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By Thomas H. Clagett

ThE Pocahontas coal field comprises the area in Tazewell County, Va. and Mcrcer and McDowell counties, W. Va., in which Nos. 3 and 4, Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305 111.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, to which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks of bone, cach from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 2.5 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficieiltly well defined to exert an influence on the mining is rare. Overlying strata arc slates, shales, and sandstones, the sandstones predominating, extending from 30 to 1000 ft. above No. 3 seam. The strata directly above the draw slate of No. 3 seam, or "black rash" of No. 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the.

Jan 1, 1923

By Thomas H. Clagett

ThE Pocahontas coal field comprises the area in Tazewell County, Va. and Mcrcer and McDowell counties, W. Va., in which Nos. 3 and 4, Pocahontas seams of bituminous coal are mined. It is a mountainous region with ridges rising 300 to 1000 ft. (91 to 305 111.) above the very narrow valleys and, except for some poorly kept farms along the tops of the ridges and the area occupied by the mining and coking plants and villages, is wooded throughout. First shipments were made from Pocahontas in 1883. There are now about 100 mines, 95 per cent. of which are drift mines. At the end of 1920, approximately 298,000,000 gross tons of coal had been produced and 35,000 acres of coal mined. No. 3 seam is from 4 to 10 ft. (1.2 to 3 m.) thick, but the change is gradual and the thickness fairly uniform in any particular locality. It has a streak of bone, about 2 in. (5 cm.) thick, to which the coal adheres on both sides. In parts of the field this bone is replaced by hard slate, usually from 2 to 4 in. thick. The seam has a slate or fireclay bottom and a draw-slate top. No. 4 seam is from 65 to 80 ft. (19.8 to 24 m.) above No. 3 seam, and of minable section only in the western part of the field. It is from 3 to 8 ft. thick and has two streaks of bone, cach from 1 to 2 in. thick. It has a fireclay or slate bottom and is overlaid by from 4 to 24 in. of a laminated coal and slate, locally termed "black rash," which is separated from the coal seam by about 3 in. of slate. This "black rash" averages about 2.5 per cent. ash, but contains some clean coal. The coal in both seams is quite soft. Cleavage sufficieiltly well defined to exert an influence on the mining is rare. Overlying strata arc slates, shales, and sandstones, the sandstones predominating, extending from 30 to 1000 ft. above No. 3 seam. The strata directly above the draw slate of No. 3 seam, or "black rash" of No. 4 seam, and up to the first sandstone are usually slates or shales, about 10 ft. thick over the.

Jan 1, 1923

By E. Miller, J. M. Eldridge, K. L. Komarek

The liquidus curve of the Mg-Pb system was accurately redetermined. The compound Mg2Pb decomposes peritectically at 538.2° ± 0.3°C to liquid and to a compound p' which melts congruently at 35.0 at. pct Pb and 549.0° ± 0.3°C. The solidus curve of ß' was determined. X-ray diffraction studies indicate that 4' has an orthorhombic structure. Activity values of magnesium calculated from the phase diagram agree with those published in the literature. EXPERIMENTAL thermodynamic properties of binary metallic systems have to be consistent with values calculated from the phase diagram. In systems forming intermetallic compounds the shape of the liquidus curve near a compound is determined by the thermodynamic properties of the coexisting solid and liquid phases. Hauffe and Wagner' neglected the temperature dependence of the chemical potentials and obtained the potential differences of the components of the liquid alloys, relative to stoichiometric liquid. Their calculations were based on the liquidus curve and on the heat of fusion of the compound, and were only valid near the congruent melting point. Steiner, Miller, and Komarek2 developed equations which account for the temperature dependence and obtained the chemical potentials of liquid Mg-Sn alloys over the entire phase diagram from the liquidus and solidus curves and from enthalpy values with the pure components as the standard states. The Mg-Pb phase diagram has been studied by several investigators whose results have been compiled and critically evaluated by Hansen.3 Although the liquidus curve was poorly defined, the general features of the diagram, i.e., one congruent melting compound, Mg2Pb, of essentially stoichiometric composition, two eutectics, and limited terminal solid solubilities, seemed to be suitable for a similar thermodynamic analysis. A careful redeter-mination of the liquidus by thermal analysis revealed, however, the existence of another compound. The liquidus curve between the two eutectics was precisely delineated and the structure and solidus curve of the new compound were investigated. The revised phase diagram was thermodynamic ally analyzed to evaluate the activity of magnesium in the liquid alloys. EXPERIMENTAL PROCEDURE The magnesium metal (Dominion Magnesium Ltd., Toronto, Canada) had a purity of 99.99+ pct; lead (American Smelting and Refining Co.) contained 99.999 pct Pb. Most experiments were carried out in graphite crucibles. Several experiments were made in high-purity alumina (Triangle R.R., Mor-ganite, Inc.) and in Armco iron crucibles to test the inertness of the graphite crucibles. Chemical analysis of magnesium and detailed description of the procedure for thermal analysis have been given previously. For the determination of the solidus curve of the compounds, specimens of initial composition Mg2Pb were equilibrated in a closed isothermal system with magnesium vapor. The source of the magnesium vapor was an alloy which had a gross composition lying in the 0' + L field at the temperature of equilibration. As equilibrium was approached, the specimens lost magnesium to the two-phase reservoir thereby lowering the activity of magnesium in the specimens until activity and composition equaled that of the ß'/ß' + L boundary. Crucibles (1.9 cm ID by 2.2 cm OD by 4.1 cm high) and tightly fitting lids were machined from a molybdenum rod; small, shallow trays were fashioned from thin (0.005 in.) molybdenum sheet, and all the molybdenum components were degreased in hot carbon tetrachloride and then dried. The pieces were then degassed in vacuum at 950°C for about 6 hr. The two-phase alloy was placed at the bottom of the crucible and small specimens of the Mg2Pb compound, weighed on an analytical balance, were placed in two molybdenum trays above the two-phase alloy. The crucible was closed by forcing its lid on and then inserted in a titanium crucible. This crucible was evacuated, flushed twice with argon, and welded under argon. The specimens were equilibrated for about 1 week in a resistance furnace regulated by a Celectray controller, and the runs were terminated by water quenching. The specimens were again weighed and the equilibrium compositions were calculated on the basis that the weight losses were solely due to a loss of magnesium to the two-phase alloy. The structure of the B' phase was investigated by the Debye-Scherrer X-ray diffraction technique. Selected ingots from thermal-analysis experiments containing about 35 at. pct Pb were re-melted, slowly cooled, and crushed in an argon-filled glovebox until the entire ingot passed through a 50-mesh sieve. The powder was thoroughly

Jan 1, 1965

By A. R. Orban, R. B. Engdahl, J. D. Hummell

The initial phase of a research program on smoke abatement from Bessemer converters is described. In work sponsored by the American Iron and Steel Institute, a 300-lb experimental Bessemer converter was assembled to simulate blowing conditions in a commercial vessel. Measurements of smoke and dust were also made in the field on a 30-ton commercial vessel. During normal blows the dust loading from the laboratory converter averaged 0.51 lb per 1000 lb of exhaust gas. This was similar to the exhaust-gas loading of a commercial vessel. The addition of hydrogen to the blast gas of the laboratory converter caused a decided decrease in smoke density. Smoke was also reduced markedly when methane or ammonia was added instead of hydrogen. The research is continuing on a bench-scale investigation of the mechanism of smoke formation in the converter process. DURING the past 2 years, on behalf of the American Iron and Steel Institute, Battelle has been conducting a research program on the control of emissions from pneumatic steelmaking processes. The objective of the research program is to discover a practical method for reducing to an unobjectionable level the emission of smoke and dust from Bessemer converters. PRELIMINARY INVESTIGATION Although conceivably some new collecting technique may be devised which would be economically practicable for cleaning Bessemer gases, no such system based on presently known principles seems feasible because of the extremely large volume of high-temperature gases involved. Hence, the research is being directed toward prevention of smoke formation at the source. A thorough review was first made of former work to determine the present status of the cleaning of converter gases. No published work was found on work done in the United States on collecting smoke or on preventing its formation in the bottom-blown, acid-Bessemer converter. In Europe, however, a number of investigations have been made on the basic-Bessemer converter. Kosmider, Neuhaus, and Kratzenstein1 conducted tests on a 20-ton converter to obtain characteristic data for dust removal and the utilization of waste heat. They concluded that because of the submicron size of the dust, special equipment would be necessary to clean the exhaust gases. Dehne2 conducted a large number of smoke-abatement experiments at Duisburg-Huckingen in a 36-ton Thomas converter discharging into a stack. A number of wet-scrubbing and dry collectors were tried unsuccessfully. A waste-heat boiler and electrostatic collector with necessary gas precleaners was felt to be the best solution for this particular plant. Meldau and Laufhutte3 determined that the particle size was all below 1 µ in the waste gas of a bottom-blown converter. Sel'kin and zadalya4 describe the use of oxygen-water mixtures injected into a molten bath in refining open-hearth steel. They claim that with use of oxygen-water mixtures the amount of dust formed was reduced between 33.3 and 20 pct of its previous level, and emission of brown smoke almost ceased. Pepperhoff and passov5 attempted unsuccessfully to find some correlation between the optical absorption of the smoke, the flame emission, and the composition of the metal in a Thomas converter in order to determine automatically the metallurgical state in the melt. In a recent U. S. Patent (NO. 2,831,762)' issued to two Austrian inventors, Kemmetmuller and Rinesch, the inventors claim a process for treating the exhaust gases from a converter. By their method the inventors claim that the exhaust gases from the converter are cooled immediately after leaving the converter to a degree that oxidation of the metal vapors and metal particles to form Fe2O3 is inhibited in the presence of surplus oxygen. Gledhill, Carnall, and sargent7 report on cleaning the gases from oxygen lancing of pig iron in the ladle. They claim the Pease-Anthony Venturi scrubber removed 99.5 + pct of the smoke, thereby reducing the concentration to 0.1 to 0.2 grain per cu ft, which resulted in a colorless stack gas after the evaporation of water. Fischer and wahlster8 developed a small basic converter and compared the metallurgical behavior of the blow with that of a large converter. Later work by Kosmider, Neuhaus, and Hardt9 on the use of steam for reduction of smoke from an oxygen-enriched converter confirmed that the cooling effect of steam is detrimental to production. From review of all of the published information on the subject, it was concluded that a practical solution to the smoke-elimination problem had not been found. Accordingly, it was deemed desirable to investigate the feasibility of preventing the initial formation of smoke in the converter.

Jan 1, 1961

All the Technical PUBlications and Preprints published in 1932 are available at Institute headquarters. They are also on file in public, university and technical libraries and, when so indicated in the list, may be found in the Transactions. Copies of the Technical Publications have been mailed to those members of the Institute who are registered in the different classes: A-Metal Mining; B-Milling and Concentration; C-Iron and Steel Division; D-Nonferrous Metallurgy; E-Insti-tute of Metals Division; F-Coal Division; G-Petroleum Division; H-Nonmetallic Minerals; I-Mining Geology; K-Mining Administration; L-Geophysical Prospecting. Other members may obtain copies by writing to the Secretary. Copies of Preprints may also be obtained from the Secretary's office. Additional copies are sold at the rate of 1 cent per page; minimum charge, 25 cents. Nonmembers may purchase the papers at the prices mentioned. Metal Mining Technical Publi- Title Author Meeting Trans. cation Volume No. and Page* 484-A.45 Propeller Type Mine Fan at Moose A. S. Richardson New York 102, 88. 95 F.52 Shaft. Butte. Montana (13) Feb.. 1932 487-A.46 Progress in the Improvement of Max H. Barber New York. Methods and Equipment at Feb., 1932 Open-pit Iron Mines on the Lake Superior Iron Ranges (9) Pre. Sand Filling through Pipes and Lucien Eaton New York, 102, 33, 41 Boreholes (9) Feb., 1932 Milling And Concentration 445-B.37 Soap Flotation of the Nomulfides Will H. Coghill New York. H.19 (18) J. Bruce Clemmer Feb., 1932 461-B.38 Principles of Flotation—An Experi- I. W. Wark New York. mental Study of the Effect of A. B. Cox Feb.. 1932 Xanthates on Contact Angles at Mineral Surfaces (48) Iron and Steel Division 447-C.81 Electrochemical Potentials of Nitri- Shun-ichi Satoh New York, fied Steels (31) Feb.. 1932 450-C.82 Reaistanoe of Iron-aluminum Alloya N. 9. Ziegler New York. 100, 267. 670 E.143 to Oxidation at High Tempera- Feb., 1932 tures (8) 451-C.83 Pressure Welding of Low-carbon C. H. Austin New York. Steela with Theoretical Con- W. S. Jeffries Feb.. 1932 siderations on the Mechanism of Such Welding (42) 453-C.84 Inclusions—Tbeir Effect, Solubility C.E.Sims New York, 100, 154, 176 and Control in Cast Steel (24) 0. A. Lillieqvist Feb., 1932

Jan 1, 1932

By Robert H. Richards

(Read at the Amenia Meeting, October, 1877.) DURING the winter of 1868-9, I was called upon by Professor F. H. Storer, to put up the Bunsen filter pump in the chemical laboratory of the Masschusetts Institute of Technology. As the laboratory is on the lower floor, it was found necessary to use a great deal of cumbrous pipe and apparatus to get the necessary column. It was then suggested that if an instrument could be devised which would make use of the hydrant pressure in the service-pipe, it would do away with the objectionable features of the Bunsen pump. After using the Bunsen pump for some years, the jet pump to be described in this paper was devised, and it has since been adopted by Prof. Wing in the quantitative laboratory of the school. I am much indebted to Prof. Wing for his aid in starting the manufacture in brass of these jet pumps, which are now to be had in Boston. The jet pump much resembles the Giffard injector in form. It differs from it in the fact that water is the impelling fluid, and air the impelled, while with the injector steam is used to impel water. The principle of condensation of steam, which is Water availed of in the injector, is entirely wanting in the jet pump. The jet pump consists in a water jet, w, (see figure), a constriction or waist, a, and a waste tube, o. The success of the jet pump depends on the following conditions 1. The relation between the sizes of sectional area of a and w. 2. The proximity of a and w. 3. The form and angle of the two hollow cones whose vertices make the constriction a. 4. The relative size of sectional area of w and o. 5. The zigzags, or an equivalent means of making foam. In the following discussion, let a, w, and o represent the diameters of these parts of the jet pump. A number of experiments were tried to ascertain

Jan 1, 1878

By R. C. Canby

The American water-jacket furnace is the outgrowth of lead-smelting at Eureka, Nev., subsequently developed in Utah and Colorado. Early smelting in Virginia, New England, or the Missouri-Kansas-Illinois district seems to have had no direct influence in the development of this furnace. First Decade, 1870 to 1880. In 1870, the draft-furnace was used in Nevada, coincidentally with the first crude sandstone blast furnaces, which were 2 or 3 ft. sq., with Sturtevant fans or small Roots blowers, and sheet-iron or water-cooled tuyeres. Stetefeldt, Liebenau, Albert Arents, and 0. H. Hahn mere the early metallurgical leaders. Arents's siphon-tap and the re-establishment of the furnace-bosh mere the most important improvements of the decade. Second Decade, 1880 to 1890. At first there were ,many small local plants; but by the time Eureka had reached its maximum production of lead-ore, two custom works (the Eureka and the Richmond) were treating all the ore, the miners having learned the economy of smelting at larger and well-constructed plants. In the early eighties, most of the features of modern practice were already discernible at Salt Lake, Leadville, Denver, and Pueblo. The small square sandstone furnaces had been replaced by rectangular water-jacket furnaces, mostly 36 in. wide, and from 60 to 80 in. long at the tuyeres, which entered between cast-iron water-jackets. Subsequently, Franz Casin, at the Holden works, Denver, put the tuyeres in the middle of the jackets, and many other improvements were made by him and others, so that the cast-iron jacket became the favorite, until the possibility of

Jan 1, 1913

By Joseph F. Shirley, Harry K. Burke

On January 9, 1964, the molybdenite extraction circuit at the San Manuel concentrator was switched from the sodium hypochlorite-ferracyanide process to the present process described in this paper. This moment was the culmination of over three years of laboratory and pilot plant test work. COPPER CONCENTRATOR Even though San Manuel's molybdenite plant begins with the thickened copper concentrates, it is necessary to describe briefly the copper sulfide circuit, because San Manuel's excellent molybdenite recovery actually begins with the primary grinding of the ore.

Jan 3, 1965

By T. R. A. Davey

IN 1947 the author became interested in the fundamental aspects of the desilverizing of lead by zinc, conducted some experimental work, and searched the technical literature for all available fundamental data. Since then a revival of interest in the subject in Europe resulted in the appearance of quite a number of papers. It became evident that it would be more profitable to collect together and examine thoroughly the results of various workers, than to attempt to duplicate the experimental determinations. There are many inconsistencies in the various publications, and it is opportune to review at this time the present status of knowledge on the Ag-Pb-Zn system. There is also a need for a clear description, in fundamental terms, of the various desilverizing procedures. This paper is presented in four sections: 1—There is an historical review of the origins of the Parkes process, of the results of many attempts to find a satisfactory fundamental explanation for the phenomena, and of the modifications proposed to date. 2—A diagram of the Ag-Pb-Zn system is presented. This is believed to be free of obvious inconsistencies or theoretical impossibilities, although thermodynamic analysis subsequently may reveal errors. 3—The fundamental bases of the various desilverizing procedures, which have been used up to the present day, are described; and a new method is suggested for desilverizing a continuous flow of softened bullion in which the bullion is stirred at a low temperature in two stages producing desilverized lead at least as low in silver as that from the Williams continuous process and a crust which, on liquation, yields a very high-silver Ag-Zn alloy. 4—A suggestion is made for the revival of de-golding practice, following a recently published account which does not seem to have attracted the attention it deserves. The terms "eutectic trough" and "peritectic fold" as used in this paper are synonymous with "line of binary eutectic crystallization" and "line of binary peritectic crystallization" as used by Masing.' The German literature on ternary and higher systems is rather extensive and a fairly general system of nomenclature has arisen, whereas in English usage the corresponding terms are not as well established. For this reason the meanings of terms used in this paper, together with the equivalent German terms, are given as follows: 1—Eutectic trough—eutektische rinne: line at which a liquid precipitates two solids S1 and S2 simultaneously. If the composition of a liquid which is cooling reaches this line, it then follows the course of this line until a eutectic point is reached, or until all the liquid is exhausted. The tangent to the eutec-tic trough cuts the line joining S1S2. 2—Peritectic fold—peritektische rinne: line at which a solid S1 and a liquid L transform into another solid S2. If the composition of a liquid which is precipitating S1 reaches the line, on further cooling only S2 is precipitated. The liquid composition moves from one phase region (L + S1) into the other (L + S2), and does not follow the course of the boundary. The tangent to the peritectic fold cuts the line S1S2 produced nearer S,. 3—Liquid miscibility gap, or conjugate solution region—mischungslucke: the region within which two liquid phases coexist in equilibrium over a certain range of temperature. A system whose composition is represented by a point in this region comprises one liquid at high temperature; then as the temperature is progressively reduced, two liquids, one liquid and one solid, one liquid and two solids, and finally three solids. 4—Liquid miscibility gap boundary—begrenzung der flussigen mischungsliicke: the line along which the surface of the miscibility gap dome, considered as a solid model, intersects the surrounding liquidus surfaces. 5—Tie lines—konoden: lines joining points representing the compositions of two liquids, a liquid and a solid, or two solids, in equilibrium. In binary systems the only tie lines customarily drawn are those through invariant points, e.g., through the eutectics of the Pb-Zn and Ag-Pb systems, or the various peritectics of the Ag-Zn system, as in Figs. 1 to 3. In ternary systems it is desirable to draw sufficient tie lines to indicate the slopes of all possible tie lines. 6—Ternary eutectic point—ternares eutektikum: point at which liquid transforms isothermally to three solids, S1, S2, and S Such a point can lie only within the triangle 7—Invariant peritectic (transformation) point— nonvariante peritektische umsetzungspunkt: (a) — On the miscibility gap boundary, the point at which two liquids and two solids react isothermally so that L, + S, + L, + S2. (b)—On the eutectic trough, the point at which a liquid and three solids react iso-thermally so that L + S, + S2 + S3. Such a point must lie on that side of the line joining S,S which is further from S,. (c)—A further possibility, not found in this ternary system, is that the point is at the intersection of two peritectic folds when the reaction concerned is L + S, + S, + S Historical Introduction Karsten discovered in 1842 that silver and gold may be separated from lead by the addition of zinc.2 Ten years later Parkes used this fact to develop the well known desilverizing process which bears his

Jan 1, 1955

By W. C. Smith, P. J. Hickey

After a short historical background of the process evolution, this article descvibes present-day plant facilities and operating techniques utilized for high-purity bismuth production. The plant is one of the world's largest, with an annual output of some one million pounds of refined bismutlz. PREVIOUS papers1 written by staff members of Cerro de Pasco Corp. have referred briefly to the production of refined bismuth. Since the Corporation is one of the world's foremost producers of high-purity bismuth, a detailed description of the process for extracting the metal may be of general interest. Following a short historical background of the development of the actual process, this presentation will trace the progress of bismuth from its entry into the primary smelting circuits to its concentration in electrolytic lead cell slimes. Our facilities for the treatment of anode muds will be described and the extractive methods given in some detail, with particular emphasis on the techniques which result in the production of refined metal. HISTORICAL BACKGROUND Shortly after Cerro de Pasco began smelting operations at Oroya, Peru in 1922, it became apparent that the dust carried by copper converter gas contained appreciable amounts of bismuth. Although dust collection efficiency was poor prior to building of the 550-ft stack and installation of the central cottrells in 1938, a large stock of dust was accumulated during the intervening years, having the following approximate composition: Oz. per ton Ag - 11.0 Pct Sn — 0.5 Pct Pb - 49.0 Pct Zn - 6.5 Pct Bi - 2.0 Pct Insol. - 1.5 Pct Cu - 0.7 Pct Fe - 2.3 Pct Sb - 3.0 Pct S - 10.0 Pct As - 7.5 In the mid-1920's, experimental crucible melts of this dust with carbon indicated that most of the bismuth and silver, and some of the lead, could be reduced to a fairly clean bullion. Other products were a small amount of leady copper matte and a slag high in zinc, arsenic, antimony, and lead; this slag contained some tin but only small quantities of silver, bismuth, and copper. After the laboratory results had been confirmed by operation of a small reverberatory, a dust reduction furnace was constructed. The ±10 pct Bi-Pb bullion produced from this operation was stocked until 1930, when an Oroya-designed converter type furnace3 was installed for the elimination of arsenic, antimony, and some lead from the bullion. This process concentrated the bismuth from 10 to about 60 pct. By means of the bismuth process developed4 by W. C. Smith at East Chicago (1909-1914) and the discovery of a method5 for separation of lead from bismuth with chlorine gas in 1929, it became possible to begin production of refined bismuth. Unfortunately, bismuth deleaded with chlorine always contained residual chlorides, and the removal of the chlorides by caustic soda left a lead content of 0.02 to 0.04 pct. This final problem was solved6 by substitution of air-blowing for the caustic treatment, which effectively removed all excess chlorine and gave bismuth which was practically lead-free. In 1934, a pilot electrolytic lead refinery began operations at Oroya. Lead smelting was resumed in 1935 and two years later a 100-ton-per-day lead refinery was put into service. In conjunction with the latter, the present-day Anode Residue Plant was constructed. Until 1940, the plant treated both lead anode slimes and dust reduction bullion. The dust reduction furnace was shut down in that year, and all cottrell dusts (with the exception of the product from the arsenic cottrell) were mixed with pyrite and treated in a Wedge roaster to eliminate all possible arsenic. Calcine from this operation joined the sinter plant feed; hence the bismuth from the copper and lead circuits was collected in the lead bullion and subsequently in lead anode slimes from the electrolytic lead refinery. The latter source has been the only bismuth-bearing material of any consequence entering the Anode Residue Plant from late 1940 to the present. A copper refinery began operating in 1948, and the cell mud from this plant is mixed with lead slimes and processed through the same circuit, though only a small quantity of bismuth is present in electrolytic copper cell residues. BISMUTH INTAKE Present-day routes which are followed by the new bismuth feed from its entry into the primary smelting circuits to its arrival at the Anode Residue Plant are traced schematically in Fig. 1. As illus-

Jan 1, 1962

By Orville R. Lyons, A. C. Richardson

THE Coal Division of Battelle Memorial Institute, during the course of an investigation conducted for a coal producer, carried out extensive sampling of the fine-coal section of a preparation plant. The information obtained during the testing period was used to evaluate the overall plant operation and provided the basis for. recommended changes in the plant flowsheet. Included in the information were data for the results obtained when using a Bird centrifuge to dewater the particular coal being treated. At the completion of the investigation, the coal producer gave Battelle permission to publish the operating data for the Bird centrifuge in order to make this information available to the coal industry. The authors wish to take this opportunity to express their appreciation to this company and its officials for their willing cooperation in the matter of publication. Description of Preparation Plant: The preparation plant on which the data were obtained is primarily a jig plant. The coal being washed is a low inherent moisture, friable, Pocahontus-type. The washed coal from the jig is dewatered and sized at 1/4 in. by a Parrish-type dewatering and sizing screen. The minus 1/4-in. coal and the underflow water from the shaker screen flow to a sludge tank where, after the coal settles, the coal is conveyed by a flight conveyor into a pump sump. The pump, a 300-gpm sand pump, delivers a mixture of fine coal and water to a Bird centrifuge which produces a dewatered cake and an effluent. At the time when the investigation was made, the centrifuge effluent was returned to the sludge tank and the cake product was mixed with a coarser size of coal prior to loading into railroad cars. The semiclarified water from the sludge tank overflowed into a pump sump from which it was pumped to the jig. As originally designed, the preparation plant was intended to operate with a closed water system but it was found, soon after operations started, that for various reasons this could not be done. In order to provide for removal of excess water from the system and to aid in maintaining the desired minimum percentage of solids in the circulating water, the plant was operated to provide from the sludge tank a continuous overflow of water and fine solids that was sent to waste. The centrifuge feed represents approximately 35 tons per hour of minus 1/4-in. coal in the form of a pulp averaging 38 pct solids. Sampling Procedure: In order to obtain basic data for determining the relationships existing in the fine-coal section of the preparation plant, samples were obtained for: (1) the centrifuge feed, (2) the centrifuge cake, (3) the centrifuge effluent, and (4) the circulating water. The sampling was arranged to start on a Monday, in order to begin operations with a sludge tank full of fresh water, and was continued long enough to allow sufficient time for fine solids to build up within the water system or for a two-day, four-shift operating period. All samples were taken at half-hour intervals except when plant stoppages occurred. After a stoppage the plant was

Jan 1, 1951

By Orville R. Lyons, A. C. Richardson

THE Coal Division of Battelle Memorial Institute, during the course of an investigation conducted for a coal producer, carried out extensive sampling of the fine-coal section of a preparation plant. The information obtained during the testing period was used to evaluate the overall plant operation and provided the basis for. recommended changes in the plant flowsheet. Included in the information were data for the results obtained when using a Bird centrifuge to dewater the particular coal being treated. At the completion of the investigation, the coal producer gave Battelle permission to publish the operating data for the Bird centrifuge in order to make this information available to the coal industry. The authors wish to take this opportunity to express their appreciation to this company and its officials for their willing cooperation in the matter of publication. Description of Preparation Plant: The preparation plant on which the data were obtained is primarily a jig plant. The coal being washed is a low inherent moisture, friable, Pocahontus-type. The washed coal from the jig is dewatered and sized at 1/4 in. by a Parrish-type dewatering and sizing screen. The minus 1/4-in. coal and the underflow water from the shaker screen flow to a sludge tank where, after the coal settles, the coal is conveyed by a flight conveyor into a pump sump. The pump, a 300-gpm sand pump, delivers a mixture of fine coal and water to a Bird centrifuge which produces a dewatered cake and an effluent. At the time when the investigation was made, the centrifuge effluent was returned to the sludge tank and the cake product was mixed with a coarser size of coal prior to loading into railroad cars. The semiclarified water from the sludge tank overflowed into a pump sump from which it was pumped to the jig. As originally designed, the preparation plant was intended to operate with a closed water system but it was found, soon after operations started, that for various reasons this could not be done. In order to provide for removal of excess water from the system and to aid in maintaining the desired minimum percentage of solids in the circulating water, the plant was operated to provide from the sludge tank a continuous overflow of water and fine solids that was sent to waste. The centrifuge feed represents approximately 35 tons per hour of minus 1/4-in. coal in the form of a pulp averaging 38 pct solids. Sampling Procedure: In order to obtain basic data for determining the relationships existing in the fine-coal section of the preparation plant, samples were obtained for: (1) the centrifuge feed, (2) the centrifuge cake, (3) the centrifuge effluent, and (4) the circulating water. The sampling was arranged to start on a Monday, in order to begin operations with a sludge tank full of fresh water, and was continued long enough to allow sufficient time for fine solids to build up within the water system or for a two-day, four-shift operating period. All samples were taken at half-hour intervals except when plant stoppages occurred. After a stoppage the plant was

Jan 1, 1951

By H. E. Stauss

THE short-time tensile-strength test, while it has not attained the practical importance of the creep test at elevated temperatures, has the advantage of being rapid and is satisfactory for determining relative strengths. For platinum and its alloys, which frequently are used at high temperatures under ill-defined conditions, the short-time test provides useful data that aid in the selection of the most promising alloy for each use. Measurements have been made upon 13 alloys, using wire 0.0031 in. in diameter, the choice of alloys and of wire size being determined by the requirements of a practical problem. Later the measurements were extended to higher temperatures by the use of heavier wire, 0.010 in, in diameter. APPARATUS The use of fine wire made the necessary equipment very simple, since no effort was made to measure the elongation. A small platinum-wound horizontal furnace was used, ½ in. in internal diameter by 1 ft. long, with the wire threaded through it. The wire, clamped at one end while the other ran over a ball-bearing pulley, of a total length of 29 in., was loaded directly by running sea sand into an attached basket, making the rate of loading and not of elongation constant. The sand flowed from a funnel, under a head of 3 in., into the basket. 3 in. below the orifice, at the rate of 170 grants per minute. Pulley friction and bending resist-ance of the wire as well as the force of impact of the sand could be neg-lected, being 3 grams and 0.5 grams, respectively. Furnace temperature .was uniform within ±5° C. over a length of 3 in., although a drift with time occurred, the largest amount being 20° C. No provision was necessary for a controlled atmosphere.

Jan 1, 1939

By F. V. Lene, E. J. Westerman

The recrystallization behaviors of two extruded and cold-drawn experimental sintered aluminum powder alloys, containing 1.75 and 3.0 pct Al2O3 by weight, were compared with that of extruded and cold-drawn commercially pure alumirzum. The kinetics of recrystallization of the alloys are described semiquantitatively. For the alloy containing 1.75 pct A l203 the rates of nucleation and of growth were also semiquantitatively determined. THE most striking property of aluminum alloys strengthened by a dispersion of Al2O3, the so-called SAP alloys, is their stability at elevated temperatures. One of the manifestations of this stability is their resistance to recrystallization after they have been cold worked. Most of the commercial grades of either the Swiss SAP or of Alcoa's Aluminum Powder Metallurgy Products have not been recrys-tallized after cold working, even when they are heated for a long time at a temperature near the aluminum melting point. Lenel, however, observed that the dispersion strengthened aluminum alloys with a larger spacing between the oxide particles than that of most commercial grades would recrys-tallize.1 It appeared to be of interest to further investigate the mode and kinetics of recrystallization of these alloys, and to compare their recrystallization behavior with that of commercially pure aluminum. Because homogeneous deformation of these SAP alloys in tension did not provide sufficient cold work to induce recrystallization, they were cold worked by wire drawing; the nonuniformity of this deformation unavoidably complicated the interpretation of the recrystallization studies. EXPERIMENTAL DETAILS Extrusions—Two types of sintered aluminum powder extrusions were used in this study. One type, designated AT-400, was produced from Reynolds atomized aluminum powder consisting of spherical particles averaging 3µ in diam and containing 1.75 wt pct of Al2O3. This powder was very similar to the R3M powder from which extrusions were previously prepared with an average spacing of 0.9µ between oxide particles.2 The second type, designated MD 2100, was produced from Metals Disintegrating Co. flake powder containing 3.0 wt pct of Al2O3, with an average flake thickness of 0.8µ. The average spacing between oxide platelets in MD 2100 extru- sions was 0.45µ.2 Powder compacts of 3/4-in. diam were extruded at 1000°F into 0.097-in. diam (AT-400) and 0.093-in. diam (MD 2100) wires by methods previously described.3 In order to compare the recrystallization behavior of sintered aluminum powder extrusions with that of wrought commercially pure aluminum 3/4 in. rod stock of 1100 F aluminum was extruded at 1000°F into 0.102-in. diam wire. Wire Drawing—Tungsten carbide dies were used for the AT-400 and 1100 F alloys. They had an included angle of about 15 deg and reduced the wire area approximately 7 pct per pass. Steel dies with an included angle of 11 to 13 deg and an average reduction per pass of 10 pct were used for drawing the MD 2100 alloy, because drawing this alloy through the carbide dies produced overdrawing defects. Heat Treatment—The cold-drawn wires were cut into small samples, and the deformed ends were etched off. The samples were each wrapped tightly in a single layer of aluminum foil, and individually isothermally annealed in a lead bath. Metallography—The modes and kinetics of recrystallization were determined by metallography. Mounted and polished specimens were anodized in a solution of 1.8 pct HBF4;4 examination under polarized light clearly revealed their grain structures. The recrystallized grains were generally much larger than those of the unrecrystallized matrix, and could clearly be distinguished because they alternated between maximum and minimum light reflection when the microscope stage was rotated, while the unrecrystallized matrix had a comparatively homogeneous "salt and pepper" structure. The fractional recrystallized volumes of the dispersion hardened alloy wires were determined by cutting and weighing of recrystallized and total transverse areas on photomicrographs. The recrystallized grains in the 1100 F alloy were too small to be cut out individually; therefore a combination of cutting and lineal analysis was used in this case. RESULTS AND DISCUSSION Modes of Recrystallization—The modes of recrystallization of the three alloys varied widely. In the 1100 F alloy nucleation and growth started in the region midway between the center and the surface;

Jan 1, 1961

By Edward Martinez

The literature on the beneficiation of serpentine asbestos ore states that the specific gravity of the rock containing fiber and the fiber itself is the same so that specific gravity cannot be used as a basis for separation. An investigation of particles of asbestos ore ranging in size from 4 mesh to 5 in. revealed that the fiber-bearing particles generally had a lower specific gravity compared to essentially barren particles. The difference in specific gravity was sufficient to make an effective gravity separation of ore particles and reject a low grade tailings product. The results of laboratory tests are described that lead to the running of pilot plant tests with 20-ton ore samples using a heavy media separator. These tests indicated that a large amount of low grade material could be rejected from the raw ore at a coarse size. Although the process used is a wet separation, it can be added to a conventional dry asbestos mill. A gravity preconcentration would allow the mining of lower grade ore, increase the life of the mine, increase the value of the feed to the fiber recovery section, and decrease maintenance costs by removal of low grade tailings early in the flowsheet.

Jan 1, 1975

By L. A. Panek

Research on roof bolting has developed significant information with respect to the functions that can be performed by properly installed bolts and the methods that can be used to ensure that effective installation has been achieved. The present paper brings together results that are pertinent to the design of a bolting system for reinforcing horizontally stratified mine roof with vertical bolts. The design approach is based on the use of bolts to decrease the maximum bending stress within an arbitrary sequence of roof strata (see Fig. 2) in which the only loads are the weights of the individual beds. The use of rock bolts to achieve other objectives, such as to stabilize a fractured rock mass on the periphery of a mine opening,3 is excluded from consideration herein. Reinforcing effect of bolts consists of a friction effect and a suspension effect in combination. "Friction effect" refers to reduction of bending in a stratified roof due to clamping action of tensioned bolts, which compress the strata, thereby creating frictional resistance to displacement along planes of stratification. (The flexure of a sequence of strata is necessarily accompanied by horizontal displacement of each bed with respect to its neighbor, along mutual surfaces of contact.) Although the sliding of the beds tends to exert a shearing effect on the bolts, tests show that this has negligible influence on the bending stresses, probably because the bolts are so far apart. "Suspension effect" refers to the transfer of part of the weight of the weaker or thinner strata to one or more thick strata, which occurs when strata with differing tendencies to deflect are constrained to have equal deflection. The term "suspension" is used irrespective of whether the thin strata are above or below the thick strata, because tests show that the supporting effect of the flexurally more rigid roof strata is for practical purposes independent of their position in a bolted sequence of strata. An earlier paper4 dealt with a bolted unit in which all strata had equal flexural rigidity; bolting could develop a friction effect but no suspension effect. The present paper deals primarily with a bolted unit in which all strata do not have equal flexural rigidity, so that bolting can develop both a friction effect and a suspension effect. The reinforcing effect of the bolts was determined by a combination of experimental and theoretical stress analysis.4'6'7 The experimental method consisted of loading a model by rotating it in a centrifuge to simulate the gravity forces exerted about an underground excavation. Models were made from two different limestones, although the results are independent of the material because the elastic constants were taken into account in the data analysis. Electric resistance strain gages were bonded to the highest and lowest members of the bolted unit to determine the outer-fiber bending strains at the center and at the supports. The quantitative results presented herein are based on tests of 56 different models, which included a range of values for the structural parameters. Each model was tested several times, with different combinations of spacing and tension of the bolts. DESIGN EQUATIONS The effects of friction and suspension on the flexure of a bolted stratified roof can be expressed by the following equations, which give the maximum bending stress, the deflection, the friction effect, the suspension effect and the transferred load for any (the i-th) member of the bolted unit in terms of the structural parameters:

Jan 1, 1964

By C. R. Boyd

In 1874 I received an appointment in the corps organized by Col. William P. Craighill, U. S. Engineers, to survey the New or Upper Kanawha river for the purpose of ascertaining the practicability and

Jan 1, 1895

By Whitmell J. Smithwick, Charles E. Dandois

The first operating unit of the ring or annular shaft kiln as developed by Mr. Karl Beckenbach of West Germany was rated 185 tpd and was installed in 1962. Today, there are approximately 84 units in operation in Western Europe and Japan. Twenty-one additional units are under construction or on order. Sizes range from approximately 100-330 tpd, however, units of over 500 tons are in the design stage. The ring shaft kiln is a single vertical cylinder to which are attached the required service platforms, combustion chambers, injectors and distribution piping (Fig. 1). Auxiliaries such as recuperator, exhaust fan and dust collector are mounted on the appropriate service platform. The charging pan, feed conveyor, etc., are mounted at the kiln top and at the bottom of the kiln is the silo with a vibrator type discharger. Mounted separately is the required blower for injector air and fan for internal cooling.

Jan 4, 1973

By William M. Barret

In order properly to evaluate the data related to geomagnetic surveys, it is highly desirable to have all the available information concerning the magnetic properties of the involved media. This is true of surface as well as subsurface formations and becomes of increasing importance as advancement in instrumental and interpretative technique permits the analysis of relatively slight magnetic reflections of buried masses. Since the magnitude and areal distribution of the usually complex resultant magnetic field at the surface of the earth is a function of the susceptibility1 of the involved matter, it necessarily follows that susceptibility determinations will afford valuable data for geophysical correlation. The present method was developed to meet the need for routine analysis of specimens in the geomagnetic laboratory where it was felt that general utility, ruggedness, ease of manipulation and reproducibility of results were of greater importance than extreme sensitivity. Electric Circuits The electric circuits are illustrated diagrammatically in Fig. 1. Current from the battery, B, passes through the variable resistance R1, reversing switch S1, double-break key K1 to switches S2 and S3. From S2 it flows through variable resistance R2, ammeter A2, to the primary coils of the fixed mutual inductance Mf and variable mutual inductance M,. . From S3 it passes through variable resistance R3, ammeter As, to the primary winding of the fixed mutual inductance MI, which is designed to receive the test sample. The ballistic galvanometer2 or fluxmeter3 G is connected in the circuit containing the double-break key

Jan 1, 1932

By John Henderson

Densities of melts 0f the lime-iron oxide-silica system in contact with solid iron have been measured by the maximum bubble pressure method in the temperature range 1250° to 1440°C and the composition range 0 to 40 mol pct lime, 15 to 100 mol pct iron oxide, and 0 to 55 mol pct silica. Densities range from 4.65 g cm 3 for wustite at 1440°C to 2.75 g cm-&apos; at 1350°C for a melt containing 30 mol pct lime, 20 mol pct iron oxide, and 50 mol pct silica. The results are interpreted in terms of a postulate that the melts can be regarded as a random array of oxygen ions in which regions of local order exist to satisfy the coordination requirements 0.f the cations. An understanding of the nature of metallurgical slags is basic to the development of a sound theoretical description of heavy metallurgical extractive and refining processes. Because these liquids are complex, direct measurements of their properties has not thrown much light on their structure. This has led to the approach of measuring the properties of simpler liquids, and building up their complexity until slag compositions are reached. In this way the density of liquid iron silicates was measured in a previous study1 and the present work represents a further stage in this synthesis. EXPERIMENTAL The technique used in the measurement of density was the maximum bubble pressure method. Details of the apparatus and procedure were similar to those previously reported,&apos; with the exception that a constant voltage transformer was used to supply the power input to the furnace and six silicon carbide resistance elements were used in place of the molybdenum winding. With these modifications melt temperature could be maintained within 1 centigrade degree during the course of a run. The silica used to prepare the melts was washed natural quartz ignited at 1000°C; wustite was prepared by air-melting A.R. grade ferric oxide in an iron crucible and lime was prepared by air ignition, at 1000°C, of weighed quantities of A.R. grade calcium carbonate, previously air-dried at 110°C. The finely ground constituents were intimately mixed in a glass ball mill prior to melting. Temperatures quoted are accurate to * 5°C and the standard deviation of the density values, calculated by the method of least squares, ranged from 0.5 to 1.8 pet. However, replicate determinations of density on different melts of the same nominal composition at the same nominal temperature did not vary by more than 1 pct, Table I, and this figure has been taken as an estimate of the accuracy of the density results. The density of carbon tetra-chloride was also measured as a check on the absolute performance of the experimental method. At 20°C a value of 1.593 * 0.002 g cm"3 was obtained; this compares with the literature value2 of 1.595 g cm"3. Results of experiments designed to measure the dependence of the density of lime-iron oxide-silica melts, in contact with solid iron, on composition and temperature are shown in Table I. Because iron sometimes precipitated in the sample during quenching, the Fe203 chemical analyses were only poorly reproducible and should be taken as a guide rather than as absolute values. Fig. 1 shows the data from various sources for the density of liquid iron silicates and Fig. 2 shows isodensity contours at 1410°C for lime-iron oxide-silica melts, calculated by graphical interpolation of smoothed curves drawn through the experimental results, together with the 1400°C results of Adachi and ogino3 and Pope1 and Esin.4 Fig. 3 shows the isothermal variation with composition of the volume of melt per gram ion of oxygen at 1410°C and Fig. 4 shows regions in which the temperature coefficient of this volume is negative, positive, or negligible (<0.005 cm3 deg-I). DISCUSSION a) Disparity Between Reported Density Results. Consider the system iron oxide-silica, the results for which are summarized in Fig. 1. Although there is some difference in the temperatures at which the various densities apply, this difference is not sufficiently large to account for the observed discrepancies. The reliability of the present results for the low-silica region has been confirmed by measurement of the density of liquid wustite by three different techniques. At 1410°C the density measured by a balanced-column method was 4.55 g cm"3, by a combination balanced-column and gas-densitometer method 4.59 g emd3, and by a pycnometer method 4.53 g cm"3. Schenck, Frohberg, and Hoffermann&apos; have also reported a value of 4.55 g cm"3 for the density of liquid wustite at 1400°C. It must be concluded, therefore, that neither Pope1

Jan 1, 1964