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By Ben L. Seegmiller

Roof, pillar and floor stability is a most important requirement in underground coal mines because it represents the leading cause of fatalities. In future coal mine planning the geotechnical aspects must often be minimized owing to unavailability of data. It is, therefore, very important to include a geotechnical investigation as part of the coal exploration program. Data which should be collected basically include the geological discontinuity characteristics, the rock strength characteristics and the groundwater aspects. Many times the only source of such data is the exploration boreholes. Therefore, an emphasis must be made to extract the maximum information from the core and the borehole. Data collection methods can include the preparation of detailed geotechnical logs for the roof, coal and floor. Laboratory strength testing will provide estimates of the mechanical properties of the various lithologic types. Slaking and swelling tests provide an insight into the stability characteristics of the rock materials immediately enclosing the coal seam. The boreholes may be used to provide permeability and pressurization data. Having developed a suitable geotechnical data base allows various stability prediction methods to be used. These include a geotechnical appraisal of the roof and floor units, depositional environment methods, multiple overlay techniques, hazard analysis techniques and classification/rating systems for roof and floor. The basic geotechnical data and the stability prediction methods form a basis upon which preliminary rock mechanics designs may be developed. Of particular importance to future coal mining are those designs relating to longwall mining because most companies are thinking in terms of the greater safety and greater productivity offered by longwalls. Items which should be ad- dressed when considering a longwall system include the panel layout in terms of a total mining plan, strata mechanics, roof/floor bearing capacities, face support requirements, face length, entry and pillar

Jan 1, 1983

Aamot, Olav Crone, Research Engr., Norsk Elektrokemisk Kongensgt, 18, Oslo, Norway ?29 Abbott, Clarence E., V.P., Charge of Raw Materials, Tenn. Coal, Iron & R. R. Co., 1242 Brown-Marx Bldg., Birmingham, Ala. '04 Abbott, Louis Benjamin, Chief Engr., Consolidation Coal Co., Watson Bldg., Fairmont, W. Va. '15 Abel, Raymond L., Asst. Prof., Dept. of Pet. Refin., Univ. of Pittsburgh Pittsburgh, Pa. '28 Abel, Walter D., Cons. Engr 364 S. Cloverdale Ave., Los Angeles, Calif. '32 Abele, Louis T 1457 Euclid Ave., Zanesville, Ohio. '31 Abell, Leo M Sullivan Concentrator, Chapman Camp, B. C., Canada. '31 Aborn, Robert H., Asst. Met., Research Lab., U. S. Steel Corp., Lincoln Highway, Kearny, N. J. '19 Abraham, W. E. V., East India United Service Club, 16 St. James' Sq., London, S. W. 1, England. '28

Jan 1, 1936

By C. H. Mathewson, A. U. Seybolt

As is well known, many questions affecting the properties and uses of a metal cannot be answered without careful consideration of the state of purity realized in the various operations of preparation, refining, or special treatment that have brought the product to its final state. The occurrellce of an impurity in minute quantity offers no guarantee that its effect will be inconsiderable because dissolved substances often modify the properties of a metal quite out of proportion to their concentration. In such cases the determination of solid solubility must be accepted as a problem of considerable difficulty, requiring the use of especially purified materials and methods not ordinarily used in the metallurgical laboratory. This aspect of the equilibrium between metals and their oxides is especially important and a search for fundamental data reveals little except conflicting opinion even concerning the systems of common occurrence and interest. Rhines and Mathewson1 in 1934 studied the solubility of oxygen in solid copper by annealing to equilibrium in air and gave reasons for the failure of other methods, notably the ordinary procedure of heat-treatment followed by microscopic examination, to yield reliable results. In this paper, a similar study has been made of the solubility of oxygen in cobalt. The simple theory of this method is outlined below. Conditions of Equilibrium between a Metal and Dissolved Oxide at Different Temperatures and Pressures The general conditions of equilibrium in any system are conveniently summarized in a projection of the triple curves of the P, T, X space

Jan 1, 1935

By A. U. Seybolt, C. H. Mathewson

As is well known, many questions affecting the properties and uses of a metal cannot be answered without careful consideration of the state of purity realized in the various operations of preparation, refining, or special treatment that have brought the product to its final state. The occurrellce of an impurity in minute quantity offers no guarantee that its effect will be inconsiderable because dissolved substances often modify the properties of a metal quite out of proportion to their concentration. In such cases the determination of solid solubility must be accepted as a problem of considerable difficulty, requiring the use of especially purified materials and methods not ordinarily used in the metallurgical laboratory. This aspect of the equilibrium between metals and their oxides is especially important and a search for fundamental data reveals little except conflicting opinion even concerning the systems of common occurrence and interest. Rhines and Mathewson1 in 1934 studied the solubility of oxygen in solid copper by annealing to equilibrium in air and gave reasons for the failure of other methods, notably the ordinary procedure of heat-treatment followed by microscopic examination, to yield reliable results. In this paper, a similar study has been made of the solubility of oxygen in cobalt. The simple theory of this method is outlined below. Conditions of Equilibrium between a Metal and Dissolved Oxide at Different Temperatures and Pressures The general conditions of equilibrium in any system are conveniently summarized in a projection of the triple curves of the P, T, X space

Jan 1, 1935

By Sir John Cadnan

The production from the Maidan-i-Naftun area for the year ending March, 1928, was 5,340,000 tons (40,210,200 bbl.), an increase of nearly 13 per cent. over the production for the year ending in March, 1927. The production by years since the opening of the field in 1912 is given in the following table: [1 ton = 7.53 bbl.] Year Ended MaRch Tons Year Ended March Tons 1912 (7 months)............. 43,084 1920....................... 1,354,631 1913........................ 80,800 1921....................... 1,767,070 1914........................ 273,635 1922....................... 2,310,098 1915........................ 375,977 1923....................... 2,913,908 1916........................ 449,394 1924....................... 3,583,000 1917........................ 644,074 1925....................... 4,221,000 1918........................ 897,402 1926....................... 4,470,000 1919....................... 1,081,919 1927....................... 4,735,000 1928....................... 5,340,000 The actual production represents only a small proportion of the potential production. The Masjid-i-Suleiman Field The present producing unit, the Masjid-I-Suleiman, is an anticline some 20 miles in length by 4 miles in width, including three domes as so far proved. The producing horizon is a very open, dolomitic limestone, which, while showing much variation in texture and porosity, is so connected that the lowering of pressure in one section of the field quickly affects the whole area. The topography of the field is broken, with differences of well elevations of as much as 1100 ft. The depths to the limestone vary from about 800 to about 5000 ft. The closed-in pressure of the oil wells varies from zero to 50 lb. per sq. in. and that of the gas wells is about 450 Ib. Although the static oil level may be several hundred feet below ground level, in oil wells of high elevation, all the wells flow continuously or intermittently after they are once started—the majority continuously. The gas is very high in hydrogen sulfide in gas-oil wells of low pressure.

Jan 1, 1929

By Wm. E. Milligan

DESPITE the zero weather of Monday, the morning meeting on nonferrous ore-reduction metallurgy got under way promptly under the efficient control of Arthur A. Center. The first and third portions of the paper on the Tri-State region were presented by B. M. O'Harra, with R. H. Lowe substituting for co-author Illidge on the second part. Attention was drawn to the compact design attainable when "Sink and Float" was employed as a means of primary concentration. The teaching profession should welcome this concise correlation of data from such a renowned center of lead and zinc production. H. R. Hanley and C. R. Hay¬ward were much interested in the mechanics of charging, manufacture, and life of the retorts.

Jan 1, 1943

By A. H. Daane, J. J. Hanak, F. H. Spedding

IN the study of the rare earths at the authors' laboratory, methods have been devised for preparing pure rare earths and many of their properties'- have been determined. Very little information is found in the literature on the properties of europium metal, which is the scarcest of the naturally occurring rare earths. Since we have recently prepared europium in appreciable quantities, we are in a position to describe the method of preparation and some of the properties of this metal. Metallic europium was first prepared by Klemm and Bommera who heated a mixture of potassium and EuC1, in vacuo, and obtained a mixture of europium metal and KC1. Trombe4 electrolyzed a molten salt bath containing NaC1, KC1 and EuC1, and collected the europium in a molten cadmium electrode as a cadmium alloy. The cadmium was subsequently removed by distillation but the properties of the residue indicate that pure europium was not obtained. McCoy attempted to prepare europium by distilling mercury from europium amalgam, but the stability of the Eu-Hg intermetallic compounds prevented a complete separation of the two elements. The preparation of very pure europium was suggested by previous work in this laboratory1,4 in which ytterbium and samarium were prepared by heating the respective sesquioxides with lanthanum metal turnings to 1450°C in vacuo and distilling ytterbium and samarium from the mixtures. Subsequently, small test experiments showed that europium also could be prepared by this method. A more detailed version of this method was used in the work described in this paper. Klemm and Bommer3 found by X-ray analysis that the metal had a body-centered cubic (bcc) structure with a lattice constant a, = 4.582 ± 0.00A (converted from kX units), with a bulk density of 5.30 g per cu cm. They found europium to be weakly ferromagnetic. Spedding, et al,7 found a,, to be 4.606 ± 0.001A. Barrett' showed that the bcc structure persists throughout the temperature range 5 to 3000K with a, at 5, 78, and 300°K being 4.551 ± 0.003, 4.551 ± 0.004 and 4.577 ± 0.001A, respectively. Trombe1 described europium as being iron gray, malleable, about as hard as neodymium and not attacked by cold water. He estimated the melting point to be between 1100 and 1200°C. Previous work in this laboratory indicated that the pure metal corrodes rapidly in air and reacts with cold

Jan 1, 1959

By Nathan C. Rockwood

The title of this paper, I understand, was suggested by professional mining engineers as an opportunity for someone to pose problems rather than to offer solutions for them, but the paper will merely attempt to pose some of the production problems of the mineral aggregates industry in such manner as to create professional interest in their solution by the mining engineer, geologist, geochemist and the metallurgical engineer versed in mineral-dressing technique. Definition of Mineral Aggregates " Mineral aggregates" may be defined as fragments of any rock or mineral used to supply bulk or volume or substance for a given purpose. For example, mineral aggregates are used in portland-cement concrete to provide bulk with the expenditure of the least feasible amount of cement binder. Mineral aggregates are used without port-land cement or bituminous binder in highway and railway roadbeds to provide substance, to carry loads. provide aggregates are used in filters for water and sewage, to provide substance for the liquid to trickle through. These are only a few typical uses of mineral aggregates. For all these uses, strength, toughness, resistance to abrasion, and durability are usually prime requisites; so, while theeretically mineral aggregates may be of any mineral composition, what really is desired is a mineral composition as nearly inert under the conditions of use as anything of earthly origin can be. Exploration and Possibilities of Development The distribution of deposits of mineral aggregate is universal, since almost any part of the earth,s crust can be used; even the top soil is sometimes used in stabilized surfacing for highways or airport runways. Nevertheless there is ample opportunity for exercise of the talents of the mining engineer and the geologist in finding the particular deposit that will be most economical to develop, keeping in mind such considerations as: (i) quality and mineralogical characteristics of the material; (2) size and permanency of the market; (3) final cost of opening and development of the deposit; (4) selection of location of plant; (5) methods and costs of distribution of the product in comparison with both present and possible future competition; (6) selection of the proper type of operating equipment; (7) estimation of probable waste products and their disposal. These are typical but there are other considerations also. Because profit margins are far smaller than in any comparable mining operation, it is all the more logical to assume that a higher degree of expert knowledge is desirable to make a success of a mineral-aggregates business. The industry has much engilleering talent in its operating branch, but it lacks, to a considerable degree, engineering talent with experience

Jan 1, 1948

By Nathan C. Rockwood

The title of this paper, I understand, was suggested by professional mining engineers as an opportunity for someone to pose problems rather than to offer solutions for them, but the paper will merely attempt to pose some of the production problems of the mineral aggregates industry in such manner as to create professional interest in their solution by the mining engineer, geologist, geochemist and the metallurgical engineer versed in mineral-dressing technique. Definition of Mineral Aggregates " Mineral aggregates" may be defined as fragments of any rock or mineral used to supply bulk or volume or substance for a given purpose. For example, mineral aggregates are used in portland-cement concrete to provide bulk with the expenditure of the least feasible amount of cement binder. Mineral aggregates are used without port-land cement or bituminous binder in highway and railway roadbeds to provide substance, to carry loads. provide aggregates are used in filters for water and sewage, to provide substance for the liquid to trickle through. These are only a few typical uses of mineral aggregates. For all these uses, strength, toughness, resistance to abrasion, and durability are usually prime requisites; so, while theeretically mineral aggregates may be of any mineral composition, what really is desired is a mineral composition as nearly inert under the conditions of use as anything of earthly origin can be. Exploration and Possibilities of Development The distribution of deposits of mineral aggregate is universal, since almost any part of the earth,s crust can be used; even the top soil is sometimes used in stabilized surfacing for highways or airport runways. Nevertheless there is ample opportunity for exercise of the talents of the mining engineer and the geologist in finding the particular deposit that will be most economical to develop, keeping in mind such considerations as: (i) quality and mineralogical characteristics of the material; (2) size and permanency of the market; (3) final cost of opening and development of the deposit; (4) selection of location of plant; (5) methods and costs of distribution of the product in comparison with both present and possible future competition; (6) selection of the proper type of operating equipment; (7) estimation of probable waste products and their disposal. These are typical but there are other considerations also. Because profit margins are far smaller than in any comparable mining operation, it is all the more logical to assume that a higher degree of expert knowledge is desirable to make a success of a mineral-aggregates business. The industry has much engilleering talent in its operating branch, but it lacks, to a considerable degree, engineering talent with experience

Jan 1, 1948

By Ralph Arnold

Since 1915, when Elk Basin field was brought in, eight oil fields have been developed in Montana and the production has arisen from 50,000 bbl. in 1916 to 8,000,000 bbl. in 1926. The production for 1927 was 5,121,600 bbl. and the total up to the end of 1927 was 27,138,239 barrels. The eight fields, with their year of discovery are: Elk Basin, Carbon County (1915); Devil's Basin, Petroleum County (1919); Cat Creek, Petroleum County (1920); Soap Creek, Big Horn County (1921); Kevin-Sunburst, Toole County (1922); Lake Basin, Stillwater County (1924); Pondera, Pondera and Teton counties (1927); and Bannatyne, Pondera and Teton counties (1927). About 170 wildcat wells have been drilled in the state of which 8 were successful and 162 gas wells or dry. Hence, only one out of 21, or 5 per cent. of the wildcats have been successful. The average depth of the wells has been 1600 ft. and the average cost $32,000 apiece. • Up to Jan. 1, 1927, a total of 1376 wells have been drilled in the producing oil and gas fields of the state, of which 904, or 66 per cent., were successful oil wells; 82, or 6 per cent. were gas; and 390, or 28 per cent, were dry. The percentage of dry wells is excessive as compared to the wells in other states, due to the uncertainties of limestone reservoir rocks. The total production of oil to Jan. 1, 1927, has been 22,016,639 bbl. and of gas 7,265,330,000 cu. ft. The average price at the well for the oil has been $1.47 per bbl. and for gas about 4 c. per 1000 cu. ft. The value at the point of consumption has been about 23 c. per 1000 cu. ft. The total revenue derived from crude oil has been $32,277,119, and from gas 8350,613, or a total income to the producer of $32,727,832. Each well drilled in the state has already (to Jan. 1, 1927) averaged a return of over $21,000. The total number of domestic (Montana) oil and gas companies organized, since the first in 1885, has been 844, of which 702 are still alive and 142 dead. Oil and gas companies registered for operation within the state to Jan. 1, 1927, but with outside charters, are 143, of which 66 are alive and 77 are dead. Outside the few distributing companies, which are included in these figures, the grand total of 997 companies which have

Jan 1, 1928

By E. L. Kinsella, F. S. White

Solids fluidization utilized in two ways for the commercial production of lime is described. Crushed —6 mesh limestone is dried and dedusted in a single bed reactor, then calcined in a 5-stage reactor. Construction, operation, and results obtained with both processes are given. THE solids fluidization process brought out by the Standard Oil Development Co. in the early forties for catalytic cracking of petroleum enabled rapid transfer of large quantities of heat from gases to solids under closely controlled temperature conditions. As the technique developed, it was natural that applications for the process outside the petroleum industry would be sought. The calcination of limestone to quicklime requires the transfer of a large quantity of heat at high and closely controlled temperatures. For these reasons, this possible application was one of the first to be investigated. Laboratory work conducted by the Dorr Co. indicated that lime produced by solids fluidization would be of exceptional quality. Calculation of theoretical heat balances showed that savings in fuel could be expected. Accordingly, a pilot plant of 10 tons per day capacity was erected by the New England Lime Co. at Adams, Mass., to substantiate the process. Operation of the pilot plant, previously described,' showed that excellent lime could be produced with a saving in fuel. The work also indicated the necessity of preparing feed to remove the finer fractions if dust losses were to be held to a tolerable level. A single-stage fluidized solids drier-stripper was developed to accomplish this sizing.' Subsequently a commercial lime reactor was erected at Adams, beginning operation in June 1949, followed by air sizer facilities completed in November 1951. This paper will describe these two applications of solids fluidization. The raw material for preparation of lime by the fluidizing process is a high calcium crystalline lime- stone having the approximate chemical composition shown in Table I. The stone is first prepared by crushing and grinding in conventional equipment to pass a 6-mesh sieve. After crushing, the material is sent to the FluoDry unit for drying and dedusting. Fluidized Drying and Sizing As shown in Fig. 1, the FluoDry unit consists of a -in. cylindrical steel shell, 9-ft ID by 22-ft 6-in. overall height. Attached horizontallv at the bottom is a 6-ft diam combustion chamber, approximately 7 ft 6 in. long. A refractory constriction dome, consisting of first-quality fire brick shapes, partitions the inside into two sections. The lower section is called the windbox and the upper the bed compartment. The vertical wall of the shell is lined with 6-in. first-quality rotary kiln blocks. In the bottom of the windbox a hoppered section is formed with castable refractory to permit the windbox to be readily cleaned via a 6-in. screw conveyor. The combustion chamber is similarly lined with refractory shapes. On the 6-in. diam section, 4 in. of insulating brick, HW-26, are used. The inner combustion chamber is lined with 4-in. Korundal brick to withstand the high flame temperatures encountered. An 8-in. steel feed pipe, entering vertically through the top and fitted with an air-operated cone valve,

Jan 1, 1953

By F. S. White, E. L. Kinsella

Solids fluidization utilized in two ways for the commercial production of lime is described. Crushed —6 mesh limestone is dried and dedusted in a single bed reactor, then calcined in a 5-stage reactor. Construction, operation, and results obtained with both processes are given. THE solids fluidization process brought out by the Standard Oil Development Co. in the early forties for catalytic cracking of petroleum enabled rapid transfer of large quantities of heat from gases to solids under closely controlled temperature conditions. As the technique developed, it was natural that applications for the process outside the petroleum industry would be sought. The calcination of limestone to quicklime requires the transfer of a large quantity of heat at high and closely controlled temperatures. For these reasons, this possible application was one of the first to be investigated. Laboratory work conducted by the Dorr Co. indicated that lime produced by solids fluidization would be of exceptional quality. Calculation of theoretical heat balances showed that savings in fuel could be expected. Accordingly, a pilot plant of 10 tons per day capacity was erected by the New England Lime Co. at Adams, Mass., to substantiate the process. Operation of the pilot plant, previously described,' showed that excellent lime could be produced with a saving in fuel. The work also indicated the necessity of preparing feed to remove the finer fractions if dust losses were to be held to a tolerable level. A single-stage fluidized solids drier-stripper was developed to accomplish this sizing.' Subsequently a commercial lime reactor was erected at Adams, beginning operation in June 1949, followed by air sizer facilities completed in November 1951. This paper will describe these two applications of solids fluidization. The raw material for preparation of lime by the fluidizing process is a high calcium crystalline lime- stone having the approximate chemical composition shown in Table I. The stone is first prepared by crushing and grinding in conventional equipment to pass a 6-mesh sieve. After crushing, the material is sent to the FluoDry unit for drying and dedusting. Fluidized Drying and Sizing As shown in Fig. 1, the FluoDry unit consists of a -in. cylindrical steel shell, 9-ft ID by 22-ft 6-in. overall height. Attached horizontallv at the bottom is a 6-ft diam combustion chamber, approximately 7 ft 6 in. long. A refractory constriction dome, consisting of first-quality fire brick shapes, partitions the inside into two sections. The lower section is called the windbox and the upper the bed compartment. The vertical wall of the shell is lined with 6-in. first-quality rotary kiln blocks. In the bottom of the windbox a hoppered section is formed with castable refractory to permit the windbox to be readily cleaned via a 6-in. screw conveyor. The combustion chamber is similarly lined with refractory shapes. On the 6-in. diam section, 4 in. of insulating brick, HW-26, are used. The inner combustion chamber is lined with 4-in. Korundal brick to withstand the high flame temperatures encountered. An 8-in. steel feed pipe, entering vertically through the top and fitted with an air-operated cone valve,

Jan 1, 1953

By J. A. Berger, R. J. Towner

A method for determining the disorientation between sub-grains that .share a common tilt or twist boundary from measurements on X-ray micrographs is described. The method may be applied to subgrains of almost any orientation having boundary angles in the range from about 1 min to 1 deg of arc. An adaption of the Berg-Barretl method has been made to allow calculation of small angles of about 1 min to 1 deg of arc: between adjacent subgrains from the relative displacement of their images on X-ray micrographs. schulz4 gave equations for determining subgrain disorientation where the axis of subgrain rotation and the reflecting crystal planes were parallel to the specimen surface. In our method, however, they need not be parallel and, indeed, may have va:rious orientations. His camera setup, involving a fine-focus X-ray tube for obtaining high resolution with continuous radiation, was different from the one used in the present investigation (a 1.2 x 1.2 mm effective source and characteristic radiation). The X-rays produced a slight magnification in the subgrain images in the Schulz method, but not in the present method. The commonly available type of X-ray tube used permitted relatively short exposure times with a fine-grained photographic plate, which could be magnified optically many times for the examination of small subgrains. The method is illustrated by results obtained on {110} tilt boundaries in a polygonized aluminum single crystal of 99.994 pct purity. The flat crystal was stretched 6 pct in tension and then heated at 500°C for a cumulative time of 5 hr. Earlier, cahn5 studied subgrains formed by polygonization in stretched and heated aluminum crystals. He observed kink bands to have well defined boundaries lying in (110) planes, at least in the early stages of deformation. Cahn found that the (110) tilt boundaries formed by polygonization were directly related to the lattice bending about the <211> axis which lay in the (111) slip plane perpendicular to the <110> slip direction. DESCRIPTION OF THE CAMERA Our camera was similar in detail to Honey-combe&apos;s.6 Unfiltered copper radiation was used at 35 kv, and 20 ma. With the particular specimen used here to illustrate the method, a 1.2 by 1.2 mm effective source and a distance of 780 mm from target to specimen were employed, giving a convergence of 5 min of arc in the primary beam. The distance was 9.5 mm from the specimen to film. The upper 1/2- by 1-in. portion of the front surface of the specimen was irradiated with the film in parallel position. Since the specimen had been inhomogeneously deformed, many regions were in a position to reflect characteristic radiation at the Bragg Angle. Five minute exposures were satisfactory with Kodak Single Coated X-ray Film, 1 1/2 hr

Jan 1, 1961

By R. B. Moore

IN 1914 the writer and K. L. Kithil announced, through Bulletin 70 of the U. S. Bureau of Mines, that the United States possessed the largest deposits of radium-bearing ore in the world. At that time a small amount of this ore was being, mined and shipped to Europe, and one company was preparing to start the extraction and recovery of radium in this country. The importance of our carnotite deposits in Colorado and Utah up to the publication of Bulletin 70 was not, however, appreciated by anyone.. At that time the total supply of radium in the world was about 10 g., calculated as element, and when I estimated that at least 100 g. of radium element, and probably 200 g., would be produced from our carnotite deposits, there was a considerable amount of skepticism. Up to the present time almost 200 g. have come from this ore, and some writers put this figure a little higher. Uranium ore was found in the copper mines of the Union Minière du Haut-Katanga at Luiswishi in 1913. Later discoveries were made at other places in the Belgian Congo, and the Katanga ores now represent the largest supply of radium-bearing ore in the world. In fact, since 1923 very little radium has been produced from American carnotite, the Katanga ore representing practically the world&apos;s supply of. radium since that date.

Jan 1, 1930

By J. A. Reeves

The unit-train for moving mineral raw materials is considered by many railroad men as the most significant contribution toward reducing rail freight costs since the introduction of the diesel locomotive. The idea, which involves assigning railroad equipment to only one haul and keeping the equipment rolling, results in the maximum utilization of equipment. For example, before the unit train was placed in service on coal hauled to one of Mid-Continent Coal and Coke Company&apos;s customers, about 900 cars and 9 locomotives were needed to move its coal requirements, compared with 250 cars and 6 locomotives currently used. If unit-train economics are to be achieved, coal companies cannot use railroad cars as small storage bins on wheels to accumulate their coal in train-load shipments, instead, these companies must build large bins which incorporate facilities for the rapid loading of railroad cars in a minimum length of time.

Jan 1, 1970

By F. W. Warnaars

The Los Bronces copper deposit is located on the west side of the Andes Mountains in central Chile about 54 km (34 miles) northeast of Santiago. The deposit consists of a hydrothermal breccia complex that was emplaced in the western part of an earlier porphyry copper mineral system. This system exhibits propylitic, sericitic, silicic to potassic hydrothermal alteration with disseminated and stockwork copper-molybdenum sulfide mineralization over an area of about 12 km2 (5 sq. miles). The Rio Blanco Mine, about 800 m (0.5 miles) east of Los Bronces, is an underground block caving mine located in the central part of this system. The mineralization at Rio Blanco is characterized by one of the highest primary copper grades known for porphyry copper systems. Present mining at Los Bronces is from an open pit located in the northern part of a hydrothermal complex that is composed of at least seven different copper-bearing tourmaline breccia pipes. Together these pipe form one large contiguous body about 2 km (1.2 miles) long and 0.7 km (0.4 miles) wide at the surface. This body crops out at elevations of between 3.5 and 4.1 km (2 and 2.5 miles) above sea level. Each breccia pipe has a characteristic shape, matrix, type and degree of mineralization and alteration. Quartz, tourmaline, specularite, anhydrite, pyrite, chalcopyrite, sericite, chlorite, and rock flour are present in significant quantities in most breccia matrices. The clasts consist mainly of quartz monzonite, andesite, and minor amounts of quartz latite porphyry. The seven different breccia pipes are identified, from oldest to youngest, as Ghost, Central, Western, Infiernillo, Fine Gray, Anhydrite, and Donoso. The breccias coalesce, interfinger, or display gradational contacts. In places they form sharp contacts with each other and with the surrounding intrusive or volcanic host rocks. Early breccias are disrupted or cut by subsequent breccia events. Explosive emplacement was followed by collapse after the release of hydrothermal pressures. The Los Bronces breccia complex has sharp contacts with the surrounding intrusive stock and andesites. These contacts dip inward on the north, west, and south margins. The eastern contact is nearly vertical and/or dips slightly outward and suggests a westward tilting of the body of about 15° after emplacement. The breccia complex is not fully explored at depth, but is known to extend at least below the 3 km (2 mile) elevation in its central eastern part, about 1 km (0.6 miles) below the highest breccia outcrop. Primary copper sulfide minerals are most abundant in the Donoso and Infiernillo breccias where the matrix consists mostly of tourmaline, quartz, sulfides, specularite, and/or chlorite with minor amounts of rock flour, sericite, and clay. The volume of the matrix material is mostly smaller than in the other breccias. Secondary enrichment enhanced the primary copper grade in the southern two-thirds of the Los Bronces breccia complex and in much of the surrounding porphyry system. Primary grades of 0.3% to 0.5% copper have been enriched to approximately 0.9%. This enrichment may extend from surface outcrops to a depth of more than 500 m (1, 640 ft) in more permeable parts of the breccia complex. A leached and oxidized capping of variable thickness covers a large part of the enriched layer and exceeds 300 m (984 ft) at one location. The shape and depth of the enrichment blanket and the capping suggest that the enrichment process is still active.

Jan 1, 1983

By Bernard Kopelman

The particle size of tungsten metal powder used to make tungsten wire for use in radio tubes and incandescent lamps must be closely controlled if the highly desirable feature of nonsagging is to be achieved. Our knowledge of factors affecting the particle size of tungsten powder has been limited by lack of any extensive literature on the subject. There are no publications in the English language discussing this problem in any but the most general terms. In this paper the published experimental information on this subject is reviewed. Based on the published data and some experimental data obtained by the author, a general theory of factors affecting the particle size of tungsten metal powder obtained by reduction of the oxides in hydrogen is offered. The various forms of the tungsten oxides and tungsten metal used in the experiments described were those prepared by commercial processes, so that the experimental results obtained on a laboratory scale would be more directly applicable to actual commercial reduction conditions. The effects of time of reduction, temperature. and concentration of water vapor in the hydrogen were observed by counting microscopically the particle distribution obtained. Such a method is very time-consuming, but the data obtained by this method are more reliable and more fundamental than those obtained by any of the several instruments available for measurement of particle size. Review of the Literature Until the year 1912, tungsten powder was prepared commercially by reducing tungstic oxide in the presence of carbon or one of its compounds (such as glycerin or ethylene glycol), usually by firing the mixture in a closed fire-clay crucible. The resultant powder contained about 95 per cent tungsten.&apos; When reduction of tungsten oxides by H² was introduced Commercially,²,³ several investigators became concerned with the equilibria between tungsten metal and its several oxides in an atmosphere of hydrogen and water vapor.4-l3 While these many investigations were in fair agreement on the equilibria WO³ + H2 = WO2 + H2O and WO2 + 2H2 = W + 2H2O there arose considerable controversy as to the chemical composition of the oxide between WO³ and WO², because of very poor agreement on the equilibrium constant of this intermediate oxide reduction reaction. The controversy on this intermediate oxide has been in progress for more than 100 years. Thus Malguti,14 in 1835, decided it was WO2.56 In 1857, Riche15 proposed the formula W³O8. Usler,16 in 1865, confirmed the existence of WsO8 and further proposed W1O1l. In 1897, Desil7 agreed upon W²O8 but also found W5O14.

Jan 1, 1947

By Bernard Kopelman

The particle size of tungsten metal powder used to make tungsten wire for use in radio tubes and incandescent lamps must be closely controlled if the highly desirable feature of nonsagging is to be achieved. Our knowledge of factors affecting the particle size of tungsten powder has been limited by lack of any extensive literature on the subject. There are no publications in the English language discussing this problem in any but the most general terms. In this paper the published experimental information on this subject is reviewed. Based on the published data and some experimental data obtained by the author, a general theory of factors affecting the particle size of tungsten metal powder obtained by reduction of the oxides in hydrogen is offered. The various forms of the tungsten oxides and tungsten metal used in the experiments described were those prepared by commercial processes, so that the experimental results obtained on a laboratory scale would be more directly applicable to actual commercial reduction conditions. The effects of time of reduction, temperature. and concentration of water vapor in the hydrogen were observed by counting microscopically the particle distribution obtained. Such a method is very time-consuming, but the data obtained by this method are more reliable and more fundamental than those obtained by any of the several instruments available for measurement of particle size. Review of the Literature Until the year 1912, tungsten powder was prepared commercially by reducing tungstic oxide in the presence of carbon or one of its compounds (such as glycerin or ethylene glycol), usually by firing the mixture in a closed fire-clay crucible. The resultant powder contained about 95 per cent tungsten.&apos; When reduction of tungsten oxides by H² was introduced Commercially,²,³ several investigators became concerned with the equilibria between tungsten metal and its several oxides in an atmosphere of hydrogen and water vapor.4-l3 While these many investigations were in fair agreement on the equilibria WO³ + H2 = WO2 + H2O and WO2 + 2H2 = W + 2H2O there arose considerable controversy as to the chemical composition of the oxide between WO³ and WO², because of very poor agreement on the equilibrium constant of this intermediate oxide reduction reaction. The controversy on this intermediate oxide has been in progress for more than 100 years. Thus Malguti,14 in 1835, decided it was WO2.56 In 1857, Riche15 proposed the formula W³O8. Usler,16 in 1865, confirmed the existence of WsO8 and further proposed W1O1l. In 1897, Desil7 agreed upon W²O8 but also found W5O14.

Jan 1, 1947

By Pierre R. Hines

Diphenyl guanidine is used as an accelerator in vulcanizing rubber. Other rubber accelerators are also flotation collectors, e.g., dithiocarbamate, thiazole, and the xanthates. Urea and its derivatives are good flotation collectors,1 and guanidine is the nitrogen analog of urea. These two characteristics suggested testing diphenyl guanidine as a flotation collector. Diphenyl guanidine has been tried in flotation work previously, but references give no details.2-4

Jan 2, 1959

By John R. Don

In the South Island of New Zealand an unusually favorable opportunity is offered for the analysis of the older crystalline rocks, underlying the sedimentary rocks which form the "country" of the gold-deposits. The Manipori formation of this island, comprising the greater part of the mountainous district west of lakes Manipori and Te Anau, in Otago, consists of an enormous thickness (estimated by Professor Hutton1 at 160,000 ft.) of crystalline schists, gneiss, syenite, and syenitic gneiss, with associated masses of granite; the whole forming the most picturesque fart of New Zealand. The famous West Coast sounds occur in this formation. No paying auriferous lodes have been discovered in it; and for this reason an examination of the rocks is specially interesting here. Stelzner, Posepny, and others, who have criticised the conclusions of Sandberger, have laid great stress on the fact that all the silicates analyzed by him were taken from the vicinity of orebodies containing those heavy metals which he notes as occurring in the silicates of the country-rock; their contention being that these metals, supposed to occur as silicates in gneiss and other crystalline rocks, were really contained as sulphides, and were therefore impregnations from the neighboring orebodies. The same objection, whatever be its weight, might be urged with equal justice against Mr. Becker&apos;s derivation of the mercury and gold of the lodes of the Pacific Coast from&apos; the granite underlying their country-rock. In the case of the

Jan 1, 1913