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By Bever M. B., Floe Carl F., Hung Liang

Data on the solubility of hydrogen in iron-silicon alloys are of practical interest, as hydrogen causes gas unsoundness and embrittlement in iron and steel and is also a factor in the metallurgy of cast iron. Zapffe and Simsl have recently demonstrated that in deoxidized iron and steel castings gassiness and bleeding are primarily functions of the hydrogen content of the liquid metal. These authors also suggested convincingly that ordinary melting and casting practice provides several sources of hydrogen for absorption by the metal. Norbury and Morgan2 and Boyles3 have shown that dissolved hydrogen affects the graphitization of cast iron. Schneble and Chipman4 included in a study of the effects of superheating on the properties of cast iron an investigation of the part played by hydrogen. Schwartz, Guiler and Barnett5 investigated the significance of hydrogen in the metallurgy of malleable cast iron. A complete interpretation of investigations of this kind calls for equilibrium values of the solubility of hydrogen in the liquid and solid phases. From the theoretical standpoint, the solubility of gases, especially hydrogen, may reveal some aspects of the nature of liquid metallic solutions. This was implied in the early work of Sieverts6 on the effect of alloying additions on the solubility of hydrogen in copper. Recent work by Bever and Floe7 has shown that in liquid copper-tin alloys the solubility of hydrogen changes abruptly at the composition of the intermetallic compound Cu3Sn. It is of considerable interest to determine whether liquid alloys of iron also have compositions at which the solubility of hydrogen exhibits a discontinuous or otherwise unique behavior. Among the various alloys of iron the iron-silicon system is important because of the presence of silicon in various structural alloy steels, transformer sheets, cast irons and special acid-resisting castings. The various grades of ferrosilicon, finally, cover almost the entire remainder of the composition range. Published information on the solubility of hydrogen in iron-silicon alloys is meager. The first values for pure liquid iron were reported in 1910-191 1 sieverts and Krumbhaar8 and by sieverts.9 They found that 100 grams of iron absorbs 28 c.c. of hydrogen at 1550°C and One atmosphere pressure, and 31 C.C. at 1650°C. In 1938 Sieverts, Zapf and Moritz10 published a value of 26.3 C.C. at 1550°C. Data on the solubility of hydrogen in liquid iron-silicon alloys and in liquid silicon are not available. Martin1' investigated

Jan 1, 1947

By Jelena Milinovic, Sofia Martins, Anna Lichtschlag, Sven Petersen, Fernando J. A. S. Barriga, Bramley Murton, Adeline Dutrieux

"The active TAG hydrothermal mound is one of the largest and better studied. Due to its size (one of the largest, known so far, in the Atlantic Ocean) and concentration of specific economic valuable elements, it might be considered for future exploitation due the increasing demand for raw materials. However, the discovery of several large inactive sulfide mounds, some partially buried under sediments may relieve this pressure. Exploration for such occurrences can, together with geophysical exploration tools, be accomplished by mineralogical and geochemical studies of those sediments.INTRODUCTION AND GEOLOGICAL SETTINGThis study reports the results from sediment samples collected during the Meteor M127 cruise on the TAG hydrothermal area in 2016. The Trans-Atlantic Geotraverse (TAG) hydrothermal field (Mid-Atlantic Ridge, 26°08’N) is located 2.4 km east of the main rift valley at depths of about 3650 m. The TAG hydrothermal field stands as an example of a massive sulfide deposit on a low-spreading ridge associated with an active detachment fault (Humphris et al., 2015).Associated with the TAG hydrothermal field are metalliferous sediments/deposits that can reach several meters in thickness (Metz et al., 1988). These sediments/deposits can be formed by hydrothermal plume fall out (buoyant and nonbuoyant), massive wasting of hydrothermal edifices (sulfide mounds and chimneys) and authigenic mineralization (Gurvich, 2006). By studying the grain size, mineralogy and geochemistry of the sediments it is possible to identify each type of sediments and find their proximal or distal distribution. The location of the gravity cores studied here is distributed along sediment ponds near inactive hydrothermal mounds [Three mounds area (B)], actively venting low-temperature mound [Shimmering mound (A) and Mir zone (D)] and an area of sediment accumulation [Central area (C)] (Rona et al., 1993) (see Figure 1)."

Jan 1, 2017

By G. E. Rennick

A siphon-type laboratory airlift well was designed and operated by the Bureau of Mines to study the concurrent flow of air and oil in small-diameter pipes. Glass flowstrings ranging from to 1 inch in diameter and from L4 to 64 feet in length were equipped with various-sized air-entry jets. Liquid holdup volumes were used to correlate vertical-flow data, and empirical equations for computing gas requirements were derived. The effect of flowstring length on flow efficiency, oil- and gas-production rates, and gas-oil ratios was relatively minor at all flowing pressures. Air-jet sizes and flowstring diameters, however, had a marked effect on flow efficiency and oil- and gas-production rates. The small air-jet sizes increased maximum flow efficiency and its associated flowing pressures but decreased oil- and gas-production rates. An increase in flowstring diameter increased maximum flow efficiency with its associated flowing pressure and also increased the production rates of oil and gas.

Jan 1, 1965

By Carolina Achiardi, Álvaro Videla

"In pre-design a flotation plant copper sulphide is useful to apply a simplified methodology of economic analysis that relates the size of the grinding product (P80) with the throughput and their effect on copper grade and recovery of copper in the final concentrate of a flotation circuit. With this objective, this methodology covers the following analysis, perform by evaluation of five sizes grinding products (P80): 120, 150, 180, 210 and 240 microns. • Flotation testwork (rougher kinetic, open and closed cycle test) has been done to determinate the optimum operating conditions and to estimate the maximum recovery of copper in the concentrate, using Klimpell model • Flotation plant design assumes a conminution product obtained from a 25,000 HP ball mill • Metallurgical balance is performed for the circuit based on flotation testwork • Economic evaluation by calculating capital expenditure (CAPEX), operating expenditure (OPEX) and sales revenue, to determine NPV is performed and relates to the NPV. An optimum grind size is then obtained based on cash flow. The results shows a grinding product size that optimizes flotation and allows a higher NPV and INPV over other alternatives."

Jan 1, 2014

By AIME AIME

COMMITTEES. LOCAL COMMITTEE.-R. C. Crawford, Chairman; Harrison W. Craver, Secretary; Julian Kennedy, Taylor Allderdice, E. W. Pargny, Charles L. Miller, W. H. Rea, S. A. Taylor, M. E. Wadsworth, W. A1. Henderon, A. C. Dinkey, W. L, Jones, E. L. Messier, George Mesta, J. H. Jones, G. H. Riley, S. L. Goodale, D. W. McNaugher. LADIES' COMMITTEE.-Mrs. Taylor Allderdice, Chairman; Mrs. Julian Kennedy, Mr. R. C. Crawford, Mrs. E. M. Herr, Airs. J. W. Marsh, Mrs. T. A. Mellon, Mrs. G. P. Bassett, Jr., Airs. D. M. Clemson, Mr. Charles L. Miller, Mrs. V. L. P. Shriver, Mrs. W. A. Bostwick, Mrs. A. C.. Dinkey, Mrs. A. P. Childs, Jr., Mrs. F. R. Dravo, Mrs. Hermon Griffin, Mrs. W. H. Nimmick, Mrs. F. H. Lloyd, Mrs. S. G. Cooper, Mrs. S. H. Church, Mrs. E. N. Ohl. The Institute Headquarters was established at the Hotel Schenley, where a bureau of information was maintained for the benefit of the visiting members and guests of the party during the time of the meeting. TECHNICAL SESSIONS. The first session, held Tuesday evening, Mar. 1, 1910, in the large lecture hall of the Carnegie Institute, was called to order by Harrison W. Craver, Secretary of the Local Committee. Mr. Craver introduced Dr. John A. Brashear, who extended to the members and guests a hearty welcome to the city of Pittsburg, to which David W. Brunton, President of the Institute, responded. The following papers were presented in oral abstract by the authors: Biographical Notice of Dr. Charles B. Dudley, by Dr. R. W. Raymond, New York, N. Y. Biographical Notice of William Metcalf, by Col. II. P. Bope, Pittsburg, Pa. Development of Hindered Settling Apparatus, by Robert H. Richards, Boston, Mass. (Illustrated by lantern-slides.) A Few Mining Districts in the Argentine Republic, South America, by Z. S. Beyl. (Illustrated by lantern. slides.)

Apr 1, 1910

By R. J. Grace

Primarily, this report presents the results of direct efficiency tests which were conducted on the 200-hp Coal-Pak, hot water generator, No. 3 unit, at the Otterbein College heating plant in Westerville, Ohio. Also included is a comparison of fuel consumption of the old plant with that of the new plant, a comparison of efficiencies obtained by the indirect method on a 300-hp Coal-Pak, and the cost of gas vs coal to show the savings effected by the latter. Tests were conducted with ten different bituminous coals selected by Mr. S. A. Frye, business manager at Otterbein College. The coals, of ten different suppliers, from four different states and eight different seams, represented competitive products economically available to the heating plant. The tests were conducted by the usual plant personnel which consisted of the plant superintendent and two assistants. No outside supervision of any kind was given during the tests. The tests were made from Dec. 13, 1960, through Mar. 4, 1961. During this period, 300.03 tons of coal were burned in the No. 3 unit. This tonnage represented 25.5 pct of the total coal requirements for the 1960-61 season. Operation during the period was typical for the comfort heat load which the plant supplies. Of the 78.12 operating days, the boiler was on bank or holdfire 14.31 days, or 18.32 pct of the time. The length of individual bank periods varied from 2 hr to three days. Run periods varied from about 2 hr to 7.26 days. DESCRIPTION OF PLANT AND HEATING EQUIPMENT The new Otterbein College heating plant, completed in the summer of 1957, is located in a 50x50 ft boiler room on the lower level of a combination maintenance building and heating plant. The building is situated away from the main college complex on a hillside, which allows for on-grade delivery of coal to the three bunkers. The bunkers are located directly above and connected with the individual stoker hoppers. Fig. 1 shows an in-plant view of the Coal-Pak installations. The No. 3, hot water, unit is at the extreme right. The heating plant consists of three 200-hp Coal-Paks. Unit No. 1 is operated at 4 psi and carries the steam heating load of 14,000 sq ft EDR (steam) installed in two buildings. Provision has been made to convert these buildings to hot water heating at some future date. Since the unit under maximum load conditions operates only at about 50 pct steaming capacity, a heat exchanger was installed to convert steam to hot water. Thus, during light loads in mild weather the unit can be employed more efficiently to supply both steam and hot water and eliminate the need for operating either of the two hot water units. Units No. 2 and No. 3 serve a hot water heating circuit for a large group of 21 buildings of the College. These boilers deliver hot water at temperatures up to 200°F to the various buildings through distribution lines in tunnels. The present calculated maximum heat loss for the College is 15.195 million Btu per hr. The Coal-Paks have individual design capacities of 6.700 million Btu per hr. Table I lists the design date for the Coal-Paks. The No. 3 unit was chosen as the test unit, primarily because of its bunker location at the far end of the delivery ramp. Because of this location any excess coal above bunker capacity could be conveniently stored and isolated from the adjoining bunkers. Each bunker has a capacity of 30 to 35 tons of coal. The No. 2 and No. 3 hot water generators are equipped with automatic controls which maintain the

Jan 1, 1962

By Don S. McCoy, W. Joseoph Schlitt, Donald B. Swanson

Introduction Contract miners have long played an important role in the minerals industry. Historically, these were individuals who worked as underground miners and were paid by the ton or the day. These hardy types still ply their trade in the industry. However, today the term "contract miner" generally refers to an independent contractor who supplies the people and equipment needed to conduct surface mining operations for an owner or lease holder. At one time or another, contract miners have probably handled virtually every mineral commodity from sand, gravel, and stone quarrying to mining uranium and precious metal ores. In today's market, though, contract mining appears to be concentrated in just a few areas. These include precious metal mines, various types of solid fuel operations, and reclamation of abandoned mines. This article covers market areas, contract mining organizations and services, and their advantages and disadvantages. Gold and silver mines Although exact figures are not available, contract mining represents a significant element in the precious metals segment of US mining. For gold and silver, the growth in contract mining appears to be paralleling the overall growth in mining activity itself. Nevada is the largest gold producer in the US. The latest available data (1985) on open pit gold/ silver mines is summarized in Table 1 (Nevada Division of Mine Inspection, 1986). These figures show that contractors mined nearly half of the ore in 1985. Total tonnage mined, of course, was much higher for both owners and contractors. Assuming an average stripping ratio of 2.5 to 1, the total daily mining rate for the contractors was more than 145 kt (160,000 st). This translates into a contract mining rate of about 36 Mt/a (40 million stpy) based on 250 operating days a year. This approximate 50-50 split between owner and contractor also appears to hold for the future. Metals Economics Group (1987) published a survey of 51 short-range and 54 long-range precious metal projects being considered throughout the US. Contacts with owners, requests for proposals, and past corporate philosophies all suggest that about 55% of the open pit operations scheduled for the near-term will be contract mined. This represents an estimated 26 kt/d (28.5 stpd) of ore. A similar assessment of the long-term projects indicates that about 12 kt/d (13,000 stpd), or just under 50% of the open pit tonnage will be contract mined. There are several reasons why gold and silver represent such an attractive opportunity for contract miners. One is that many properties are owned or leased by individuals or partnerships that are essentially entrepreneurial. Many of these owners or lease holders have minimal mining experience. They must therefore either rely on contract mining organizations, sell an interest in the project to another mining company that will become the operator, or hire an experienced management team. For about the same reason, junior mining companies also represent good prospects for contracting. This is because most junior companies focus on exploration. They would prefer to spend their generally limited funds on drilling programs. Thus, they tend to minimize additional capital requirements by using contract miners or farming out the property to another mining company. Another good reason for using contract miners is the size of the deposit. Many properties are so small that their reserves will not support the capital expenditures

Jan 12, 1987

By Peter Cali, David Yang, Rakshya Shrestha, Ralph Griffin, Sushil Kafle, Motoji Watanabe

"As a part of the ‘US Army Corps of Engineers - New Orleans to Venice, Louisiana Hurricane Protection Project’, deep mixing ground improvement was designed to reinforce the soft ground for the emergency repair of a 7297-foot (2,224 m) long section of the Mississippi River West Bank levee in Buras Levee District under the contract ‘Emergency Levee Repairs – Contract P-17A, Plaquemines Parish, Louisiana’. The ground improvement utilizes a series of soil-cement panels oriented in the transverse direction of the levee. Each soil-cement panel consists of overlapped columns installed by the deep mixing method (DMM). DMM is performed from the existing top of levee and extended to competent stratum at depths varying from 26 feet (7.9 m) to 47 feet (14.3 m) below the working platform. The main purpose of the DMM panels is to provide shear resistance to the levee against lateral loads at high flood water condition. The DMM panels support the additional vertical load created by the increased height of the levee for hurricane protection and control the long-term settlement of the remediated levee, which might otherwise occur due to the compression of soft foundation soils. The total amount of DMM treatment is approximately 54,000 cubic yards (41,500 cubic meters). This paper reviews the project background, subsurface soil conditions and presents the design, construction and quality control of the deep mixing work for this levee repair project.INTRODUCTIONIn August 2005, Hurricane Katrina caused major damage to the federal and non-federal flood control projects in southeast Louisiana. In September 2005, Hurricane Rita caused further damage to this flood protection system. In an effort to minimize further damage and to enhance the protection level, certain non-federal levees located on the west bank of the Mississippi river in Plaquemines Parish, Louisiana were incorporated into the New Orleans to Venice Hurricane Protection project by the US Army Corps of Engineers (USACE). DMM techniques were then utilized by the USACE in several of these emergency projects and in many medium sized projects between 2006 and 2010 (Bruce et al. 2012). This project (called P-17A hereafter) which forms a part of this protection system is executed under the contract Emergency Levee Repairs, Contract P-17A. This is one of the largest projects of the USACE after LPV- 111. The huge LPV-111 project (5.5 miles (8.9 km) of levee) treated by DMM), the largest DMM project ever conducted outside Japan, began in 2010 and was completed in 2012 (Cali et al. 2012, Bruce et al. 2012).DMM is utilized at P-17A to construct a series of soil-cement panels oriented in the transverse direction of the existing levee to provide shear resistance to the levee against the lateral load during high flood water condition. The project extends 7297-foot (2,224 m) along the West Bank Levee of the Mississippi River in the Buras district of Louisiana, parallel to Louisiana Highway 23 (LA-23) as shown in Figure 1. It is located approximately 60 miles (96 km) south-east of downtown New Orleans."

Jan 1, 2015

By Charles B. Kenahan

The chemical and galvanic corrosion properties of commercially pure titanium, a titanium-manganese alloy, a titanium-aluminum-tin alloy, 2 titanium-copper alloys, and 2 titanium-aluminum alloys were studied. Tests were conducted in synthetic ocean water, tapwater, 1-percent sodium hydroxide, 5-percent ferric chloride, and 5-percent ammonium sulfate solutions, various concentrations of sulfuric and hydro-chloric acids, and 1-percent oxalic acid. All tests were carried out at 35° or 60° C., with a flow of air or helium. Dissimilar metals involved in galvanic-couples experiments were aluminum, magnesium, and stainless steel. Electrode-potential measurements on single (uncoupled) specimens were also measured. Commercially pure titanium and the titanium-base alloys were generally quite similar in their resistance to corrosion, whether alone or in. contact with a dissimilar metal. Contact with the titanium metals caused aluminum and magnesium to corrode by galvanic action. Coupling with stainless steel in sulfuric acid solutions decreased the normal corrosion rate of the titanium metals. Contact with aluminum in air-aerated oxalic acid and 0.1N sulfuric acid solutions resulted in cathodic corrosion of titanium and the titanium-base alloys; however, there were a few exceptions. The copper alloys showed better corrosion characteristics than titanium in certain sulfuric acid solutions but were more susceptible to chemical attack in some hydrochloric acid solutions.

Jan 1, 1958

By A. T. Aldred

HIS note reports the existence of several new scandium intermetallic compounds of the A2B and AB stoichiometries where the A element is scandium and the B element is from group VIII or IB of the periodic table. Alloys were arc melted from high-purity materials using 2-g charges. Typical lot analyses for the materials used are given in Table I. Chemical analyses were not made on the alloys as there was little or no weight loss on melting. After annealing the specimens for 14 days at 600°C and water quenching, metallographic and X-ray techniques were used to determine the occurrence of the various phases. Table II lists the relevant data for the A2B phases. The Ti2Ni structure,2 has ninety-six atoms in a face-centered cubic cell, whilst the Al2Cu structure is body-centered tetragonal.3 The occurrence and stability of the various A2B type phases in transition metal alloy systems has recently been considered by Nevitt.4 The Ti2 Ni-type phases occur over a relatively narrow range of radius ratios from 1.14-1.26 with A partners mainly from the titanium group and B partners predominantly from the cobalt group. This suggests that both a favorable relative size of the atoms and a favorable electron concentration are necessary conditions for the existence of this structure type. A12Cu-type phases involving Ti group elements as A partners are formed with B partners from the cobalt and nickel groups, predominantly the latter, and have a higher radius ratio (1.27-1.29) than the Ti2Ni-type phases. The occurrence of both Sc2Pd and Sc2Co can be rationalized in terms of these considerations but SczNi is anomalous because of its high radius ratio. The radius ratios are calculated from the Goldschmidt CN12 radii and it might be postulated that the scandium atom is reduced in size in the compound. Unfortunately the apparent atomic size of scandium in the compound cannot be calculated from a knowledge of the structure and lattice dimensions. The crystal structure of the Sc2Au compound could not be identified from the powder diffraction pattern, but it is interesting to note that it is isostructural with a series of A2Au phases where A is a rare earth element.5 Pertinent information concerning the AB phases found in this investigation is given in Table 111. All the phases have the ordered bcc CsCl structure. The occurrence and stability of CsC1-type phases have recently been discussed by Dwight,7 who concludes that the relative sizes of the atoms do not appear to be a controlling factor (the radius ratios of all the known phases vary between 0.985 and 1.439) but that the occurrence of the phase is predominantly dependent on the relative positions of the A and B partners in the periodic table. As a measure of the stability of a given CsCl phase it has been customary to define a lattice contraction on alloying, DAB — dAB, where DAB is the AB interatomic spacing as calculated from the Goldschmidt CN8 radii and dAB is the AB interatomic distance in the compound (= v3/2 ao). Values of the parameter DscB -dScB for the compounds found in this investigation are given in column 5 of Table III and are plotted in Fig. 1. If the lattice contraction does in fact reflect the stability of a phase then Fig. 1 shows two very interesting trends when the position of the B partner in the periodic table is considered. Firstly, within any given group the stability increases in order on going from the first to third long period. Secondly, within any given period the stability increases in order as the B partner is taken from the copper, nickel, and cobalt groups. In the case of the second long period, the lattice contraction reaches a maximum at ScRh and then decreases to ScRu. The same trend could

Jan 1, 1962

By Jean P. Murray

The extractive metallurgical industry is a major consumer of high-temperature process heat. It is, consequently, a major contributor of C02 emissions and other greenhouse gases derived from the combustion of fossil fuels for heat and electricity generation, or from the process itself. These emissions can be substantially reduced by replacing fossil fuels with solar energy as the source of process heat. Concentrated solar radiation can supply a large amount of thermal energy to endothermic reactions at temperatures exceeding 2000 K. Another potential for C02 emission reduction exists as well. Thermodynamically, the total amount of energy required for metal production at any given temperature is the enthalpy change of the reaction, AH. Of this total, an amount equal to the Gibbs free energy AG must be supplied as high quality energy in the form of electrical work or by using a reducing agent. Many conventional metallurgical electrolytic processes and reducing furnaces are unable to add only the thermodynamically required electricaVchemical energy due to process inefficiencies. Electrical work in excess of AG for the reaction may be used to supply process heat, or excess chemical reducing agent burned to supply process heat. Besides the thermodynamic and economic drawback, this inefficiency results in excess C02 emissions derived from the combustion of fossil fuels. With these practices, energy becomes an important cost element of the metal produced. Highly-concentrated solar energy has the demonstrated potential to separate these two forms of process energies, allowing the addition of only the thermodynamically required high-quality energy while high-temperature solar process heat is provided to the reactor. Examples of metal oxide reduction processes that have been studied experimentally in solar furnaces include the production of Fe, Al, AI/Si alloy, Mg, Zn, Tic, Sic, CaC2, TIN, Si3N4, and AIN by carbothermic reduction of their oxides, and the high-temperature solar- electrothermal reduction of ZnO.

Jan 1, 1999

By O. W. Simmons, L. W. Eastwood, C. M. Craighead

Binary alloys of titanium with silver, lead, tin, nickel, copper, beryllium, boron, silicon, chromium, molybdenum, manganese, vanadium, iron, and cobalt were studied. One-half-pound ingots of the alloys were prepared in an arc furnace, employing a water-cooled copper crucible, an argon atmosphere, and a water-cooled tungsten electrode. The half-pound ingots were fabricated by forging at 1700°F in air to 1/4-in. slab, followed by hot rolling at 1450°F to 0.060-in. sheet. Tensile properties, minimum bend radii, hardnesses, response to heat treatment and aging treatment, and phase relationships were determined for these alloys. EARLY in 1947, as one phase of the evaluation of materials for Air Force Project RAND, Battelle successfully arc melted Bureau of Mines titanium and obtained a few basic properties of unalloyed titanium and several titanium-base alloys. The low density, excellent resistance to corrosion, and high tensile properties of these materials stimulated a great deal of interest among metallurgists and designers seeking better materials of construction. As a result of this early work, Battelle, under contract with the Air Materiel Command, Wright-Patterson Air Force Base, has continued extensive studies of titanium alloys. The result of the first year's work is described in three papers. The present one deals with binary alloys, the second with ternary alloys, and the third paper with quaternary alloys. Melting Method The arc furnace for melting titanium and other refractory metals was developed at Battelle Institute under Air Force Project RAND. During the present work, considerable improvement has been made in the design and operation of this furnace for making small ingots of titanium and titanium alloys. The improved furnace design is illustrated by fig. 1, which shows a cross-sectional view with the dimensions of various parts and their relationship to one another. The essential features of the furnace are the inert argon atmosphere, the water-cooled copper crucible, and the water-cooled tungsten elec- trode. The melting crucible is formed by spinning 0.064-in. annealed copper sheet. A groove for an O-ring seal is spun into the flange of the crucible. A fiber gasket between the crucible flange and the water-cooled brass cover provides electrical insulation. The brass cover is fitted with a sight glass, an opening for the water-cooled electrode, and a tube through which the material is charged. Direct current is used for melting, with the positive electrical terminal connected to the water jacket and the negative terminal to the electrode. A typical heat is made in the new furnace as follows: Approximately, 0.20 lb of titanium under 2 mesh per in. and the alloy addition, if any is to be used, are placed in the bottom of the crucible. The cover is clamped on the crucible so that the O-rings make a tight seal. The remainder of the charge is placed in a glass bottle, and this is connected by a rubber hose to the charging tube of the furnace. The crucible and the charge are evacuated with a mechanical pump to a pressure below 50 mm of mercury, or less if desired, and held at this reduced pressure for 5 min. Hot water is circulated through the jacket during the evacuation period to assist in outgassing the crucible. Following the evacuation, tank argon of 99.92+ pct purity is flushed through the crucible for 5 min and then adjusted to give a positive pressure of 1 to 2 psi. During subsequent operation, an argon pressure regulator introduces only enough argon to compensate for the small leakage which occurs through a mercury trap. Re-evacuation and reflushing the melting chamber with argon produced a negligible reduction in contamination. Two 550-amp generators, connected in parallel, constitute the power source. Normally, about 800 amp are used for melting alloys which are not extremely refractory. The generators are set for 90 v open circuit and 200 amp. The arc is struck and the electrode is withdrawn to produce an arc of 23 to 26 v. This voltage is maintained while the electrode tip is moved slowly around a circle 1 in. smaller than the diameter of the ingot. The current

Jan 1, 1951

Abouleid,A.F.327 Ajdic, I.451 Albert, L.F. 221,307 Alem, A. 525 Asama, T. 431 Avgherinos, P. J. 373 Baker Jr, C.N. 331 Balossi Restelli, A. 171 Beckwith,G.H.33 Bellini, A. 469 Bermingham, P. 481 Bertero,M. 171,535 Beyer, M. 503 Bottiau, M.41 Botto, G. 205 Bruskeland, O. 337 Budhu, M. 349 Bullivant, R.A. 181 Bustamante, M.G. 185, 475, 535

Jan 1, 1994

By Alejandro Freyre Belliki

El presente trabajo describe las causas del abandono de mineral, las razones de su recuperación de las antiguas explotaciones, así como el estudio de las técnicas más adecuadas para realizar dicha operación. Entre las causas del abandono de mineral tenemos: la variación de precios de metales, calidad y complejidad del mineral, seguridad, derrumbes y el método de explotación. También se mencionan los casos de recuperación de materiales útiles como rellenos antiguos, mineral "in situ", derrumbes y minerales dejados en mantos. Finalmente, se evalúa la posibilidad de hallar reservas recuperables y por ende, examinar la factibilidad de su recuperación. Asi también, las zonas recuperables son generalmente las más altas, ya que la aplicación de métodós con relleno hidráulico tomado directamente de los relaves de la Planta, debe considerar la diferencia de nivel entre ella y los tajeos re-explotables.

Aug 20, 1965

By Robert M. Becker

The evaluation of existing rock stresses from measurements of the change in diameter of a borehole that is stress relieved by overcoring has become a common practice, The stress evaluation has been generally based ,on isotropic relations whereas rock in varying degrees is anisotropic. A plane elasticity solution for a circular hole along one axis of an orthotropic medium is adapted for evaluation of in situ stress from borehole deformation measurements. A consistent triaxial testing approach for determining elastic properties of the overcore is also outlined. Comparison of isotropic with anisotropic interpretations of the state of stress show that large errors in evaluation may occur if the anisotropic character of the rock is neglected.

Jan 1, 1967

By Gordon E. Forward

Sustainable development in the mining and metallurgical industrial sectors can be implemented by two distinct groups: 1. The managers and operators of existing facilities ? working with industrial processes and systems that have been operating for many years, in some cases several decades 2. Those who are engineering and building new facilities or upgrades of existing plants. The focus of this paper is to look at approaches and tools available to bring sustainable development thinking and action to these two distinct areas. The paper will discuss: 1. The concept of working with existing facilities and processes to minimize waste and maximize product from a regional raw material/resource consumption perspective using By-product Synergy. The BPS concept has been carried out in Texas (TXI-Chaparral?s STAR Example), in Calgary, Alberta with Suncor assistance, in Montreal, Quebec with Noranda as project leader and currently in Ontario with Dofasco as industry champion. Successful outcomes are presented. 2. A new approach to designing greenfield facilities or upgrades to existing plants using a sustainable development review process. Undertaking the review early in every project can institutionalize the concept of discussing and implementing waste elimination, design for environment, eco-efficiency, risk minimization and community awareness and involvement in the future capital stock. The review process has been developed as an excellent tool for measuring and reporting on a corporation?s sustainable development progress. Examples of how this approach has been adopted in the mining and metallurgical industry, and the resultant outcomes are given.

May 1, 2002

By Donald O. Capstick

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Jan 1, 1983

By Victor L. Stevens

The Oso tunnel is one of three tunnels located on the San Juan-Chama Project in south-central Colorado. The purpose of the tunnel is to carry water from the upper San Juan watershed through the Continental Divide into the Rio Grande watershed. The Oso tunnel contract was awarded by the U.S. Bureau of Reclamation to a joint venture of Boyles Brothers Drilling Co., Gibbons & Reed Co., and Cimco in February 1966. The Oso tunnel is 26,660 ft long and has a rock section of 10 ft, 2 in. with a lined diameter of 8 ft circular. The main portion of the rock formation was in the Lewis shale. This formation is quite consistent in hardness and texture with exception of localized limestone concretions and sandy siltstone inclusions. The job was bid on the basis of using a James S. Robbins mole of the P221 series. The decision for using a mole was predicated on the geology and the experience of the same joint venture on the adjoining Azotea tunnel where a mole was successfully used. The Robbins P221 series mole is capable of cutting from 9 ft, 11 in. to a 10 ft, 7 in. diam without any major change. It is 40 ft long and weighs 60 tons. The cutting head has 22 disk-type cutters and one center pilot (standard 9 7/8 -in.) tricone bit. The head is powered by four 75 hp electric motors. Gripper pads at about the midway point on each side of the machine are pushed into the tunnel ribs and thrust-cylinders in front of each gripper apply pressure to the face as the head rotates. The regripping cycle is 3 ft and then the grippers are retracted and advanced. Each cutter on the perimeter of the head is followed by a bucket which scoops up the cut material, almost in the same manner as a water wheel, and dumps it into a conveyor situated along the top of the mole. The muck from the mole conveyor is dumped into a gantry-type conveyor. The gantry conveyor is 350 ft long and is pulled

Jan 1, 1970

By John Durkin

Electromagnetic narrow-band signals were transmitted through the earth at 27 coal mines located throughout the United States. From those Bureau of Mines tests, apparent earth conductivity values were derived based upon a homogeneous half-space model of the earth. The derived conductivity values were found to be inversely proportional to the transmitted frequency and mine depth. A linear regression model relating the logarithm of the conductivity to the mine depth was formulated, and the results indicate that the mine depth can be an adequate predictor of the apparent earth conductivity above coal mines. Apparent earth conductivity was found to decrease with mine depth.

Jan 1, 1984

By John Durkin

Electromagnetic narrow-band signals were transmitted through the earth at 27 coal mines located throughout the United States. From those Bureau of Mines tests, apparent earth conductivity values were derived based upon a homogeneous half-space model of the earth. The derived conductivity values were found to be inversely proportional to the transmitted frequency and mine depth. A linear regression model relating the logarithm of the conductivity to the mine depth was formulated, and the results indicate that the mine depth can be an adequate predictor of the apparent earth conductivity above coal mines. Apparent earth conductivity was found to decrease with mine depth.

Jan 1, 1984