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By W. E. Rudolph, R. E. Baylor

DUE to its location in the Atacama Desert, one of the most barren of the earth's surfaces, Chuquicamata's water supply presents unusual problems. Yearly rain-fall averages less than one tenth of an inch at the plant. However, there are summer showers above 12,000 ft in the Cordillera to the east, the resulting run-off flowing through old river valleys buried beneath more recent volcanic formations, to be impounded within sediment-filled basins. This water emerges at springs where the outlets of these basins are blocked by lava flows, and here are formed the small streams which feed the only important river of the region, the Rio Loa. Chuquicamata's water is obtained from these springs and rivulets. [ ] The map above indicates four pipe lines from which potable and industrial water are supplied. Potable water, amounting to 4500 metric tons per day, is conveyed in the Toconce pipe line from springs 59 miles due east of Chuquicamata. This water is used not only for drinking, but also for boilers and other needs requiring high quality. For industrial water at the oxide plant, there are two 12-in. pipe lines from the Rio San Pedro, carrying a total of 17,000 metric tons per day of slightly brackish water. This water is at present used mainly for leaching and for hygienic purposes. Water Source Found For the present and future needs of the sulphide plant, it was calculated that at least 32,000 metric tons per day of make-up water would be required. For this purpose, a pipe line of 44 miles length was constructed to bring in the entire flow of the Arroyo Salado, one of the eastern tributaries of the Loa. The salt content of this water is so high (over 5000 parts per million of solubles, mostly chlorides) that it is highly detrimental to farming, and the Chilean Government had been studying projects to separate these waters from others of the Loa system in order to improve agricultural conditions in the fertile valley of Calama. So it happened that the Government was willing to award rights to the Arroyo Salado waters under agreement whereby the Mining Company removes waters from the Rio Loa system above Calama for all time. The outlet of these waters, after serving their purpose at the new concentrator and leaving the plant in tailing, is the Salar de Talabre, an old salt lake which presents fully ten square miles of surface to serve as an evaporating pan, the outlets having now been blocked by dams. Here the dry climate of Chuquicamata is a favourable factor, evaporation averaging slightly above 1/4 in. per day. The Toconce and San Pedro pipe lines have been functioning from 26 to 34 years, and through the use of special cleaning tools which were developed at the plant, as well as deaeration of the more active Toconce water, these pipes are now maintained at capacities which do not diminish as years go on. Constructing the Dam The Arroyo Salado pipe line design and construction involved certain special and interesting features, and inasmuch as this line and its intake works are solely for the needs of the new sulphide plant, more detailed description is given. The waters are impounded at a gravity dam constructed of concrete to a height of 100 ft above the river bed, keyed into the precipitous Dacite walls of the narrow canyon (barely 6-ft wide at the bottom, only 25-ft width at 50 ft above). A small secondary dam was built 100 ft down stream from the main dam, providing a pool of 15-ft depth to protect the main structure from flood flows over the spill-way during the rainy season. A system of four 36-in. syphons was designed for discharging these flood waters from the lower depths of the lake, in order to avoid eventual sedimentation behind the dam. The lake has a length of 3300 ft, and its water level is controlled by an adjustable spillway permitting draw-down of eighty inches, amounting to 41,000 metric tons of available capacity. This regulation is necessary because of wide fluctuations in stream flow between day and night due to freezing of feeders. During the construction of the dam the entire river flow was handled within a 36-in. pipe line some 2000 ft in length. As the excavations proceeded

Jan 1, 1952

By J. D. Anthony

The Australian continent possesses significant reserves of a wide range of minerals, including bauxite, coal, copper, diamonds, gold, iron ore, lead, manganese, mineral sands, nickel, phosphate, silver, tin, uranium, and zinc. Australia's identified economic resources of many minerals are very large as indicated in Table 1. A sophisticated and highly experienced mineral industry is now an established feature of the Australian economy and Australia is the world's largest exporter of iron ore, alumina, mineral sands and refined lead and amongst the leading suppliers of many other commodities such as coal, lead and zinc ores/concentrates, nickel, refined zinc, tungsten concentrates and bauxite. The industry exports 70% of its production. This is reflected in the value of Australian mineral exports which have grown from about $200m in 1960/61, comprising 10% of total export receipts, to about $1265m or 29% of export income in 1970/71 to around $7061 representing 37% of Australia's total export income in 1980/81. Details of the more significant minerals are as follows: Japan (42.1%) USA (11.3%) ASEAN (6.3%) UK (5.9%) F.R. Germany (3.8%) Republic of Korea (3.4%) New Zealand (2.6%) Also see Table 2. AUSTRALIA'S MINERAL RESOURCES POLICIES Federal and State Governments' Responsibilities Australia has a federal system of government comprising six States, a self-governing Territory and a Federal Government. Under the Australian federal system the Constitution sets down the powers of the Federal Government. All powers not assigned to the Federal Government in the Australian Constitution reside automatically with the States. Certain of these broad powers result in the Federal Government having a significant influence on resources development. For example, in being responsible for economic management, the Federal Government's fiscal and monetary policies have an important effect on industry as well as on State finances. In particular, the taxation regime employed by the Federal Government is of direct importance to decision-makers in the resources industry. The Federal Government is responsible also under the Constitution for external trade matters; and international trade and commodity matters are increasingly important in Australia's international relationships. Foreign investment is another area where the Federal Government has a role to ensure that national interests are protected. This foreign investment power flows from the Federal Government's control of foreign exchange movements into and out of Australia. However, before enlarging on these and others of the Federal Government's powers and policies, it should be emphasized that the State governments, by virtue of their wide powers to regulate matters within their own boundaries, are more directly involved in the day-to-day administration and regulation of mining operations. For instance, the powers of the State governments include the responsibility-for the granting of exploration rights and mining leases, the approval of mining operations and the levying of royalties and other like charges. Administrative arrangements covering the granting of minerals and petroleum exploration and development titles vary from State to State. Before development rights are granted, State governments consider environment protection and rehabilitation aspects of development proposals. The provision of infrastructure within State borders is a matter primarily of State government responsibility. It is usual practice in Australia for State governments to construct and operate infrastructure services such. as railways, ports and electricity generation and transmission. The States may also provide certain public services such as electricity. and water, port and loading facilities, communications, health and education services which form part of the infrastructure of mining operations. In remote areas the mining companies themselves usually are expected to provide much of this infrastructure. However, the Federal Government is primarily responsible in some fields, such as telecommunications and parts of the railways network. State governments carry out preliminary exploration and geological mapping and some are directly involved in the mining of coal for power generation. The Federal Government's responsibilities in addition to economic management, taxation, international relations, foreign capital and investment, include regulation of exports, environmental matters and matters affecting the Aboriginals of the Northern Territory. FEDERAL GOVERNMENT POLICIES The continued sound development of the minerals and energy resources sector is regarded by the Federal Government as being of very great importance. However, the Government does not seek to participate directly in resource developments. It sees its role rather as that of establishing a sound economic and policy climate in which private companies can identify opportunities, seek out customers and marshall the necessary capital for the development of resource projects.

Jan 1, 1982

By K. E. Nelson

The effect of thorium and zinc variations on the strength and 100-hr creep characteristics of Mg-Th-Zn-Zr alloys was investigated. Optimum resistance to creep at 650° and 700°F are attainable within a certain range of thorium and zinc contents. This range does not conform to that which develops maximum tensile properties. RAPID advances have been made during the last few years in the development of magnesium alloys for elevated temperature applications demanding high resistance to creep. The beneficial effect of rare-earth metals on the creep resistance of magnesium alloys has been emphasized by a number of publications1-13 and such alloys are now in commercial production. The use of thorium as an alloying ingredient in magnesium was mentioned by McDonald and also in two Alien Property Custodian patent applications.10,17 The initial observation of Sauerwald that thorium contributes still higher creep resistance to magnesium than is attainable with rare-earth metals has recently been substantiated.'" " In fact, it has been demonstrated that the useful temperature range of magnesium alloys is appreciably extended by the use of thorium. In all cases, it was observed that zirconium must be included in the alloys in order to render them fine grained and more readily castable. Several recent publications:2-27 indicate that a still further improvement in creep resistance and a further extension of the useful temperature range can be realized by the addition of zinc to alloys. The primary purpose of this paper is to present the results of a comprehensive study of the effect of zinc on the strength and creep characteristics of Mg-Th-Zr alloys. Compositions covering the range of thorium content from M to 6 pct and zinc content from 0 to 5 pct have been investigated. The creep characteristics at 650" and 700°F reported in this paper are based on results of tests of 100-hr duration. It is appreciated that creep tests of 100-hr duration might not yield adequate data for design purposes for parts with much longer expected life. However, for the purposes of the present discussion, it is felt that the combination of stresses and temperatures used in the 100-hr creep tests have yielded a clear representation of the compositional variation of creep resistance at the temperatures investigated. Creep tests of 1000-hr duration are now in progress on a few of the most promising alloys. Preparation and Testing of Alloys The alloys studied in this intensive investigation were prepared in 25 lb capacity mild steel crucibles. The thorium, zinc, and zirconium were alloyed and poured as described in earlier publications The thorium was introduced into the melt in the form of a Mg-Th hardener," the zinc added in the metallic form, and the zirconium alloyed in the form of the commercial hardener containing magnesium and 30 to 50 pct Zr.10, 26 Fluxing practices for melting and refining were the same as for magnesium-rare-earth metal-zirconium alloys. The melts were sampled for analytical determinations and poured into separately cast 1/2 in. diameter standard tensile bars. The test bars were given a precipitation treatment of 16 hr at 600°F in laboratory furnaces. It has been shown by other tests that a high temperature solution treatment followed by an aging treatment is unnecessary for the development of optimum properties in Mg-Th-Zn-Zr alloys. The selection of 600°F as the aging temperature was based on an attempt to achieve metallurgical stability without coalescence of the undissolved phases and the attendant loss in strength. The thorium, zinc, and zirconium contents of each melt were determined chemically. The zirconium contents are reported in two parts, "soluble" and "insoluble," referring, respectively, to the portions present in the alloy which are soluble and insoluble in dilute HCl acid. Distinction is being made between these two components of the zirconium content in the alloys because it has been found that only that portion of the zirconium content which is soluble in dilute inorganic acids affects the structure and properties of the alloys. The usual impurities consisting of copper, iron, manganese, and nickel were determined spectroscopically. The analysis for each melt is listed in Table I. A description of the methods of tension and creep testing has been detailed in earlier papers. The tests were performed with the cast skin re-

Jan 1, 1954

By R. M. Hudson, P. E. Perry

Weight-gain data obtained by nitriding low-carbon sheet steel in an amrnonia CNH,) atmosphere indicated that the process obeyed a parabolic rate law. The calculated actization energy for nitriding in the range 964" to 1268°F agreed reasonably well with published data. At 1358"F, rate data indicated that the activation energy decreased. Weight-gain data obtained by uszng mixtures of NH3 -Nz at 1268°F containzng jrom 10 to 100 zol pct NH3 also obeyed a parabolic rate law. The rate of 'nitriding increased with an increase in the NH3 content of the gas Mixture. It is well-known that steel heated in gas mixtures containing ammonia (NH3) takes up much larger quantities of nitrogen than steel heated in nitrogen, both gases having a total pressure of 1 atm;' this phenomenon can presumably be attributed to the catalytic decomposition of NH3 on the steel surface to furnish nascent (monatomic) nitrogen. This process was studied bv Brunauer. Jefferson, Emmett, and Hend-ricks at furnace temperatures of 752" and 831°F2 using mixtures of NH3 in Hz. Englehardt and wagner3 reported that, at a furnace temperature of 914°F and under their experimental conditions, both nitriding and denitriding were controlled by the rate of gas-metal reactions at a steel surface rather than by the rate of diffusion of nitrogen in iron. The present study was undertaken to obtain information on the kinetics of nitriding low-carbon steel strip at higher temperatures so that practical rates for short-time strip-annealing treatments could be estimated. Variables studied included time: temperature, and NH, content in the annealing atmosphere. Mechanical and chemical characteristics of steel nitrided in this manner will not be considered in the present article. MATERIALS AND EXPERIMENTAL WORK The samples used were from a commercial low-carbon steel, 0.0244 cm thick, in the cold-reduced condition. The chemical composition of this steel is given in Table I. Panels were cut to 5.1 by 17.8 cm, degreased in toluene, and weighed just before treatment. Four specimens were nitrided under each of the experimental conditions. A study was made of the nitriding rate of steel in a 100 vol pct ammonia atmosphere, 740 mm pressure, at five specific temperatures within the range 964" to 1358°F. The nitriding rates of steel in ammonia-nitrogen gas mixtures containing 10, 18, 26, 50, and 100 vol pct ammonia, 740 mm total pressure, at 1268°F were also determined. All atmospheres used were dried by successively passing them through drying towers packed with soda lime and with Linde Molecular sieve Type 4A. Quoted gas compositions refer to those entering the furnace. Specimens were held in the constant-temperature zone of a vertical annealing tube furnace for times of 14, 3, 5, 10, or 15 min. Gas flow rates were maintained at 3.8 cu ft per hr, which was nineteen volume changes per hour for the system used. The rate of flow was selected to provide a high level of free NH3 for cracking on the steel surface where the ammonia gas is most effectively used as a nitriding agent. The vertical annealing tube furnace consisted of a Hevi-Duty tube furnace with a 2 1/2-in.-ID mullite ceramic high-temperature tube. The constant-temperature zone (controlled within 10°F) was about 10 in. long. After each specimen was degreased, a hole was punched in one end, for attaching the specimen by hook to a chain so that it could be lowered into or raised from the high-temperature portion of the tube by means of a power-driven winch. A stainless-steel access port with O-ring seals was connected by suitable glass-to-metal seals to the cool upper portion of the furnace tube. After the weighed specimen was placed in the access port, the furnace tube was evacuated to approximately 10"3 torr, and then the system was flushed thoroughly with the atmosphere under study. When the gas flow rate and constant-temperature zone of the furnace were established, the specimen was lowered into the constant-temperature zone. The atmosphere flowed from the top to the bottom of the vertical furnace tube and was then vented. For all these runs, during the first 3 min of the time the specimen was in the constant-temperature zone of the furnace the specimen was heating up to the tempera-

Jan 1, 1970

By Roland J. Lapee

A history of the progress made in copper refining in Montana is presented. The casting furnaces and the newly rebuilt electrolytic refinery are descmbed and operating details are given. Experiences with various addition agents, effects of rernoval of chlorine from the electrolyte, and effects of separan on electrolyte and on copper deposit are discussed. Observatzons are made on the effect of various impurities in anode copper and behavior of thezr salts in electrolyte and in slime, on treatment of slime for removal of copper, and on electrolyte puriification problems. The improved method for production of starting sheets is described. Attention is given to new materials for construction and to improvement in matrials handling ad quality control. COPPER refining at Great Falls, Montana, dates back to 1892. The original plant produced 65,000 lb. of cathodes per day at a current density of 16 amp per sq ft. In this plant anodes and cathodes were handled to and from the cells by means of hand-operated chain blocks and were moved about the plant by hand trucks. Cathodes were melted in coal-fired, reverbera-tory furnaces, were charged by hand, and refined copper was dipped by hand and cast in iron molds. In 1916, a new, modern plant with capacity to refine 18 million Ib. copper per month was built. Two refining furnaces were built, and each was provided with a twenty-mold Clark casting wheel. In 1922, the furnaces were converted to the use of pulverized coal: in 1923, to the use of oil; and in 1928, to the use of natural gas. In 1926, the plant was enlarged 50 pct. The use of pulverized coal for refining copper at Great Falls was discussed1 in a paper presented at the Salt Lake City meeting of the Institute in September 1925. Also, a comparison of the use of various fuels in copper refining furnaces was discussed 2 in a paper presented at the New York meeting of the Institute in February 1932. Prior to 1943 the cellroom was operated with twenty-five 720-lb. anodes and 26 cathodes per cell. Four cathodes, each weighing about 165 lb., were produced from an anode. In 1943 the weight of the anode was reduced to 460 lb., and two cathodes, each weighing about 190 lb., were produced from an anode. In 1949, a Billet Casting Wheel for the production of 3 in. diam phosphor deoxidized billets was built. A paper, presented at the New York Meeting of the Institute in February 1956, describes3 this plant and operation. In the electrolytic refinery, production is controlled by variation of the current density. In 1956, germanium rectifiers on a separate electrical circuit were provided for the starting sheet section so that current density could be maintained at any desired figure. Also in 1956, a program of modernization and enlargement of the Electrolytic Copper Refinery was started. This program, when completed, will raise the capacity of the Great Falls Electrolytic Copper Refinery approximately 33 pct, to 33 million lb. of cathodes per month. Furnace capacity for melting cathodes and anode scrap is ample to take care of the increased production from the electrolytic refinery. This fortunate condition came about primarily as the result of the three following changes: 1) The furnaces were lengthened 10 ft in 1922 when the fireboxes were found unnecessary for burning powdered coal. 2) Furnace life, and hence furnace capacity, was increased by the successful efforts of the Copper Refinery staff to develop a method for sanding furnace side walls and roof. 3) The natural gas being used has a very low sulfur content. As a result, it is possible to reduce time spent rabbling and poling and thus greatly increase tons per furnace charge. TANKHOUSE AND TANKS The Great Falls Electrolytic Refinery is 535 ft long by 252 ft wide. The cells are arranged in four crane bays in which are operated seven 10 ton Whiting Cranes. The old cells are in groups of ten, with circulation of electrolyte through five cells in cascade and with aisles between the groups of ten cells. The new cells are nested in groups of sixteen with no aisles, are all on one level, and have individual circulation to each cell. Present plans call for 1792 commercial cells and 128 stripper cells. To reduce confusion during the construction period, and to use the cranes, cars, and other equipment already on hand, the cells in the rebuilt section were designed to have approximately the same inside dimensions as the cells in the old part of the

Jan 1, 1962

By William T. Folwell

The Lucky Friday mine east of Mullan, Idaho, is an outstanding example of a property in the Coeur dlAlene district where a small and insignificant-appearing silver-lead-zinc vein at the surface has changed at depth into a large vein of great importance. The Lucky Friday vein has little if any surface expression and above the 1200 level the ore shoots are small and discontinuous. Between the 1200 level and 2450 level, the lowest developed level, the main ore shoot has shown remarkable improvement on each succeeding lower level, and today the mine is one of the major lead-silver producers in the Coeur d'Alene district. History: The Lucky Friday property is on the north side of the South Fork of the Coeur d'Alene River in sections 25, 26 and 35, T. 48 N., R. 5 E., Hunter mining district, Shoshone County, Idaho. This is about one mile east of the town of Mullan, which serves the eastern portion of the Coeur d'Alene district. The southern part of the property is crossed by a branch line of the Northern Pacific Ry. and by U. S. Highway 10. This highway is the principal road crossing the panhandle of Idaho and connects the district with Spokane, Wash., on the west and Missoula, Mont., on the east. The main portal and surface plant of the Lucky Friday mine is at an elevation of 3365 ft, only a short distance above the valley floor and a few hundred feet from U. S. Highway 10. so the mine is readily accessible for year-round operation. The property is comprised of six claims, known as the Lucky Friday group, owned outright by the Lucky Friday Silver-Lead Mines Co. There are four patented claims, Good Friday, Lucky Friday, Northern Light, and Lucky Friday Fraction No. 2 (Mineral Survey No. 3028), and two unpatented claims, Hunter and Creek. In addition, Lucky Friday owns an undivided one-half interest in the Hunter Creek property. which adjoins the Lucky Friday group on the north: a 90 pct interest in the mineral rights in the Jutila Ranch (160A), which adjoins the Lucky Friday group on the east; and a 60 pct interest in the Lucky Friday Extension claim group, which adjoins the Lucky Friday group on the west. The company also has a long-term mining lease on the Hunter Ranch, which adjoins the Lucky Friday group on the west. The claims of the Lucky Friday group were located between 1899 and 1906. The Lucky Friday Mines Co. was organized in 1906 and did considerable exploration work by surface trenching and shallow underground workings, only to see the property sold by the Shoshone county sheriff to satisfy labor claims totaling $2000 in 1912. Another firm, Lucky Friday Mining Co., bought the claims in 1914 and spent 12 years driving what is now known as the tunnel level crosscut. This tunnel intersected a vein previously exposed in a higher tunnel, but it was only a few inches wide. The vein was followed a short distance westerly but was so unpromising that the work was discontinued in favor of extending the main crosscut tunnel several hundred feet north. No ore was found and all work was discontinued. The property was held in such low esteem by the firm that it let taxes amounting to less than $15 a year go delinquent for nine years. Then the property lay idle for two more years until in 1938 John Sekulic, a Mullan service station operator, took a lease, with a $15,000 purchase option, on the advice of an old miner who had worked in the mine. Sekulic re-opened the tunnel level crosscut and explored the vein with an easterly drift for about 200 ft. The vein was too narrow to be of commercial value but was believed interesting enough to warrant further exploration at depth. Lacking funds to explore the vein at depth, Sekulic tried to get the district's larger operating companies to take over his lease and option. They were not interested because of the lean tunnel level showing and the fact that the property lay between the White Ledge fault on the north and Osburn fault on the south, an area which geologists always considered unworthy of exploration. Sekulic then organized the present company and assigned his lease and option to it for stock. This was in 1939. Enough stock was sold locally to finance sinking of a shaft 100 ft from the tunnel level east drift. The vein at this additional depth still was not commercial but showed some improvement. Treasury stock was offered at 5 to 10C a share

Jan 1, 1959

By A. O. Dady

PRODUCERS of both bituminous and anthracite coal have for many years been worrying about the gradually decreasing consumption of their product in the United States. Twenty years ago production had climbed to a peak of 569,000,000 tons of bituminous and around 90,000,000 tons of anthracite. In 1939 it was about 393,000,000 tons of bituminous and 51,000,000 tons of anthracite. This does not mean that the country is using less polver than it did twenty years ago, or that we are living in colder buildings. As a matter of fact, about three times as much electric power is now being consumed as in 1920. In general two factors are responsible for the drop in coal consumption: (1) greater efficiency in the use of coal, and (2) increased competition from hydro power, internal-combustion engines, and substitution of oil and gas for coal in conventional furnaces. As to the first factor, it took 3 Ib. of coal to generate a kilowatt-hour of electricity in central stations in 1920 and it takes 1.4 Ib. now; it took about 175 Ib. of coal in railroad locomotives to haul 1000 long tons a mile in 1920 where it talces about 120 lb. now: and marked economies have also been made in the use of coal in steel plants: in by-product coke ovens, and in other large uses. As to the second factor, three times as much electricity is produced from water power now as was produced in 1920; three times as much oil is being consumed in various ways; and likewise, the consumption of natural gas has tripled- whereas coal consumption in the same period has dropped by 40 per cent

Jan 1, 1941

By F. M. Morris

A large reserve of thick coal in southwest Virginia was developed by Clinch-field Coal Co. in 1957-1958 to produce a nominal rate of 1500 tph raw coal. Operation features coal cleaning in transit. Refuse removed averages 35 pct. Evolution of plant from initial conception to completion is discussed, selection of means applied is explained, and performance to date us design expectation is described. The long-range plan calls for ultimate handling capacity of 40,000 tpd raw coal with anticipated clean coal capacity of 26,000 tpd on a three-shift basis. From a materials handling viewpoint, the Moss No. 3 operation is principally of interest as an ensemble. Generally speaking, it uses time-tested equipment and ideas but some of these are employed on a scale that may be new in the industry. At present about 20,000 tpd of raw coal are being handled. This is expected to increase to 40,000 tpd as soon as business conditions justify it. The purpose of this paper is to discuss the evolution of the plant as to materials handling practices, to describe briefly the equipment and methods used, and to comment on performance in certain areas. The subject divides itself naturally into four phases: 1) operations at the mines, 2) transportation from mines to plant, 3) raw coal handling into the plant, and 4) transportation of refuse away from the plant. OPERATIONS AT THE MINES The coal reserves for Moss 3 are in southwest Virginia where Dickenson, Russell, and Buchanan Counties come together. This area contains about 15 square miles and over 100,000,000 tons of coal. Here the No. 4 (Tiller) and No. 5 (Jawbone) seams of the Norton Formation lie so closely together that for practical purposes, they constitute one seam of coal. This seam, which is called the Thick Tiller, varies from 10 to 18 ft in thickness and underlies Sandy Ridge, a mountain cresting between 2400 and 3300 ft above sea-level. At 18-ft seam height (which is considered to be the maximum practical mining height) the parting between seams will be about 3 1/2 ft thick and each bench of coal will be about 7 ft thick. Depending on the amount of impurities in the seams Drover. total reject in this height coal may approximate 50 pet by wt. It will average about 35 pet for the property as a whole. The first move in developing this resource was made in 1953 when contour maps of several square miles around Duty, Va., including all the known outcrop, were made by photogrammetry. At this time, it was felt the prospective operation would be served by the Clinchfield Railroad and a photogrammetric route survey was made by this railroad from Haysi to Duty. Study of the resulting maps indicated only one site—adversely owned—which might accommodate the size washing plant to be erected. Water resources of a dependable nature seemed nonexistent. In 1954 bulk washability tests were made on the Tiller Bench at the Moss No. 1 preparation plant. The tests indicated this portion of the seam, mined separately, would wash to 4 pet ash with good recovery. Also in 1954, development of Moss No. 2, south of Sandy Ridge, was begun. This mine is in the Tiller Seam where it is about 100 ft below the Jawbone Seam. The reasons for developing a mine in the normal Tiller Seam before tackling the Thick Tiller seemed compelling: Railroad service could be established quickly, communications were better (though not good!), more was known of the seam (it had been mined in the years 1911 to 19241, and there was no essential property to be acquired. After some legal skirmishing, the Norfolk & Western Railroad was granted the right to serve the new mine. Three decades earlier the old mine had been served by the Clinchfield Railroad. The event which triggered active development of Moss 3 was the Appalachian Power Co.'s decision in 1956 to build a 450,000-kva power plant on Clinch River at Carbo. This solved the problem of marketing the steam coal which inevitably must be a product of a mine in the Thick Tiller. Management promptly decided to build the preparation plant at Carbo where an excellent site was owned; where railroad service existed; where telephone service could be obtained; and where roads, bridges, and water supply were tolerable.

Jan 1, 1961

By W. J. Schuster

Safety in coal mines depends largely upon adequate training of the foreman. Although management must provide modern and safe equipment and at all times keep mines in first class condition from a safety viewpoint, final results will be determined by the quality of supervision. HANNA COAL CO., Division of Pittsburgh Consolidation Coal Co., operates three large underground mines in eastern Ohio. The section of Pittsburgh No. 8 coal seam in which these mines are located varies in thickness from 52 to 64 in. It is immediately overlain by a stratum of shaly material 12 to 15 in. thick locally known as draw slate, which is structurally very weak and which disintegrates rapidly upon exposure to atmosphere. Immediately above the draw slate as it is normally found is a band of extremely high ash material 6 to 12 in. thick known as roof coal or rooster coal, and above this is a stratum of conglomerate material varying from 4 to 10 ft in depth. Overlying the conglomerate is a relatively thick stratum of limestone, the first stable material above the Pittsburgh coal seam in eastern Ohio. With the method of full-seam mining that has been adopted, draw slate is shot down, loaded with the coal, and removed in the preparation plants. The roof coal then becomes the permanent roof. The major problem in mining the No. 8 seam in eastern Ohio is control of the roof. Since the strata above the draw slate contains no material with a structure firm enough to provide self-support, the roof begins to sag in a relatively short time after the coal and draw slate have been removed. The problem thus becomes one of getting temporary safety posts under this roof as quickly as possible to prevent a break or separation from occurring either in the roof coal or in the conglomerate above it. Haulage System The Pittsburgh No. 8 seam in eastern Ohio is relatively level, with only minor local dips. Throughout the Hanna Coal Co. mines, entries are generally 12 ft wide. Rooms are driven on a 60" angle on 30-ft centers and are 22 ft wide. No attempt is made to extract the 8-ft pillars between. The entire length of main line haulage is gunited in one mine, and a major portion in another. Two of the mines have single-track main haulage roads with passways. The third, a new mine, is double-tracked, and the roof is supported by steel crossbars, 60 lb or heavier, spaced on 4-ft centers and lagged. In recent years timbering on main line and secondary haulage roads has been accomplished by one of two methods: 1—crossbars are supported on a small section of post set in a hitch hole in the rib, or 2—or a hole is drilled in the rib about 12 in. below the roof, of sufficient depth to fasten securely a short length of 40-lb rail, the bottom of the rail facing the roof, on which a short post is set directly under the crossbar. At present the hitch-hole timbering method is favored. At two of the mines the main line haulage locomotives are 26-ton, 8-wheel units. These locomotives are of the axleless type, each wheel being individually mounted on the frame. The motorman's compartment is encircled by 3-in. armor plate for the protection of the occupants. At the third underground mine conventional 15-ton locomotives are being used. However, these locomotives have been completely rebuilt in the company's shops. Equipment has been streamlined and quarters have been provided for two people, who are protected by heavy steel plate in much the same way described above. This modernization program has been completed on all secondary haulage locomotives at the three mines, and the company is well on the way to similar equipment of the 6-ton section locomotives. The following additional features have been included in their modernization: 1—additional support for the motors to prevent their falling to the middle of the track and derailing the locomotives should a break occur in the suspension bar support; 2—installation of additional bracing to prevent brake rigging from becoming displaced and causing derailments; 3—enclosure of all electric wiring in conduit or raceway; 4—provision of an enclosed compartment for the storage of re-railers, jacks, and other equipment, so that they need not be carried on the outside of the motor; and 5—redesign of the end of the locomotive opposite the operator's compartment to prevent anyone's mounting from that direction. It is interesting to note that some

Jan 1, 1955

By R. E. Hudrlik, G. W. Preckshot

The diffusion coefficients of silver from solid silver in molten lead were measured to within ± 0.8 pet in a columnar type diffusion cell ower, the temperature range of 326° to 530°C. Fick's law describes the process up to 530°C where the laminar mechanism appareltly breaks down. These is negligible resistance at the interface as shown by mathematical analyses. The diffusion coefficients are found concentration independent. IT would seem that diffusion in liquid metals would be free of such effects as molecular structure, dissociation. polarization. and compound formation. This view was taken by Gorman and preckshot in their study of diffusion of copper from solid copper into molten lead. They reported diffusion coefficients which were independent of the concentration over the range of 478° to 750°C. They found that the Stokes-Einstein equation with constant radius of the diffusing specie represented the diffusion data better than Eyring's rate theory equation and Sheibel's correlation. The radius of diffusion was found to be that of the doubly charged copper. There appeared to be no resistance across the solid-liquid boundary. In the present work the diffusion coefficients for silver in liquid lead were measured over a range of temperatures of 350° to 505°C. The solubility of silver in lead over the range of 303° to 630°C was also obtained. These results are compared with calculated or correlated values or with data in the literature. EXPERIMENTAL Procedure—The experimental equipment techniques and procedures were those reported in detail by Gorman and preckshot9 and will not be repeated here. Measured values of WT, Co, A. L were obtained for various diffusion times and the diffusion coefficient was computed for the case of no resistance at the interface9, 11 by: WT/CoAL = 1- 8/p2 n=1 1/(2n - 1)2 exp[-(2n - 1)2p2 Dt/4L2] [1] or where there was resistance at the interface by: WT = 1- ?n=1 2h2/ap2L [sxp [-Dan2t]/[(h2 + an2) L + h] The roots an are those of the transcendental equation3 tan (an L) = Iz/cun. The diffusion coefficient is that defined by Hartley and Crank.7 The total silver in the lead cylinder and equilibrium slug was determined by a cupellation technique' with proper correction for losses. Analysis of known samples showed that this method is surprisingly accurate. The amount of silver in the lead adhering to the silver cylinder was obtained in the same fashion as shown by Gorman and preckshot.9 The small errors involved in this determination are not critical since the silver in this adhering lead layer is only 3 to 15 pet of the total diffused. Materials—Electrolytic silver containing 99.9+ pet Ag obtained from General Refineries of Minneapolis, Minn. was used for all but runs 7 and 8. For the balance of the runs this silver was reduced with hydrogen at 1100°C and its oxygen content was found to be about 0.017 pet. For the runs. 7 and 8, phosphorous-reduced silver of the same purity was obtained from Handy and Harman Co. of Chicago, Ill. The densities of the phosphorus-reduced silver and the hydrogen-reduced electrolytic silver were 10.484 and 10.487 g per cm3, respectively. These values agree with those reported for pure silver. Silver which was reduced at 900 C had an average density of 9.998 g per cm3, indicating porosity. This silver was used for a number of runs which were not tabulated in Table I. These are indicated by crosses on Fig. 2. The 99.999 pet Pb was obtained from the National Lead Co. Research Laboratory of Brooklyn, New York. DISCUSSION OF RESULTS The diffusion and solubility results are reported in Table I for eleven runs using either phosphorus-reduced electrolytic silver or hydrogen-reduced silver at 1100° C. The solubility data shown in Fig. 1 show the excellent agreement with that reported by Heycock and Neville.8 The data of Friedrichs5 apparently are in error. The experimental solubility data of this work are reported to 0.3 pet. The experimental diffusion coefficients computed from Eq. [1] are reported within 1.2 pet of the mean and are plotted in Fig. 2. These are expressed within +0.8 pet of the experimental values over the entire temperature range by: D= 8.26 x 10 -5 e-1925/RT . [3] There appears to be little difference due to the

Jan 1, 1961

By M. Bevis, A. F. Acton, P. C. Rowlands

Defor~nation twinning and transformation twinning modes most likely to be operative in Fe-Ni and Fe-Ni-C martensites have been determined using a new theory of the crystallography of deformation t~inning.~ This analysis shows that potentially important conventional and nonconventional twinning modes1 have been omitted in previous analyses. Discussion is given on the relevance of the predicted twinning modes to the lattice invariant shear associated with the martensite transformation in steels and to anomalous deformation twinning in Fe-Ni-C martensites. THE two most important criteria which appear to govern operative twinning modes in metallic structures1 are that the magnitude of the twinning shear should be small and that the twinning shear should restore the lattice or a multiple lattice in a twin orientation. The latter criterion ensures that the shuffle mechanism required to restore the structure in a twin orientation is simple. These criteria have been adhered to in the prediction of twinning modes2"6 in bcc and bct single-lattice structures with axial ratios in the range y = 1 to 1.09 as, for example, encountered in martensite occurring in steels. Refs. 2 and 3 in particular consider the martensite transformation in steels and the twinning modes in these cases relate to transformation twinning, and hence the lattice invariant shear associated with the martensite transformation. The list of twinning modes which can be compiled from these sources is incomplete and the ranges of magnitude of shear considered could be unrealistically small, particularly in the case of deformation twinning. The latter consideration is supported by the fact that twinning modes with magnitudes of shear large compared with the smallest shear consistent with a simple shuffle mechanism have been established in, for example, the single-lattice structure mercury7 and the multiple-lattice structure zirconium.&apos; In addition the anomalous deformation twins reported by Ftichrnan4 to occur in a range of Fe-Ni-C martensites still remain unexplained. It is clear that a comprehensive analysis of twinning modes likely to be operative in martensite In steels is required. The results of the application of a new theory of the crystallography of deformation twinningg to these structures are presented in this paper. The theory has been used to determine all shears which restore the lattice or a multiple lattice in a new orientation with magnitude of shear up to a required maximum. The orientation relationships between parent and twinned lattices are not restricted to the classical orientation relationships of reflection in the twin plane or a rotation of 180 deg about the shear direction. PREDICTED TWINNING MODES Twinning modes which restore all or one half of lattice points to their correct twin positions will be referred to as m = 1 and m = 2 modes, respectively. These modes are the most likely to describe operative modes in single lattice structures. The bcc m = 1 and m = 2 modes which have magnitudes of shear s in the range s < 2 and s < 1, respectively, have been given10 and are reproduced here in Tables I and 11. Detailed discussion of the crystallography of these modes and cubic modes in general will be discussed elsewhere (~evis and rocker, to be published). The four twinning elements Kl, &,ql,7)2 as well as the magnitude of shear s are given for each twinning mode, and the twinning modes are given in order of increasing shear. Two twinning modes are given in each row of the tables, the twinning mode Kl, Kz, ql, q2 and the reciprocal twinning mode with elements Kl = K,, Ki = Kl, q: = q2, and 17; = ql. The m = 1 and m = 2 twinning modes which describe twinning shears with small magnitudes of shear and simple shuffle mechanisms in bct crystals with -y = 1 to 1.09 are given in Tables I11 and IV, respectively. On increasing the symmetry of the tetragonal lattice to cubic, that is making y = 1, all modes listed in Tables 111 and IV must reduce to crystallographically equivalent variants of the modes given in Tables I and 11, respectively, or become twinning modes with both shear planes as symmetry planes in the cubic lattice and hence not considered in Tables I and 11. With the exception of this last type of mode only those tetragonal twinning modes which reduce to modes 1.1, 1.2, 2.1, and 2.2 of Tables I and I1 are considered in Tables 111 and IV. For values of y in the range -y = 1 to 1.09 the tetragonal modes and the corresponding cubic twinning modes have approximately the same magnitude of shear. The twinning modes listed in Tables 111 and IV are therefore by the criteria given above the most

Jan 1, 1969

By W. A. Tiller, W. C. Johnston

A high-temperature electromagnetic stirrer is described in which heating and stirring are accomplished by independently controlled power sources. The appavatus is suitable lor use at temperatures up to 1700°C in a variety of ambient atmospheres. Some typical examples of the homogenizatimz capabilities of the system are given. THERE are few processes in solidification that are not markedly affected by motion in the melt during freezing. In many instances, the mechanisms are diffusion-controlled, and the transport in the melt may be greatly accelerated by deliberately stirring the melt. In zone-refining, stirring1 assists the removal of rejected impurities from the interface, so the process proceeds at a faster rate. The transition from a planar to a cellular interface is caused by constitutional undercooling in the melt ahead of the interface: and stirring delays its onset. Stirring is valuable for homogenization of melts: and chemical reaction with sluggish kinetics may be accelerated. Finally, it has been observed that grain refinement is related to motion in the melt. Fine grain castings are usually produced by the addition of catalysts to the -melt,&apos; catalysts which are thought to act simply as hetereogeneous nucleation centers. Even here motion is important. Richards and Rostoker 5 applied ultrasonic vibration to a solidifying A1-Cu alloy which had been innoculated with a catalyst and found that the grain diameter fell linearly with the amplitude, the peak acceleration and the power input to the melt from the transducer. Finally, mechanical and electrical stirring alone have been used to generate a fine-grained structure.6,7 Johnston ef a1.&apos; have carried out a series of systematic investigations of grain refinement by electromagnetic stirring in a number of low melting point alloys. They found, for example, that the number of grains per unit volume in Pb-Sn alloys could be increased several orders of magnitude by stirring an undercooled melt at the moment of recalescence. In general, a relation AT .H = constant prevailed for a given grain size, where AT was the undercooling of the melt and H the field strength. In more recent work, deliberate homogeneous nucleation of slightly undercooled melts established that the mechanism of refinement must be one involving crystal fragmentation and subsequent multiplication, rather than a "shower" of nuclei effect.9 It is the purpose of this note to describe a stirring device suitable for use up to 1700°C. At low temperatures mechanical stirring and direct-current methods are feasible, but at high temperatures the problem of a protective atmosphere and of electrode corrosion rules out such procedures. The most convenient method for high temperatures is to use externally generated ac fields for both stirring and heating. With rf induction heating alone, considerable stirring and agitation can be achieved, but in general the penetration of field into the melt is small, and the stirring cannot be controlled independently of the heating. In the present experiments, separate power sources of different frequencies for heating and for stirring were used. A susceptor design was chosen so that the 450 kc rf heating field was completely absorbed in the susceptor. The stirring frequency, 400 cps, hereafter called the af field, was chosen so that a high penetration of the melt proper was achieved. EXPERIMENTAL APPARATUS The apparatus, Fig. 1, consists of a quartz tube and end plates, surrounded by an rf induction coil and six equally spaced af stirring coils, four of which are shown in full and a fifth in section. Each af stirring coil is a transformer of which the secondary is a single-turn water-cooled copper loop and the primary is composed of two 10 amp-117 v Variac cores as shown. These cores are cooled by forced air, as each of the six pairs will carry maximum currents of 15 amp for short periods. Each set of Variac windings are connected in series, but opposite sets are connected in parallel with a three-phase 400 cps 400-v source. By properly phasing the coils in this way, a rotating field is produced. Capacitors C1, C2, and C3 in Fig. 2 are used to match this inductive load to the generator. Fig. 3 shows a cutaway view of the quartz tube. The sample (1 in. diam by 1 in. high) is placed in a tapered alumina crucible. An axial W-26 pct Re thermocouple, enclosed by a protection tube, is provided. The cruci-

Jan 1, 1968

By O. F. Kammerer, W. E. Miller, D. H. Gurinsky, J. Sadofsky, J. R. Weeks

Zirconium and titanium inhibit solution mass transfer of steels by liquid bismuth, mercury, and lead. It is shown that in bismuth and mercury, these adsorb on the surface of the steels and subsequently react with nitrogen and possibly carbon from the steels to form inert, adherent surface layers of ZrN, TiN, or TiN + Tic. Data are presented which describe the condition under which thase deposits form. These inhibitors decrease the solution rate of iron into bismuth, and require a higher supersaturation for precipitation of iron from bismuth. USE of the low-melting heavy metals (bismuth, lead, mercury, and their alloys) as coolants has been limited because solution mass transfer of steels occurs in these liquids; i. e., iron dissolves in the hot sections of the heat transfer circuit and deposits in the colder sections. The rate of solution of iron and the temperature coefficient of solubility are sufficiently great to cause complete or partial stoppage by the deposition in the coldest section of a closed circuit in finite time, even though the actual solubilities are extremely low. In the development of the mercury vapor turbine by the General Electric Co., Nerad and his associates1 discovered that the addition of as little as 1 ppm Ti or Zr to magnesium-deoxidized mercury reduced the mass transfer of ferrous alloys by mercury to a negligible amount. Reid2 reported that titanium was detected chemically on the surface of steels contacted with this mercury alloy in amounts varying from 2.0 to 2.6 mg per sq in., the greatest amount being found in the hottest portion of the circuit. Reid stated that the titanium forms the intermetallic compound Fe2Ti by reaction with iron on the surface of the steels. This compound was presumed to be highly insoluble in mercury. More recently, El-gert and Egan3 have reported a greater than 100-fold reduction in the rate of mass transfer of a 5 pet Cr steel by liquid bismuth upon the addition of titanium (in excess of 50 ppm) and magnesium (350 ppm) in the liquid metal, during experiments performed in thermal convection loops* over the temperature differential 700° to 615° C. Also, Shep-ard and his associates&apos; have reported that the addition of titanium to liquid bismuth and Pb-Bi eutec-tic produced a marked decrease in the rates of solution of both iron and chromium from type 410 steel capsules under static conditions. This inhibiting effect increased with repeated reuse of the capsules. Tests performed in this laboratory under carefully controlled conditions have shown that the addition of zirconium and magnesium, or titanium and magnesium, to liquid bismuth or lead greatly reduces the rate of mass transfer of chromium alloy steels and carbon steels in thermal convection loops with a maximum temperature of 550°C.5-9 The present paper will review the data obtained to date at this laboratory on the behavior of iron and steels in contact with liquid bismuth alloys containing titanium or zirconium, and will attempt to explain the role of the above additives in reducing solution mass transfer. Reaction between the Zirconium or Titanium Dissolved In Liquid Bismuth and an Iron or Steel Surface Reaction between Zirconium Dissolved in Bismuth and the Surface of Pure Iron-—A small pure iron crucible (analyzed by the supplier to contain 0.8 ppm N was contacted with bismuth containing approximately 0.1 pet Mg and varying amounts of a radioactive zirconium tracer. The crucible was then inverted at the temperature of contact. The thin residual layer of adherent bismuth was dissolved in cold, concentrated nitric acid. The crucible surface and the solidified bismuth were then analyzed for radioactive zirconium. An analysis of the activity loss on the crucible surface and the weight loss of the crucible during the nitric acid treatment showed that the acid treatment removed the zirconium that had originally been dissolved in the adherent bismuth, but not any zirconium that may have reacted with the crucible surface. The crucible was then pickled in warm aqua regia to remove all surface activity, hydrogen-fired at 600°C, and recontacted with a new liquid alloy. The results of the experiments contacted 1 hr at 450°C show, Fig. 1, a Langmuir-type adsorption with an adsorption free energy of approximately 17 keal per g atom Zr.5 This deposit was estimated to contain 1 atom of zirconium for each 7 to 8 iron atoms on the crucible surface, assuming a surface roughness factor of the pickled crucibles to be five. Increasing the temperature to 520°C caused consi-

Jan 1, 1959

By Henry M. Rowe

Professor Ruff&apos;s most illuminating paper&apos; describing his extremely valuable investigation of the carbon-iron equilibrium diagram assigns definite temperatures to certain very important lines in the diagram, viz.: the eutectic temperature EB (Fig. 1) and the " near-eutectic " temperature E&apos;B&apos;, but without indicating clearly enough the reasoning which warrants these assignments. These temperatures are of such primary importance that it is well to indicate the apparent gaps in the reasoning, in the hope that they may be filled in. In his diagram, Fig. 1, the point X and the liues B&apos;D&apos;lII represent the experimental results reached by Ruff and Goecke;a AB, AEB, and SE represent the liquidus, the solidus, and the solubility of the carbon in solid austenite in the absence of graphite according to Gutowsky,3 AB being also in accordance with Spiecker&apos;s data; S&apos;E&apos; the solubility of carbon in solid austenite in the presence of graphite according to Ruer and Iljin;5 GS the line A3 after Ooerens and Meyer; 6 and EBC the true eutectic line as inferred by Prof. Ruff from Gutowsky&apos;s data. B&apos;D&apos; is the carbon content found in cast-iron which had apparently saturated itself with carbon in the presence of the graphite of the enclosing graphite crucibles in which the determinations were made. From its straightiiess and its sharp inflection near 6.67 per cent. of carbon, the carbon content of

Jan 1, 1913

By G. B. Misra

While investigating on the aerodynamic resistance of airways supported with peripheral timber sets, at regular intervals, the following theoretical equations were developed by the author to estimate the resistance coefficient of such airways: [ ] for S < 1, where f is Darcy-Weisbach resistance coefficient of the airway, C is modified drag coefficient of the supporting member, D is equivalent diameter of the bare airway, 8 is ratio of the approach velocity over the sets to the average velocity of the bare airway, A is cross-sectional area of the bare airway, a is projected frontal area of the sets, A., is cross-sectional area of the air stream at the vena contracta inside the set, S is spacing of the sets, f, is resistance coefficient of the bare airway, l is length of aerodynamic influence of sets, p is perimeter of the bare airway, p, is setted portion of the perimeter of the bare airway, pe is unsetted portion of the perimeter of the bare airway, and P shielding factor. The equations were verified experimentally in a model rectangular airway supported with one- (bars), three-, and four-piece sets of square-section timber of three different sizes and were found to hold true. The work has been further extended to one-, three-, and four-piece sets of round timber of 2.6, 3.2, and 3.8 cm diam with the same experimental set up. Tests have been carried out for spacings of 25, 50, 75, 100, 150, and 200 cm over a regime of flow defined by the Reynolds number (with respect to the equivalent diameter of the bare duct) ranging from about 1.5 X 106 to 5 X 106 using the same experimental techniques. The values of f are calculated in the manner indicated in [Ref. 1]. Unlike with square-section timber, the resistance coefficient f of the airway setted with round timber shows a distinct variation with the Reynolds number of flow. This conforms to observations made by Sales and Hinsley.2 In order to have a comparable value of f for all types of sets with all sizes of timber, it was necessary to select the value of f at a fixed Reynolds number of flow. Since f is chiefly a function of the drag coefficient of the sets, the appropriate Reynolds number RE is that with respect to the diameter of timber in the set. Considering the diameters of timber used and the regime of flow over which measurements were made, f was chosen at a value of RE = 20,000 in all cases. The f vs. S curves are maximal in nature and in conformity with theory, the f vs. 1/S curves are straight lines up to a value of S = 1 beyond which they show a distinct flexure. The observed values of 1, the length of aerodynamic influence of sets, agree with the relation 1 = 42 e, developed for square-section timber sets, thus suggesting that the shape of timber has little influence on the length of aerodynamic influence. The value of the modified drag coefficient CD for round timber was calculated in the same way as for square timber in Ref. 1, taking the contraction factor Z = 1.5 for round-edged constrictions. CD has an average value of 0.96 with a standard deviation of 6.08% as compared to the free stream drag coefficient of 1.2 at RE = 20,000 for long cylindrical obstructions The shielding factor [ ] is plotted against S/1 in [Fig. 1]. The curves are more or less independent of the size of timber, but are different for the different types of sets, possibly due to their different degree of symmetry. Values of f calculated by the author&apos;s [Eqs. 1 and 2], using experimental values of CD&apos; and [ ] and taking I = 42 e, are plotted in [Fig. 2] against experimentally measured values of f for different types of sets with different sizes of round timber. The values agree closely with a standard deviation of only 5%, thus establishing the veracity of the theoretical equations developed by the author for round timber as well. A comparison was made between the Xenofontowa4 equations (the only other reasonable relations available for the estimation of the resistance coefficient of supported airways) and the author&apos;s [Eqs. 1 and 2] by comparing in [Fig. 3] the values of the resistance coefficient f computed by the Xenofontowa relations with those experimentally measured by the author. In order to make

Jan 1, 1975

By C. S. Barrett, A. Smigelskas

There have been no publications of the deformation and recrystallization orientations of the metal beryllium, yet pronounced textures would certainly be anticipated since it is close-packed hexagonal in structure. Having an axial ratio approximately that of magnesium, beryllium probably deforms by nearly the same slip and twinning mechanisms that operate in magnesium, and the textures are likely to be similar or but slightly different from the magnesium textures. In the tests reported below this is found to be the case; the textures are found to differ from those of magnesium only in the details of the scatter from the average orientation. This report covers not only samples rolled at room temperature, but some rolled at elevated temperatures. Since magnesium has been suspected by some investigators of altering its crystallo-graphic deformation mechanism at elevated temperatures, it was considered possible that beryllium might do so and alter its textures accordingly. No pronounced alterations were found, however. Unfortunately, the theory of deformation textures is not in a state of development that permits one to deduce the deformation mechanism from a knowledge of the textures, which means that the similarity of textures at different rolling temperatures, reported here, cannot be taken as definite evidence that the deformation mechanism is actually the same at all temperatures. The general similarity of the deformation textures of magnesium and beryllium also extend to the recrystallization textures of the two metals, judging by the pole figures for recrystallized sheet presented in this report. Samples were prepared in the form of composite sheets made up of small pieces stacked in a pile. Each piece was trimmed with scissors so that an edge was parallel to the rolling direction, dipped in paraffin, and assembled into the pack by aligning it under the cross hair of a microscope. As the desired orientation was obtained on each piece it was secured in place by touching with a hot wire to melt the paraffin. A stack of ten or fifteen pieces was built up in this way, then trimmed to the shape of a T; the portion to be X rayed was then etched to the shape of a wire about 0.045 in. diam with 6N HCl. This method of shaping the sample is a modification of that used by Bakarian on magnesium.&apos; The absorption of the rays in the sample was so slight that it caused no difficulty in interpreting the films. Exposures were made with a 0.030 in. diam pinhole, using molybdenum radiation (40 kv, 25 ma, Type A film at 5 cm, 2 to 3 hr exposures). With the recrystallized specimens it was found necessary to oscillate the specimen so as to reduce the spottiness of the lines. A range of oscillation of 5" was SUB- cient to produce reasonably satisfactory patterns, though the quality was somewhat inferior to that of the deformation texture patterns, and only two degrees of intensity were read from the arcs on the films. Typical photo-grams for each of the deformation textures and the recrystallization texture are assembled in Fig 1. The pole figures were plotted in the usual way with the intensity of the various portions of the diffraction rings estimated by eye. Seven to nine films were made of each sample and each was carefully read in plotting the pole figures. Typical series included exposures with the beam normal to the rolling direction and at 11, 26, 41, 56 and 71" to the cross direction, plus two exposures with the beam normal to the cross direction, and at 11 and 79" respectively to the rolling direction. The rolling was in each case considered sufficient to develop the final texture: the reduction by cold rolling was 84 pct (from 0.0045 to 0.0007 in. thickness), following prior hot rolling in longitudinal and transverse directions and recrystallization; the reduction by hot rolling at 800°C was 90 pct (0.010 to 0.001 in.), following similar prior treatment; the reduction by rolling at 350°C was 88 pct (from 0.005 to 0.0006 in.) after similar prior treatment. The recrystallization texture was determined on a sample rolled at 350" to a reduction of 88 pct (0.0165 to 0.002 in.) after similar prior treatment, then mounted between steel strips to keep it flat and annealed at 700" in an atmosphere of argon. Discussion of Results The results of the X ray determinations are assembled in the pole figures of Fig 2, 3, 4 and 5 for rolling at

Jan 1, 1950

By Winston W. Spencer

ALTHOUGH by far the largest A consumer of platinum metals in the world, the United States until recently has been in- significant as a producer. Writing in the "Minerals Yearbook" for 1939, H. W. Davis states that 1938 probably marked the beginning of an- other period of transition in the world production of the platinum metals. In that year, he says, their production in the United States jumped to 49,380 oz., or fourth in the rank of world output. In the ten-year period 1925 to 1934 United States output ,averaged only 8300 oz. Practically all of this sudden in- crease in production is attributable to the introduction of mechanized equipment at the placer deposits in the Goodnews Bay district of Alaska. This district adjoins the Bering Sea and Kuskokwim Bay (accented on the first syllable), and is approximately 120 miles south of the mouth of the Kuskokwim River. The nearest village and post office is Platinum, which lies about eleven miles north- east of the Goodnews Bay Mining Company&apos;s camp. A good gravel road connects the mine camp and the village.

Jan 1, 1940

By M. R. Tek, F. F. Craig, J. O. Wilkes, C. S. Goddin

The waterflood performance of a water-wet, stratified system with crossflow is computed by a finite difference procedure. The effects of five dimensionless parameters on tile oil displacement efficiency, water saturation con-tour.7 and crossflow rates are evaluated in the absence of gravity forces. Crossflow due to viscous and capillary forces is shown to exert a significant effect on oil recovery in a field-scale model of a two-layered. water-wet sandstone reservoir. The crossflow is at a maximum in the vicinity of the front advancing in the more permeable layer. Under favorable mobility ratio conditions, the comparted oil recovery with crossflow always is interrnerliate between that predicted for a uniform reservoir and that for a layered reservoir with no crossflow. INTRODUCTION The important erects of reservoir heterogeneity on waterflood performance are commanding increased attention in the technical literature. Much of this attention is centered on two categories of layered reservoirs: those in which layers are non-communicating and those in which crossflow of fluids occurs between the layers. In the first category, the reservoir is assumed to consist of discrete layers, each uniform within itself and differing from the others only in such properties as thickness, porosity and absolute permeability. The performance within each layer is calculated by one-dimensional flow theory, and the performance of the total reservoir is obtained by summing individual layer performances. Capillary and gravity effects usually are not considered. Representative publications dealing with thi5 type of reservoir are those of Stiles,&apos; Dykstra and Parsons,&apos; Hiatt,3 Warren and Cosgrove&apos; and Higgins and Leighton." Prediction of performance for reservoirs in the second category is considerably more difficult since viscous, capillary and gravitational forces all play important roles in causing crossflow between layers. A number of authors have investigated the simpler problem of two-dimensional displacement flow in a stratified system with a mobility ratio of unity and negligible capillary and gravity effects.&apos;; Others have considered two-dimensional, non-steady-state flow of a single, slightly compressible fluid in a stratified reservoir. A limited number of laboratory oil displacement tests in layered models with crossflow have been reported. Miscible floods (with resultant zero capillary forces) in layered five-spot models were conducted by Dyes and Braun," who studied the effect of mobility ratio with zero gravity forces, and by Craig et al. 12 who studied the effect of gravity forces at constant mobility ratio. Waterfloods in layered five-spot models (with cross-tlow due to capillary, viscous and gravity forces) were conducted by Gaucher and Lindley,"&apos; who showed the effect of gravity forces in causing underrunning of the injected water and by Carpenter, Bail and Bobek, 14 who demonstrated the reliability of Rapoport&apos;s" dimension-less parameters for scaling layered systems. Waterfloods in rectangular layered models were conducted by Richardson and Perkins."&apos; who investigated the effect of velocity at constant mobility ratio and with zero gravity forces, and by Hutchinson," who studied the effects of varying mobility, layer permeability and layer thickness ratios. The differential equations which rigorously describe waterflooding in a heterogenous porous medium are non-linear and do not facilitate analytical solution. By using finite difference approximations it is possible to obtain a solution to any desired degree of accuracy. Such a solution, using an alternating direction implicit procedure (ADJP), is described by Douglas, Peaceman and Rachford.18 In the present study, a computer program using ADJP explores systematically the effects of important parameters on waterflood performance of a two-dimensional, two-layered, field-scale model of a water-wet sandstone system. Particular attention is given to evaluation of the water saturation contours and crossflow rates at the interface between layers to gain improved understanding of the crossflow mechanism. PROCEDURE BASIC: FLOW EQUATIONS The basic flow equations for two-dimensional, two-phase, immiscible, incon~pressible flow in a porous medium are:

Jan 1, 1967

By R. L. Schmidt, J. D. Verhoeven

The diffusion coefficients and electrotralzsport mobilities of aluminum, gallium, and arsenic have been determined in molten germanium with the capillary reservoir technique. The diffusion coefficients are slightly higher than those determined by the solidification technique of Brirton et a1.1 It was found that aluminium and gallium migrate to the cathode in molten germanium) whereas arsenic migrates to the anode under an electric field. A possible technique for producing p-n junctions by means of electric field freezing is discussed in view of the experimental data. In the past decade or so the literature has abounded with articles concerning the solid-state properties of the semiconducting elements germanium and silicon. It is surprising by comparison how little has been published concerning the liquid-state properties of these materials. In the production of doped germanium or silicon the diffusion coefficient of the dopant in liquid germanium or silicon is an important quantity. Yet, to the best of our knowledge, no one has ever made a direct measurement of the diffusion coefficient of solutes in either molten germanium or silicon. A number of authors&apos;-4 have determined some diffusion coefficients by the indirect technique developed by Burton et al.1 This technique appears to give numbers of the correct order of magnitude but it does depend upon a number of assumptions in the analysis and also requires prior knowledge of the viscosity of the melt. In the present work a more direct measure of the diffusion coefficients of arsenic, gallium: and aluminum in molten germanium has been made. It has been shown independently by two groups of authors5, 6 that, when an electric current is passed through the interface of a solidifying metal containing an impurity, the resulting electrotransport in the liquid boundary layer at the interface offers a mechanism for control over the redistribution of the impurity accompanying the solidification. These authors show that the degree of control is proportional to the mobility, U of the impurity atoms in the liquid boundary layer, where the mobility is defined as the drift velocity relative to the solvent per unit electric field. In principle the results apply to the solidification of a doped semiconductor and offer an interesting technique for control of the doping of a semiconductor. In the present work the mobilities of aluminum, gallium, and arsenic were determined in molten germanium in order to evaluate the potential usefulness of electro-transport as a technique for solute control when applied to the solidification of doped germanium. EXPERIMENTAL PROCEDURE Both the diffusion coefficients and the mobilities were measured using high-purity alumina capillaries and a resistivity technique for analysis. A schematic diagram of the electrotransport apparatus is shown in Fig. 1 and the details of the cell design are shown in Fig. 2. For the mobility experiments electrical contact with the bottom of the capillary was made with graphite, as shown in Fig. 2. In order to prevent electrical contact with the melt the graphite was coated with a thin layer of alumina using a plasma spray torch. The cell assembly, a graphite electrode, and another assembly containing two blank capillaries were fastened to the lavite disk shown in Fig. 1. To make a run the system was evacuated and the doped germanium was outgassed at the operating temperature of 1030°C. The cell and blank capillaries were then submerged beneath the melt and a slightly positive pressure of helium was introduced to fill them. After passing current through the cell for a given time the cell and blank capillaries were raised from the melt and the furnace was shut off. The temperature of the cells dropped below the freezing point of germanium within 1 to 2 min after the furnace was shut off.

Jan 1, 1968

By William Romanoff

IN AN ARTICLE on "Recovering Smelter Dust and Oxide," published in the Engineering and Mining Journal (Vol. 131, No. 2), the authors briefly described some dust-recovery equipment and its operation at a secondary-metals brass-smelting and refining plant. A description of our experience in overcoming certain operating difficulties may prove of interest or value to those who are confronted with some of the same problems or who should equip their furnaces with dust-collecting equipment. The dust loss from reverberatory smelting and refining of brass con¬sists largely of zinc, with smaller percentages of lead, tin, and copper in the order named. At the Kramer plant, the fume discharge from the two stacks amounted to 3 to 4 tons of solids per day, making a dense smoke. With dust equipment in operation, the discharge from the stacks is invisible and approximately 2500 lb. of zinc and other metals is recovered daily thereby from each furnace operating.

Jan 1, 1933