driven piles have been carried out on a routine basis in Sweden for more than 30 years. The Swedish Commission on Pile research established, in the early 1980s, safety factors based on the percentage of piles tested. The safety factors could then be reduced, creating economic incentives for the contractor to carry out testing.Energy formulas adopted by our neighbour countries, Norway and Denmark, were not accepted in Sweden as a reliable method since the dynamic testing method was considered to be a more scientific and reliable method. It has become almost the only in-situ method used to verify geotechnical bearing capacity on driven piles.The typical Swedish geology of soft soils overlaying very hard strata makes it possible to verify high design loads, typically 20 kN per cm2 of pile cross sectional area for steel piles and 1,5 kN per cm2 for concrete piles.Conventionally a certain percentage of the piles have been tested in Sweden. The EC7, on the other hand states a number of piles to be tested, regardless of the total number of piles. The Swedish NA (National Annex) states a maximum distance between tested piles, which results in better distribution of the test piles over the foundation.For shaft-bearing piles in soft clays, dynamic pile tests have to be correlated by static load tests.Historical backgroundOn a research basis stress-wave measurements have been carried out for a long time in Sweden. One example is a project in the city of Gothenburg, conducted by Fellenius and Pejrud (1964). Strain gauges were glued on the reinforcement bars in 60 m long, precast concrete piles at 3 m and 25 m above the pile tip respectively. Evaluation was done with different methods such as dynamic formulas,stress-wave measurements and in addition compared with static load tests.In the 1975 Building Code in Sweden, driving criteria for concrete piles based on stress-wave theory were introduced, as summarized in Table 1. The driving criteria are intended for 3 and 4 tonne gravity hammers. For hydraulic rigs, which have free-falling hammers, the drop height indicated should be reduced to 80 % of the values in the table. The criteria were considered to include a safety factor of 3,0. For 330 - 450 kN pile load, pile with a standard a cross sectional area of 235 mm x 235 mm was used. For 600 kN load, the standard cross sectional area was 275 mm x 275 mm."

Jan 1, 2014

By Gautam Chandra Mondal

An attempt has been made to carry out qualitative assessment of mine water discharges from the different mines of West Bokaro Coalfields of Hazaribag district of Jharkhand, India. A total of 33 mine water samples were analyzed for pH, electrical conductivity (EC), major cations (Ca, Mg, Na, K), major anions (HCO3, SO4, Cl, NO3), and heavy metals (Fe, Mn, Cu, Zn, Pb, Cr, Ni, Co, As). The hydrochemical data has been evaluated in terms of major ion chemistry, solute acquisition processes and suitability of mine water for domestic, industrial and irrigation uses. The pH of the analyzed mine waters varies from 5.27?8.43 indicating acidic to alkaline nature of the water. The electrical conductivity (EC) values varied from 276 µScm-1 to 1598 µScm-1 for measured samples and spatial differences between the EC values reflects the wide variation in lithology, surface activities and hydrological regime prevailing in the region. The anion chemistry of the analyzed water samples shows the dominance of SO4 and HCO3 with a minor contribution from Cl, NO3 and F. The observed high values of sulphate in the mine water of the area may be attributed to the oxidative weathering of pyrites. On average chloride is contributing 5% of the total anionic balance (TZ-). Fluoride and nitrate have low concentrations and on average contributing <2% the total anions. The cation chemistry indicates that 79% of the samples are Ca>Mg>Na>K, while the 6% belong to Ca>Na>Mg>K and 1% of Ca>Mg>K>Na. Weathering of rock forming minerals and cation exchange processes normally control the levels of these cations. Concentrations of some heavy metals (Fe, Mn, Ni, Cr) are found to be well above the prescribed permissible levels recommended for the drinking water. The water quality assessment for drinking purposes indicates that most of the mine water having high TDS, total hardness (TH) and sulphate concentrations. This suggests that mine water of the study area is not suitable for direct use in drinking and domestic purposes and needs a proper treatment before its utilization. The assessment of water for irrigation uses show that the water is good to permissible quality and can be used for irrigation uses. However, high values of salinity, %Na, SAR, RSC and Mg-hazard at certain sites restrict its suitability for agriculture uses. Keywords: mine water; ion chemistry; quality assessment; West Bokaro coalfields

Sep 1, 2012

By József Faitli, Barnabás Csoke, Károly Solymár, Zoltán Hatvani, Dimitri Papanastassiou

In most cases, the sizing of dry or wet mills for ore processing is primarily based on the determination of the Bond grindability index, using the Bond-Kjos-Rowland relations. In several cases, grinding takes place under conditions different from these and the grinding conditions have a great influence on the postgrinding process. This is particularly true for the grinding of bauxites where grinding takes place in alkali, often in higher temperature in alkali Liquor. So it is necessary to develop a testing method providing a reliable grindability index even under these special circumstances and allowing a more precise mill sizing. Therefore we surrounded the grinding crucible of the Hardgrove mill with a crucible furnace to heat it. The new heated mill is suitable to determine the grindability index under temperature-controlled circumstances, in liquids (water and alkali) with different temperatures. The required heating power is 2000 W. In the implemented equipment we installed four 230V/500W electric heating wires made from heat-resistant alloyed material and surrounded by ceramic insulators. The temperature and temperature-gradient are measured by four PT100 platinum thermal resistors. The data acquisition and control program written in the C language in the Lab Windows CVI program environment measures the voltage proportional to the temperature. Determining the grindability indices of hungarian and greek bauxites experimentally, first dry grinding tests were carried out under standard conditions. During wet and alkaline grinding, to 50 g of dry bauxite 85.5 g of industrial alkali with a concentration of about 200 g/l was added. As grinding aid 2 g of CaO (4 % of dry bauxite) was dosed. From the data it can be concluded: - The tested bauxites show a similar behaviour against temperature change. However, the extent of grindability change is different for different bauxites. - Soft Hungarian boehmitic bauxite can be ground in a high-temperature alkaline medium better than hard Greek one. The increase of grindability according to the high-temperature Hardgrove indices: (kHT = Dry o HT,alkaline H /H ): for hard Greek and soft Hungarian boehmitic bauxites: kHT = 1.7 and 2.4 respectively; according to estimated Bond indices (kB,HT =Wi,dry/Wi,HT, alkaline ): kB,HT = 1.6 and 2.1. - The increase of grindability is the most important factor the reaction of kaolinite with caustic liquor and lime and the loosing of the textural structure at 80 ºC and especially at 100 ºC.

Jan 1, 2003

By Gloyd M. Wiles

The property of the Park City Consolidated Mines Co. is situated in the northeast part of the Park City dining district 1 mile east of the town of Park City. The elevation at the collar of No. 1 shaft is 7,250 feet. Park City is a town of approximately 4,500 population lying about 30 miles east of Salt Lake City. The two are connected by an oiled State Highway, which is cleared of snow in winter and maintained in first-class con¬dition by the State throughout the year. The district is served by branches of the Denver & Rio Grande Western and Union Pacific Railroads, each operating on a daily&apos; schedule, except Sunday. These two railroads have constructed 1 ½ miles of joint track from the town to the Park City Consolidated Mines Co. loading tipple, located 200 feet from the portal of the adit that leads to No. 1 shaft.

Jan 1, 1936

By John Chipman, Wayne L. Worrell

Using recent thermodynamic data for the carbides and oxides of tantalum, columbium, and vanadium, the stable solid phases aboue 1300°K and at 1 atm CO(g) pressure in each M-C-O system have been determined. Pourbaix-Ellingham diagrams have been constructed and are used to estimate the minimum temperatures necessary to obtain these metals by carbothermic reduction under reduced pressures. TANTALUM, columbium, and vanadium react with carbon to form the M,C and MC carbide phases. The M,C carbides have narrow homogeneity ranges, but the MC phases possess considerable ranges of homogeneity.&apos; The free energies of formation of the MC carbides were recently measured, and those of MC and M2Ccarbides have been estimated., (The notations MC and MC denote, respectively, the carbide compositions at the carbon-rich and at the metal-rich boundaries of the homogeneous MC phase.) In this paper, these carbide data are combined with thermodynamic data for the oxides to predict the stable phases in the Ta-C-O, Cb-C-O, and V-C-O systems at specified temperatures and oxygen pressures. M-C-O PHASE DIAGRAMS The solid phases in equilibrium above 1300°K for each M-C-0 system are shown in Figs. 1, 2, and 3. Above 1300°K, the vapor would consist entirely of carbon monoxide; and these phase diagrams were obtained by specifying the vapor pressure of CO(g) to be 1 atm. The temperature at which any two solid phases become unstable is specified along the two-phase equilibrium lines in each diagram. To simplify the diagrams, all solid phases in the Ta-C-O and Cb-C-O systems are represented as stoichiometric compounds. However, the homogeneity ranges of VO and VC are indicated, because they form a complete series of solid solutions.3 It also should be noted that the homogeneity range of

Jan 1, 1964

By M. R. Hieb, J. R. Britt, S. P. Harteis, M. J. Sapko, E. S. Weiss, R. K. Zipf

The design requirements for mine ventilation seals have undergone a radical transformation in recent years, and these revisions have greatly increased the cost of the seal designs and their construction. For example, in the past two years, new federal regulations have increased the minimum design requirement to withstand explosion pressures from 138 kPa (20 psig) to 345 kPa (50 psig) or 827 kPa (120 psig), depending on whether the sealed mine volume is monitored continuously or not. Moreover, there is still a possibility that under certain conditions (such as detonations), even higher pressure requirements may be necessary. The ability of a monolithic, stand-alone mine seal to reliably withstand the full range of possible explosion pressures is becoming increasingly difficult and expensive. In an effort to develop a practical alternative, the West Virginia Office of Miners’ Health Safety and Training (WVOMHST) and the National Institute for Occupational Safety and Health (NIOSH) have collaborated in a research effort to develop a practical, economic and safe mine sealing technique that can enable mine seals to meet the full range of new explosion pressure design requirements. The basic idea is to use a barrier of common mine gob and rubble, in combination with a conventional mine seal, so that the pressure resulting from a gas explosion is reflected, absorbed and attenuated, and the pressure on the ventilation seal is reduced. This paper discusses some concepts and preliminary test results for a “passive mine blast attenuator” (PMBA) that can provide a useful alternative to increasingly larger and stronger stand-alone mine seals. Numerical models and full-scale experiments conducted at the NIOSH Lake Lynn Laboratory Experimental Mine (LLEM) show that the use of a PMBA can signficantly reduce the blast pressure and impulse on conventional ventilation seals.

Jan 1, 2009