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By James Norman, E. W. Gieseke

Spodumene is a lithium-bearing pyroxene, and is an important source of lithium compounds. Because of its high alumina and lithia content, it might be a desirable constituent of glass batches. The use of spodumene in ceramics will undoubtedly increase, as it has been shown that a mixture of the proper proportions of this mineral with soda feldspar and potash feldspar has a much lower maturing temperature upon firing than has any one of the three components. The mineral spodumene seldom occurs as pure mineral, but ordinarily with feldspars, mica and quartz. All spodumene-bearing rocks tested contained these minerals as gangue. Small amounts of iron and manganese oxides also were present, usually as stains on the silicate grains. Mica associated with the spodumene carried a considerable amount of lithia. This was discovered in flotation tailings, which, although virtually free of spodumene grains when examined microscopically, showed 0.4 to 0.5 per cent Li²O on chemical analysis. The lithia content of micas picked from the North Carolina spodumene rock was variable. Some samples contained little lithia, others more than 3 per cent. In 1938, under the supervision of Oliver C. Ralston,S the Bureau of Mines, in cooperation with the Tennessee Mineral Products Corporation, undertook to study the beneficiation of North Carolina spodumene by froth flotation. The Bureau had previously published investigations on the beneficiation of spodumene.2,3The methods described in the second publication3 did not give good results with badly weathered spodumene rock, so that more thorough investigation of the subject was warranted. Study of Method of Flotation Separation Most of the work on concentration of spodumene was done with rock from near King's Mountain, North Carolina. Rock from two

Jan 1, 1942

By James Norman, E. W. Gieseke

Spodumene is a lithium-bearing pyroxene, and is an important source of lithium compounds. Because of its high alumina and lithia content, it might be a desirable constituent of glass batches. The use of spodumene in ceramics will undoubtedly increase, as it has been shown that a mixture of the proper proportions of this mineral with soda feldspar and potash feldspar has a much lower maturing temperature upon firing than has any one of the three components. The mineral spodumene seldom occurs as pure mineral, but ordinarily with feldspars, mica and quartz. All spodumene-bearing rocks tested contained these minerals as gangue. Small amounts of iron and manganese oxides also were present, usually as stains on the silicate grains. Mica associated with the spodumene carried a considerable amount of lithia. This was discovered in flotation tailings, which, although virtually free of spodumene grains when examined microscopically, showed 0.4 to 0.5 per cent Li²O on chemical analysis. The lithia content of micas picked from the North Carolina spodumene rock was variable. Some samples contained little lithia, others more than 3 per cent. In 1938, under the supervision of Oliver C. Ralston,S the Bureau of Mines, in cooperation with the Tennessee Mineral Products Corporation, undertook to study the beneficiation of North Carolina spodumene by froth flotation. The Bureau had previously published investigations on the beneficiation of spodumene.2,3The methods described in the second publication3 did not give good results with badly weathered spodumene rock, so that more thorough investigation of the subject was warranted. Study of Method of Flotation Separation Most of the work on concentration of spodumene was done with rock from near King's Mountain, North Carolina. Rock from two

Jan 1, 1942

By C. H. Fulton

Page Introduction........................... 280 Distillation Products from Reduction of Zinc Ore........... 281 Preliminary Investigations..................... 290 Equilibrium of Reaction Zn + CO2 + ZnO + CO in Temperature Range of Condensation........................ 293 Requisites for Successful Condensation................ 299 Discussion............................ 300 Introduction The study of the condensation of zinc from its vapor was undertaken to shed light on certain problems encountered in large-scale electric zine-furnace work. recently conducted. It is a matter of common knowledge that one of the disadvantages of the arc type, electric zinc furnace, is the production of a large amount of blue powder and proportionately little liquid spelter, for, as stated by Stansfield,l "A notable defect in the electric smelting of zinc ores is the difficulty experienced in obtaining the distilled zinc in the liquid state; ...... when zinc ores are smelted electrically, very little liquid metal is commonly obtained, nearly all of the zinc being in the state of powder." This difficulty was not encountered in the electric furnace used12 the main difficulty being the destruction of the fire-brick lining of the large condenser, in certain parts, and the formation of .some oxide and dross coatings on the condensing surfaces. The reason for these difficulties could readily be explained by the infiltration of air whenever the condenser was changed from one retort base to the other, the zinc absorbed in the fire-brick forming oxide and causing rupture by its expending volume. Two facts, however, tended to disprove this theory for the rupture of the fire-brick lining: first, the large distilling retort, lined with fire-brick, which was shifted from charge to charge on alternate bases at temperatures between 1200 and 1300 C., fully exposed to the air, showed no signs of disintegration whatever; second, marked disintegration in the condenser occurred .only in places that were at tempera-

Jan 1, 1919

By W. E. Ruder

Considerable interest has been shown lately in various types of furnaces for the production of high temperatures, both for laboratory purposes and for small industrial uses. Dr. J. A. Harkerl described certain modifications of the Arsem graphite resistance helix for high-temperature work at the National Physical Laboratory. Dr. Northrup also has developed two interesting forms of high-temperature furnaces. In 1911, Messrs. Winne and Dantsizen2 described two forms of resistance furnaces using ductile molybdenum or tungsten as resistors. Since the discovery, by this laboratory, of methods for producing these metals in ductile form, their use as resistance elements in the research laboratory of the General Electric Co. has grown to such an extent that they are now almost indispensable. These furnaces are used here for alloy research, annealing, heat-treating, and practically all of the thermal processes requiring temperatures above 900' C. Various types of furnaces have been developed to meet the general needs of laboratory work, and it is the purpose of this paper briefly to describe some of these furnaces. A tubular furnace wound with a metallic resistor is the simplest type to construct, and almost every laboratory has made such furnaces. The only essential difference between a furnace wound with platinum o base metal and one wound with tungsten or molybdenum is that the coils of the latter must be heated in a neutral or reducing atmosphere It has long been our custom here to build our resistance furnaces with : metallic casing from which the tube could be easily removed for rewind ing. The heat-insulating material—usually calcined magnesia, alumina silica, or silex, depending upon the nature of the resistor-—is used i: powdered form so that it can be readily removed and replaced. I order to adapt this method to tungsten or molybdenum windings,

Jan 1, 1918

By M. Streat, R. Y. Qassim

Summary This paper reviews the developments in uranium process technology as a result of modern trends in the design of ion exchange equipment. The operation of the Cloete-Streat fluidised bed contactor is described and pilot-plant development of this contactor for uranium slurry work at Imperial College is outlined. The properties of a suitable ion exchange resin for slurry treatment are discussed, together with a description of relevant rheological properties of ore suspensions. A prediction of ion exchange resin- slurry behaviour is presented in order to establish the feasible operating limits for this type of R-I-P contactor.

Jan 1, 1973

By F. P. Haver, M. M. Wong

The Federal Bureau of Mines has developed a leach= electrolysis method for recovering lead from galena (PbS). Flotation concentrate is leached with ferric chloride (FeCl3) solution to obtain lead chloride (PbC12) and elemental sulfur, PbS + 2FeC13 -+ PbC12 + 2FeC12 + Sº. Over 99% of the P~S in the condentrate is converted to PbC12. The leach solution, composed of brine (NaC1) and FeC13, dissolves the PbC12 as it is formed, leaving a residue of elemental sulfur plus gangue. Very pure PbC12 (>99.9%).is crystallized from the leach filtrate on cool- ing and is subsequently electrolyzed in a fused-salt bath to obtain molten lead metal (>99.9% Pb). Spent leach solution (FeC12) is regenerated for further use by reaction with the chlorine obtained in the electrolysis of lead chloride, and other values, such as Ag, Cu, Zn, and Sº, are recovered by additional treatments of the leach solution and leach residue. This paper discusses results and conditions of the leaching and electrolysis operations.

Jan 1, 1976

By N. J. Themelis, R. V. Flint

In a four-year period between 1973-76, the Metal Mining Division of Kennecott Copper Corp. conducted a company-wide program of operations improvement which involved line and staff personnel from all company functions. Specialist teams performed detailed studies of all operational aspects at Kennecott's four copper mines and concentrators, four smelters, and two electrore-fineries. The specific recommendations that resulted may transform the Metal Mining Division into the country's lowest cost producer of copper by the mid-1980s. In this article, N. J. Themelis and R. V. Flint describe the mechanics of MMD's integrated approach to operations improvement.

Jan 6, 1979

By R. E. Hucko

Introduction The principal thrust of the US Department of Energy (DOE) coal preparation R&D program is to develop technology to reduce ash and sulfur content of US coals to levels where the resultant product can be formulated into a high quality solid or slurry fuel that could replace oil or natural gas in both new and retrofit applications. The coal preparation program is implemented by the Pittsburgh Energy Technology Center (PETC) and is conducted both in house and by a cross section of industrial organizations, universities, and national laboratories. Research is subdivided into three areas - coal characterization, beneficiation, and ancillary unit operations.

Jan 1, 1985

As REVISED IN CONFERENCE FEB. 17, 1931 REPORT OF SUBCOMMITTEE ON MINE VENTILATION CODE OF A, I. M. E. COMMITTEE ON VENTILATION ALFRED W. HESSE, Chairman THOMAS G. FEAR GEORGE S. RICE E. A. HOLBROOK WILLIAM A. ROWE HOWARD ECKFELDT R. V. NORRIS CADWALLADER EVANS, JR. H..1. SMITH, Secretary REALIZING, from the safety and health standpoint, the need of furnishing air of good quality to workers underground and of diluting and carrying away dangerous and noxious gases, the following code is offered to unify regulatory measures in coal mine ventilation, so far as is practicable, in the United States of America.

Jan 1, 1931

COLORADO SCHOOL OF MINES On February 21, 1918, the students of the Colorado School of Mines who are members or have applied for membership in the Institute re-organized the "Scientific Society of The Colorado School of Mines,? which has for its .purpose the study and discussion of mining problems. Its meetings will be held on the first and third Wednesdays of each month, in Guggenheim Hall. Following are the officers R. W. GIBSON, President, W. A. CONLEY, Vice-president, F. L. SERVISS, Sec'y.-Treas. At present there are 62 members, seven of whom are Faculty members. F. L. SERVISS, Secretary.

Jan 5, 1918

By A. L. Mowry, John T. Norton

Isothermal sections of the pseudo-ternary carbide systems TiC-VC-ZrC, TaC-VC-ZrC, and NbC-VC-ZrC have been examined. The sections contain an inverted U shaped two-phase field whose maximum extent is 77 molecular pct for Tic, 64 pct for TaC, and 51 pct for NbC. lsoparametric lines were determined which show the curvature of the single-phase field and the direction of tie lines in the two-phase field. It was not possible to show any relationship between atom size and the extent of the two-phase field. REVIOUS work' on the solubility relationship of several pairs of refractory cubic monocarbides from the group Tic, ZrC, VC, NbC, and TaC has shown that, with the exception of the system VC-ZrC, they all form continuous series of solid solutions. It was demonstrated that the limited solubility in the case of VC-ZrC system was primarily a question of the size factor, the size difference between vanadium and zirconium being greater than the limiting value of about 15 pet. It is interesting to consider the results of combining these isomorphous carbides in groups of three and to determine the isothermal section of the pseudo-ternary diagram at, for instance, 2000°C, a temperature that will assure an equilibrium structure in a reasonable length of time. In the ternary systems where the three pairs form complete solid solution series, it is probable that the isothermal ternary section would be a single- phase field, although this is not necessarily the case. However, if one pair of carbides has a very limited solubility, then there must be a two-phase field of some extent, and it is the purpose of the present work to examine the extension of this field and to determine if it can be related to the nature of the third carbide. This condition is fulfilled if the VC-ZrC pair, which has only limited solubility, is combined with Tic, TaC, or NbC. The relative sizes of the metal atoms undoubtedly play a role in the extent of the single-phase field, and the values obtained from the pure carbides, assuming close-packing between metal and carbon atoms and a carbon atom diameter of 1.54A, are

Jan 1, 1952

All the Affiliated Student Societies report that the present college year promises to be a most successful one. The officers for this year, so far as reported, are as follows: TUFTS COLLEGE CHEMICAL SOCIETY President, CHESTER B. PIERCE Vice-president, JOHN H. SCHMUCK Secretary-treasurer, DANIEL A. PRESCOTT President, R. G. SATTERLEY Third vice-president, H. E. BLAKE First vice-president, LESTER VOCKE Secretary,-J. E. FLANIGAN Second vice-president, `V. J. KLINE Treasurer, PROF. JAMES FISHER, JR About 75 per cent. of the students are ex-service men, and most of them saw service in France.

Jan 12, 1919

By J. E. Lawver, J. D. Raulerson, Charles C. Cook

The agriculture industry has made great strides during the past decade to increase agriculture yields through increased use of fertilizers. Increased use of fertilizers may prevent, or at least delay, mass starvation due to the alarming increase in world population. Phosphate was added to soil as a plant nutrient in the form of calcined bones at least 2000 years ago (Anon., 1964), and man has used phosphate minerals as a source of fertilization in one form or another for at least 100 years. During 1977 the world produced about 116 Mt of phosphate rock, with about 86% used for fertilizers and another 4% for animal feed supplements. More than three-fourths of the total production comes from the United States, Morocco, and the Soviet Union. From a mineral beneficiation point of view, the major sources of phosphate rock and the methods of beneficiation can be classified as follows: marine deposits not containing appreciable carbonate minerals, marine deposits requiring a francolite carbonate mineral separation, igneous deposits not containing appreciable carbonate minerals, and igneous deposits requiring apatite carbonate mineral separation. [ ] Guano, mostly from Chile and Peru, accounts for 0.1% of the total world production, and the calcium phosphates from Ocean, Nauru, and Christmas Islands and the aluminum and iron phosphates from Brazil and Aruba account for less than 4% of the world production and are thus not considered in this classification (Lawver, et al.). At present, marine phosphorite deposits account for about 75% of the world's production; the igneous deposits account for 20%. The igneous deposits low in carbonate minerals are easily concentrated by crushing, grinding, and apatite flotation. The most important igneous deposits are those of the Kola Peninsula, USSR (Woodrooffe, 1972). The igneous deposits high in carbonate materials are of corn appreciably more difficult to beneficiate, but they have been concentrated by froth flotation for a number of years. An interesting but rather complicated flowsheet of this type is at Phalabonva, in the Republic of South Africa (Lovell, 1976). The Phalaborwa deposit is an igneous complex of pyroxenite with a central core of carbonatite surrounded by a serpentine- magnetite-apatite rock called phoscorite. The phoscorite containing about 10% P2O5, 35% magnetite, and 35% calcium magnesium carbonate is currently being processed. The process involves comminuting the material for fiberation and subjecting it to a copper float using a potassium amyl xanthate as collector and triethoxybutane as a frother followed by a magnetic separation of the tailings to produce a feed for phosphate flotation. This process produces a phosphate concentrate containing greater than 36% P2O5 at a P2O5 recovery ranging from 75 to 80%. Considerable success has been claimed for recovering apatite from carbonate-bearing ores at the Jacupiranga Mine of Serrana S/A (Silva and Andery, 1972). The carbonatite currently being mined contains an average of only 5% P205 and is concentrated using a unique flotation process (Andery, 1968) to yield 96% P205 concentrates. The ore contains about 12% apatite, 5% magnetite, 80% calcite plus dolomite, and minor amounts of phlogopite, olivine, zircon, ilmenite, and pyrochlore. Feed preparation consists of crushing to -31.75 mm (-1 M in.), rod milling in closed circuit with hydrocyclones to about 92% (-50 mesh), and two-stage cyclone desliming of the -50 mesh sands at 20 m. Weight recovery in the deslimed feed is normally 85 to 88% and the corresponding P2O5 recovery is usually about 90%. The deslimed feed is conditioned at 60 to 70% solids for 15 min at pH = 8-10 with 0.6 kg/t of causticized starch for iron oxide and calcite-dolomite depression. The conditioned slurry is diluted to 20 to 30% solids, about 0.2 kg/t of fatty acid or soap collector is added to the conditioner discharge, and the reagentized ore is subjected to rougher-scavenger flotation with additional fatty acid added to the scavenger float. The scavenger concentrate is returned to rougher circuit distributor, and the rougher concentrate froth is subjected to two stages of cleaner flotation to yield a final apatite concentrate analyzing 36 to 38% P205. Flotation recovery of P205 is, in general, above 90% when treating fresh carbonatite. The high-carbonate flotation tails normally analyze 1 % P2O5 or less and are suitable for portland cement production. The marine deposits. Types 1 and 2 of central Florida are representative of enormous reserves of phosphate rock that will undoubtedly account for much of the world's production in the near future. Until very recently the sedimentary deposits high in carbonate minerals (Type 2) have not been considered reserves due to the difficulty in making a francolite-carbonate separation. Although no commercial plant has yet been built to beneficiate Type 2 ore, laboratory and pilot plant data indicate the process is viable. If so, the reserves of Florida and similar deposits throughout the world will be substantially increased. A discussion of the beneficiation of these two types of sedimentary deposits and the relation of the resulting concentrates to the fertilizer industry of the United States is the subject of this paper.

Jan 1, 1981

By W. W. Vaughn, R. H. Barnett, E. E. Wilson

Soon after the search for uranium ores on the Colorado Plateau began in earnest, thousands of feet of drill core ranging from 1 1/8 to 2 1/8 in. diam became available for study. Although significant advances had been made in the technique of quantitative gamma-ray borehole logging, instrumentation was in the development stage, and complete reliance could not be placed on gamma-ray logs alone to interpret quantitatively the meaning of radioactivity in a drillhole. A method that would be faster than chemical analysis and still give reproducible and reliable results for various drill core sizes was desirable to provide additional information on the enormous footage of drill core being accumulated. A solid phosphor scintillation drill core scanner was designed and constructed. Basically the instrument was developed to measure radiation from a drill core which would not be clearly recorded by a gamma-ray logger using a Geiger tube as the sensitive element. Such data would be beneficial in constructing isorad maps to delineate ore-bearing zones. A calibration in the range 0.01 to 0.1 pct eU.,O, was provided; above 0.1 pct eU3O8 gamma-ray logs were available and were being used to calculate grade and tonnage of ore reserves. The core scanner, however, has been used to estimate equivalent uranium content of ore-grade materials containing as much as 2.2 pct eU3O8 with an accuracy of ± 10 pct, the sample being in the form of a BX drill core. Actually, an apparent calibration of eU3O8 vs counts per unit time is a straight line with a slope that is a function of the sensitive element and the geometry of the counting assembly. A true calibration that will show the expected departure from a straight line is due principally to the random nature of the pulse from a radiation source and the nonlinearity of the electron circuitry. Design and Construction: Three methods of detecting radioactivity were considered and applied in developing the core scanner now in use: 1) the Geiger tube, 2) liquid scintillation phosphors, and 3) solid scintillation phosphors. The desired sensitivity and long-term drift characteristics needed for this operation could be attained only by using solid scintillation phosphors. All three methods are discussed. Before scintillation counters were common, nine beta-gamma sensitive Geiger tubes 7/8 in. diam by 12 in. long were used, arranged to surround the drill core with tube axes parallel to the axis of the core. This arrangement of Geiger tubes was en- closed in a lead shield 1 in. thick, and provision was made to slide a 6-ft length of drill core manually into the counting chamber, one foot at a time. A count for each segment was taken with a scaler while the core remained stationary. The equivalent uranium content of the different sections of drill core could then be estimated with the aid of a calibration curve of counts per unit time vs percent equivalent uranium (eU). In rare cases the effects of the radioactivity concentrated in small areas within the core introduced errors in the readings made with the Geiger tube arrangement owing to the geometry of the measurement. The variability of counting rate due to a localized concentration of radioactivity in a spot in the wall of a drill core is illustrated in Fig. 1. This effect and the inherent low efficiency of the Geiger tube were considered major disadvantages of this counting arrangement. When liquid scintillation phosphors became available the core scanner in Fig. 2 was constructed to make a more accurate measurement of the equivalent uranium content of a sample. This instrument contains about 4 liters of liquid phosphor in a stainless steel coaxial cylinder 1 ft long, with inner and outer walls 0.060 in. and 0.125 in. thick, respectively. Four end-window type photomulti-plier tube with cathodes of 2 in. diam, immersed in the solution at right angles to the axis of the core, were used to observe light flashes in the phosphor. The liquid phosphor offered equal sensitivity to radiation originating at any point in the enclosure and represented geometrically the optimum in design. However, providing a semi-permanent leak-proof seal between the glass envelope of the phototube and the metal walls of the container proved to be a serious problem in constructing the equipment. The most effective seals were especially machined O-rings from sections of large tygon tubing. The tygon took a permanent set owing to cold flow characteristics and in most cases sealed completely. The light absorption characteristics of the liquid phosphor changed gradually with time, and after one month the counting rate had decreased to half the original value. The most sensitive liquid phosphor tested proved to be a solution containing 4 g of 2.5-diphenyloxazole and 0.01 g of 2-(1-naphthy1)-5-phenyloxazole per liter of toluene. With fresh solution in the chamber and with all photomultiplier tubes operating in parallel, the counting rate contributed by any one of the four photomultiplier tubes was about 85 pct of the counting rate from a single tube operated individually. From these observations it was concluded that owing to coincident loss and light attenuation within the liquid phosphor, the apparent sensitivity could not have been materially increased by additional phototubes. However, this approach to core

Jan 1, 1960

By M. H. Pilkuhn, H. Rupprecht

p layers on n-type substrates of GaAs were prepared by difhsion of zinc into these substrates. Experimental details about the diffusion technique will be presented. It will be shown that inhomogeneity in the substrate doping level and crystal shucture defects cause nonplanarity in the junction plane. These nonplanarities give rise to high threshold current densities. From incremental sheet-conductivity measurements the zinc profile is derived. Data about surface concentration and gradient of the zinc concentration at the junction are presented. Finally, some results of laser diodes, which were made of wafers with extremely planar diffusion fronts, are discussed. ThE investigations reported here are part of a program concerned with the materials aspect in the fabrication of GaAs laser diodes. This includes studies of the influence of crystal defects introduced both during the growth of the crystal and as a consequence of subsequent diffusion. Compared with the progress made in the investigation of structure defects, for instance by X-ray microscopy in materials like silicon, no similar results are known so far for the case of GaAs. The further progress in laser diodes will depend very much upon our knowledge of these structure defects and subsequently upon our ability to avoid them. Laser diodes of GaAs had been made mostly by diffusing an acceptor impurity, normally zinc, into an n-type substrate. This procedure was found more convenient than the inverse one for various technical reasons. We will follow this line and restrict ourselves to zinc diffusion only in this paper. DIFFUSION TECHNIQUES AND EXPERIMENTAL DETAIL In order to set a refgrence for our experiments, we used a standard procedure for the diffusion runs. The diffusions were carried out in a sealed quartz tube for 2-1/2 hr at 850°C. One milligram zinc arsenide (ZnAsz) per cu cm of the tube volume served as zinc source. Under these conditions the junction depth was between 25 and 30 p. It is assumed, according to the investigations of the zinc arsenide system by V. Lyons and V. ilvestri,' that ZnAs2 dissociates completely at 850°C into ZnsAsz and Asg. Since there is a condensed phase of ZnsAs2 in the tube at 850°C, the partial pressure of zinc is controlled by the pressure of the arsenic formed by the dissociation of ZnAs2. Due to the fact that the zinc partial pressure resulting from the dissociation of ZnsAsz is only a weak function of the partial pressure of arsenic, the amount of the initial ZnAs2 is not very critical. Experimentally, we found that the junction depth, as well as the surface concentration, was practically the same, if the ingots of ZnAs2 varied between 0.2 and 5.0 mg per cu cm. The wafers were oriented with the main faces in (100). Surface damage due to cutting and lapping was removed by a chemical-etch treatment in a rotating beaker (etch solution: HzSOr, H202, and HzO in a ratio 3:l:l). The GaAs substrate material was n type and doped with tellurium, selenium, silicon, or germanium. It was grown by means of horizontal or Czochralski techniques. After the diffusion process, junction depth and planarity were checked either by grooving with a

Jan 1, 1964

By Richard McCaffery

CONTENTS PAGE 1. Research on Blast-furnace Slags (A Progress Report). By Richard S. McCaffery 5 2. Determination of Viscosity of Iron Blast-furnace Slags. By Richard S. McCaffery, Clarence H. Lorig, Ira N. Goff, Joseph F. Oesterle and Oscar O. Fritsche 27 3. Effect of Magnesia on Slag Viscosity. By Richard S. McCaffery, Joseph F. Oesterle and Oscar O. Fritsche 55

Jan 1, 1931

By William P. Roe, James F. Ellis, W. P. Fishel

This paper reports the distribution of sulphur between iron and titanium in Fe-Ti-S alloys and in Fe-Ti-C-S alloys. The compound TiS was found, which exists as a separate phase at temperatures below 1200°C. Titanium in steels reacts with sulphur in preference to carbon. THE general relationship between sulphides and oxides suggests that titanium should form a sulphide having properties similar to those of titanium oxide. Specifically, titanium should form a sulphide having a high melting point, indicating a stable compound of low solubility in iron. In order for titanium to function as a desulphurizer or sulphur-fixing agent, it must form a stable sulphide. In order for titanium to be useful in steels as a sulphur-fixing agent, it must react with sulphur in preference to carbon; and to prevent red shortness, the sulphide formed must have solubility characteristics such that grain-boundary precipitation of a low melting phase be prevented. The terms desulphurizer and sulphur-fixing agent when used in this paper mean that titanium effectively ties up the sulphur as a stable sulphide of titanium and prevents the formation of FeS. The partition of sulphur between iron and titanium was studied first in a series of Fe-Ti-S alloys with carbon absent, and then in similar alloys with carbon present. It has been established1 a that titanium forms in Fe-Ti-S alloys a sulphide of undetermined composition which can be identified microscopically, and that the titanium-sulphur compound does not respond to sulphur printing. It has been suggested" that titanium acts as a sulphur-fixing agent, although little proof has been given. A report% f The Advisory Committee on Metals and Minerals summarizes the existing knowledge concerning substitutes for manganese in steels. Distribution of Sulphur Between lron and Titanium with Carbon Absent The alloys for this investigation were made by -melting Armco iron, metallic titanium, and iron sulphide in alundum crucibles in an induction furnace. The titanium was added after the other ingredients had melted. In these alloys it is recognized that some TiO, is present, since the melts were made in the air and since no other deoxidizing agent was used; therefore, some of the titanium was not available for union with sulphur. Thirteen ingots were made. Ingots 2 to 8 were cast in baked-sand molds; ingots 12 and 13 were cast in metal molds; and ingot 9 was allowed to cool in the crucible in which it was melted. All ingots were annealed at 900°C for 3 hr and were cooled slowly in the furnace. Subsequent information indicates that this anneal had no effect upon the titanium sulphide particle size, since the sulphide does not go into solution at this temperature. The ingots as-cast were approximately 8 in. long and % in. diam. The compositions of the eight ingots and the results of the partition of sulphur are recorded in Table I. Total titanium and total sulphur were determined by standard gravimetric methods. Sulphur, as iron sulphide, was determined by the evolution method. Boiling HC1 (1:4) liberates the sulphur from iron sulphide, but leaves the titanium sulphide practically undecomposed. A complete separation of titanium sulphide and iron sulphide cannot be made by digesting these compounds in dilute HC1 since titanium sulphide was found to dissolve slowly. The amount of titanium sulphide decomposed will depend upon the time in contact with the acid and upon the particle size of the sulphide. The point

Jan 1, 1952

By H. T. Bowman

FOR more than 80 years the Bessemer process has depended upon the ability, skill, and judgment of the blower, although as early as the 1860's it was recognized that the process would benefit by some type of instrumental control.1 However, instruments with sufficient recording speeds did not exist, so the end-point determination was always dependent upon the blower's eye. This placed an obstacle in the way of precise experimental work, because of the absence of the necessary records and reference points as reported by Work.2 The use of the photocell provided the instantaneous record of the blow that made this experimental work possible and provided more scientific control for the process as discussed by Graham3' A brief description of the progress of the blow will illustrate the recording function of the instrument. The silicon in the iron is being oxidized during the first minutes of the blow, as pointed out by Henning4 in a table showing the order of metalloid elimination. The product of this oxidation is a solid that remains in the converter, so there is virtually no flame during this part of the blow. This low flame luminosity is illustrated by the flame recording shown in Fig. I, where the silicon blow is plainly marked. The dull flame of the silicon blow gradually gives way to the brilliant carbon flame as the increasing quantities of carbon monoxide formed inside the converter burn to dioxide upon contact with the atmosphere above the vessel. When the carbon reaches some low value under 0.20 per cent, there is insufficient monoxide formed to maintain the dioxide flame, so its intensity drops. Throughout the flame drop carbon continues to burn, until at some value around 0.05 per cent an arrest occurs in the recorded drop. This point has been designated as the end point. The period of time from this end point until the vessel is turned upon its side and blowing has stopped has been called the afterblow. This afterblow is used as a measure of the degree of metal oxidation, and uniformity demands that it be controlled within narrow limits. Iron and blowing conditions are likely to change several times during a turn, necessitating corresponding changes in end-point control. It is the purpose of this paper to discuss these changes, the factors that make them necessary, and their bearing upon the quality of steel. CARBON-FEO RELATIONSHIP In the Bessemer process oxidation relationships are precise. Fig. 2 shows a graph of the carbon-FeO relationship. The familiar "bomb-test" method, described by McCutcheon and Rautio,5 was used for obtaining these and all other dissolved FeO values herein reported, all tests being taken from the converter before deoxidation unless otherwise noted. All tests of 0.10 per cent carbon and above were obtained from metal blown for duplexing. McGinley and Woodworth6 reported iron oxide values for blown metal, but their tests were obtained by a different method than were those

Jan 1, 1942

By J. B. Hess, R. L. Dietrich

IN the mechanical twinning of magnesium on the {1012} planes the crystallographic deformation is such that, in the direction of the hexagonal axis [0001], twinning is possible only under tension stress, while in the directions perpendicular to the hexagonal axis, twinning is possible only under applied compression. Thus in extruded magnesium and magnesium alloys where the extruded textures are best described as either [1010] or [ 1120] fibers,1 the basal planes are all nearly parallel to the extrusion axis and such material is susceptible to {1012 } twinning by compression in the direction of the extrusion axis. In order to explain the results of cyclic stressing in alternate tension and compression upon the tangent elastic modulus in such material, Dorn and Thomsen2 have postulated that, following twinning under applied compression, a retwinning or untwinning occurs as a result of residual microtensile stresses immediately upon removal of the applied compressive force. Recently, Carapella and Shaw3,4 have used this supposition of untwinning in magnesium in discussing their results on indentation hardness and on cold drawing of MI sheet. In trying to confirm this hypothesis by suitable direct experiments, the authors have been unsuccessful in detecting any evidence whatsoever that untwinning follows the release of applied compression and are forced to conclude that this hypothesis is improbable or, at least, that the quantity of material involved is small. For the experiments, small cylindrical specimens 0.040-in. diam. X 0.120- to 0.140-in. long were etched from 0.0625-in. extruded rod. The ends of the specimens were accurately faced off parallel in a jeweler's lathe. The compression jig was simply an ordinary set of micrometer calipers with an anvil interposed between the spindle and the specimen. The anvil was free to move axially but restrained from rotating so that the rotation of the spindle would not introduce a twisting stress on the specimen. With this arrangement the compressive strains could be estimated to 0.0001 in. and X ray pinhole photographs could be taken with or without load as required. Fig 1 is the pattern of an uncompressed specimen of AZ31 alloy, showing the approximate [1010] extrusion fiber. The X ray beam is normal to the extrusion axis; in this position it is particularly sensitive in detecting twinning if one directs his attention to the diffraction ring from the basal plane. Since twinning on {1012} planes produced an 86° reorientation of the crystal lattice, the original basal reflections occur at the 3 and 9 o'clock positions while the (0002) reflections from any twinned material will occur at the 6 and 12 o'clock positions. Fig 2 shows the same specimen compressed 3.0 pct, photographed while still under load. This strain has been just enough to produce a detectible amount of twinning and the sensitivity is roughly indicated by the fact that no alterations were produced in the intensities or maxima of the other diffraction rings. A strain of

Jan 1, 1948

By E. H. Burdick

THE difficulties inherent in table concentration operations as applied to gold, silver, lead and zinc ores, are accentuated in the scheelite mill, which has a flowsheet that is similar in general principles. The narrow "spread" in the specific gravity between scheelite and the usual gangue in which it occurs often makes its clean separation extremely difficult and in some cases impracticable. The problem is further complicated by the extreme friability of the mineral and its inordinate tendency to "slime," particularly where fine grinding is necessary and wet grinding methods are employed. The physical aspect of the occurrence of scheelite in the ores, the kind and character of the rock in which the mineral is present, its hardness, friability and the specific gravity of its constituents, as well as like characteristics of the associated mineral-spyrite, pyrrhotite, sphalerite, and others, all have a bearing on the successful milling of crude ores and the production of a concentrate that is marketable with minimum price penalty or none at all. Most tungsten deposits are considered low grade when the limiting factor of low percentage recovery by gravity methods is borne in mind. The great majority of deposits developed or in process of development carry from 0.4 to 1.0 per cent tungstic oxide. Ore carrying 1 per cent is considered good grade; 0.7 to o.8 per cent, fair, and below that figure, low grade. Steady mill feed averaging upward of 1.0 per cent WO3 is exceptional. Inasmuch as buying specifications demand a 60 per cent WO3 content, the mineral content of the ores as mined must be concentrated to at least that degree. A few companies have constructed chemical plants for the treatment of their complex concentrates and low-grade flotation product, but the smaller operator must rely on table concentration, borrowing the tools and to a large extent the methods of his predecessors in the gold, silver, lead and zinc mills. Whether these are the best tools and methods that can be devised is an open question. Except in rare instances, table concentration, however carefully conducted, does not yield a satisfactory final recovery of the mineral content in headings as indicated by chemical determinations of tungstic oxide content. A saving of 6o to 65 per cent is considered fair, and an average 80 per cent recovery, exceptional. Each scheelite ore body is a separate problem. Details of handling that give best results in one case may not be successfully operative at a property on the same contact zone and but a short distance removed. Supplementary treatment of tailings by flotation may be successful in producing a low-grade concentrate of around 15 per cent W03, which is adapted to further treatment chemically at a given property and entirely unsuccessful at another. This problem constitutes another of the unanswered questions. The following paragraphs relate solely to the beneficiation of scheelite ores by table concentration and supplementary methods.

Jan 1, 1942