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By Myron Read

DURING the past 25 years, mining engineers have seen the development of a multitude of specialized engineering curricula in the mineral industry field. Bachelor degrees are now !ranted in the fields of mineral dressing petroleum engineering, geological engineering, fuel technology, and many others. These fields were all, previously considered to be part of a mining enginieer's work and elementary phases of them were included in mining engineering curricula. Their emergence has left some mining engineers with the fear that nothing will he left for a purely mining engineer to teach, or that mining engineering graduates will be unable to complete in any of these fields with the more specialized students and thus will have neatly restricted employment opportunies. A secondary reaction has been that if students who began to say that the pining engineering curriculum taught everything but mining, since so many of :he subjects that had always been udedAuded in the mining engineering curriculum now were fields in their own right.

Jan 1, 1946

By R. I. Jaffee, H. R. Ogden, D. J. Maykuth

IN THE past year, Jaffee and Campbell&apos; and Finlay and Snyder2 reported on the mechanical properties of titanium-base alloys, some of which were in the same ranges of composition as are covered in this paper. In this paper, evidence confirming that given by Finlay and Snyder on the effects of carbon, oxygen, and nitrogen on titanium will be presented; and, in addition, new data will be given on the effects of these elements on the flow properties and phase transformation of titanium. Materials and Preparation of Alloys The preparation and general properties of iodide titanium have been adequately described elsewhere.&apos; , As-deposited iodide titanium rod, prepared at Battelle, of Vickers hardness less than 90 was employed as the base metal in the present work. This was the same material as that used by Finlay and Snyder.2 The probable analysis reported by them for standard quality metal holds here also: N 0.005 pct, 0 0.01 pct, C 0.03 pct, Fe <0.04 pct, A1 <0.05 pct, Si <0.03 pct, and Ti 99.85 pct. Carbon was added in the form of flake graphite supplied by the Joseph Dixon Crucible Co. Oxygen was added in the form of c.p. grade TiO, powder, produced by J. T. Baker Chemical Co. Nitrogen was added in Ti3N4 powder, supplied by the Remington Arms Co. Individual ingots weighed 7 or 8 g. Carbon, oxygen, or nitrogen was added by placing the corresponding powder in a capsule made from as-deposited iodide titanium rods and melting the capsule with the balance of the charge. The charge was are-melted with a tungsten electrode on a water-cooled copper hearth under a partial vacuum of very pure argon (99.92 pct minimum). Melting was practically contamination free. Vick-ers hardness increases of less than 10 points were normal for unalloyed iodide titanium control melts. Nitrogen analyses of are-melted iodide titanium showed a nitrogen content of 0.005 pct, about the same as is present in the as-deposited rod. No tungsten pickup was found in a melt of iodide titanium analyzed for tungsten. Weight losses in melting nitrogen-free alloys were very small and varied consistently from nil to 0.015 g (0 to 0.2 pct). This permitted the use of nominal composition for these alloys. Chemical analyses made for carbon, which can be analyzed conveniently by combustion methods, justified this procedure. Where nitrogen was added, considerable splattering took place. Here it was necessary to analyze for nitrogen by the Kjeldahl method. The ingots were hot rolled at 850°C to about 0.045 in. thick. After hot rolling, the strips were descaled by mechanical grinding, and then given a cold reduction of 5 to 10 pct to insure a uniform thickness throughout the length of the specimen. The edge strips and the tensile strips were annealed in a vacuum of 1x10-4 mm Hg pressure for 3 1/2 hr at 850°C and furnace cooled. Methods of Investigation Hardness Measurements: At least five Vickers hardness measurements were taken using a 10-kg load on each sample in the following conditions: (1) top and bottom of each ingot, (2) top and bottom surface of as-rolled and annealed sheet, and (3) on cross-section of annealed sheet and all quenched specimens. Tensile Tests: Tensile tests were conducted on Baldwin-Southwark testing machines having load ranges of 600 or 2000 lb. Tests were made on 1-in. gauge-length specimens, 3 1/4-in. overall length, 1/2 in. wide, 0.040 in. thick, with a reduced section 1 1/4 in. long and 0.250 in. wide. Two SR-4, A-7 strain gauges, one mounted on each side of the specimen, were used to measure the strain over a limited range to determine the modulus of elasticity. After the modulus of elasticity readings had been taken, load vs. strain readings were taken, using only one strain gauge, at increments of 0.0001 in. until the yield points were passed and then at 0.001-in. increments to the limit of the strain-gauge indicator (0.02 in.). Strain readings above 0.02 in. per in. were taken every 0.01 in., using dividers to measure the strain between the 1-in. gauge marks until the maximum load had been reached. Crosshead speed, when using the SR-4 gauges, was 0.005 in. per min, and, when using dividers, 0.01 in. per min. Flow Curves: Flow curves were determined using the true stress-true strain data obtained during the tension test. The usefulness of this type of information has been dealt with very adequately elsewhere by L. R. Jackson,&apos; J. H. Hollomon,6 and many others. Flow curves of true stress vs. true strain could be converted to the more conventional cold-work curve of 0.2 pct offset yield strength vs. percentage of cold reduction by means of the transformation, 1/1 = 1/1-R, where R is the fraction reduction in cold working. Thus, the true strains corresponding to percentage reduction can be calculated, and the 0.2 pct offset yield strengths scaled off the — 6 curve by taking the true stresses corresponding to the values of 6 + 0.002 strain. Heat Treatment: For the transformation studies, the alloys were heat treated in a horizontal-tube furnace using a dried 99.92 pct argon atmosphere, and quenched into water. Essentially no contamination was found after several hours of heat treatment at temperatures up to 1050°C. Metallography: Specimens were prepared in the

Jan 1, 1951

By Jacques Steininger, Robert E. England

Single crystals of ZnTe and ZnTe1-,Sex with x up to 0.13 have been grown from the elements by temperature gradient solution zoning using excess tellurium as a solvent. Best results have been obtained with charges with the compositions 45/55 at. pct Zn, Te, for ZnTe and increasing amounts of selenium for ZnTe1-xSex. The temperature in the molten zone was maintained at about 1070°C with a gradient of about 10°C per cm. Chemical analyses of quenched ZnTe ingots show tellurium concentrations in the molten zone as high as 70 pct with concentration differences across the zone of 1 to 2 at. pct Dark dots which are observed by transmitted light microscopy in as-grown crystals can be removed by annealing in zinc vapor at 900 C. INTEREST in wide band gap semiconductors has led to a new study of ZnTe and ZnTel-xSex crystal growth. ZnTe is the only II-VI compound with a wide band gap (2.3 ev) that can be made p type with low resistivity. Attempts to make it n type with low enough resistivity to be useful for p-n junctions have so far been unsuccessful.1 ZnSe has a band gap of 2.65 ev but can be made n type only. However, ZnTel-xSex solid solutions with x as low as 0.36 have been made both highly n and p type2 with a minimum band gap around 2.12 ev3 at room temperature and appear to hold the best promise for efficient injection electroluminescence in the visible. ZnTe has the lowest melting point of the zinc chal-cogenides (1295°C) and consequently attempts have been made to grow crystals from both the liquid and the vapor phase.4 Complicated apparatus is required for growth from stoichiometric melts because of the high vapor pressures of the elements at the melting point of ZnTe and because of the problem of quartz devitrification. Small crystals have thus been grown in high-pressure equipment by Fischer5 and by Narita et a1.6 with pressures of the order of 50 atm of argon to prevent excessive evaporation from the melt. Large crystals of ZnTe can be obtained by growth from the vapor phase4 but they often present numerous dislocations and inclusions. An improvement in the quality of vapor- grown ZnTe crystals was reported by Albers and Aten7 by equilibration of mixtures of small crystals with compositions lying on either side of the solid single-phase field at fixed temperature. The same technique was later applied by Aten8 to the growth of ZnTe1-xSex crystals with less than 1 pct inhomo-geneity. Because of the higher liquidus temperatures of the solid solutions and the high vapor pressure of selenium, previous attempts to grow ZnTel-xSex from the melt have been limited and unsuccessful.9 The phase diagram of the Zn- Te system is reproduced in Fig. 1, based on data from Kobayashi10 and Kulwicki.11 Carides and Fishher12 have reported lower liquidus temperatures on the tellurium-rich side, but their data would require confirmation. The liquidus temperature on the tellurium-rich side decreases rapidly with increasing tellurium concentration and the Te2 vapor pressure over the liquidus also decreases accordingly.&apos;3 The decrease in liquidus temperature and vapor pressure therefore makes it possible to use conventional apparatus if there is a sufficient excess of tellurium in the melt. Single crystals of ZnTe have thus been grown by Kucza,14 in a modified Bridgman technique, from solutions containing up to 60 at. pct of Te by lowering unsupported quartz ampoules through a temperature gradient at about 1200°C. Under these conditions, the phase diagram indicates that the entire charge is initially molten. Crystal growth can therefore proceed by normal freezing and rejection of excess tellurium into the melt. The modified Bridgman technique has several major limitations. Because of the rejection of excess tellurium into the melt during freezing, the melt composition and the temperature at the growth interface vary continuously. They tend to follow the liquidus until the eutectic which is very close to pure tellurium (447°C, >99 pct Te). Since the solidus composition also varies with temperature,15 crystals grown by this method are inhomogeneous. They present small variations from stoichiometry which may affect their structure and physical properties. The simultaneous increase in tellurium content and decrease in melt temperature also combine to reduce the rate of diffusion of tellurium away from the growth interface, thereby causing constitutional supercooling and possibly dendritic growth. To minimize these effects, the initial melt composition is in practice kept relatively close to stoichiometry (less than 60 pct Te). This however limits the possibilities of operating at low temperatures and pressures. This paper describes a modified method of crystal growth by temperature gradient solution zoning (TGSZ) which is an adaptation of the temperature gradient zone-melting technique developed by pfann16 and of the traveling solvent method of Mlavsky and weinstein.I7 The TGSZ method now applied to the growth of ZnTe and ZnTel-xSex crystals is characterized by its very simple experimental arrangement and sample preparation technique. Unlike the modified Bridgman technique, there is no increase in the tellurium concentration in the melt and therefore it is possible to operate at lower temperatures and pressures. This method is also suitable for maintaining a constant temperature at the growth interface.

Jan 1, 1969

By Felix A. Vogel

I (New York Meeting, February, 1912.) FLUE-DUST, to most blast-furnace operators, means a troublesome by-product, the formation of which should be curtailed, if not prevented entirely. However, with the increasing use of fine ores, larger furnaces, and high-pressure blast, the production of flue-dust is constantly increasing, and amounts annually in the United States to from 3,000,000 to 3,500,000 tons, An exceedingly high tonnage, of which a large part has been discarded as valueless. As a result of greater economy in the iron industry, the attention of our furnace-men has been directed towards the utilization of this enormous amount of waste material, a problem which had also been given due consideration by metallurgists abroad. Flue-dust is generally a fine material containing considerable coke and iron-ore, with a small admixture of lime and silica, depending upon the burden. The iron-ore is partly reduced, which shows that the dust originates largely in the reducing-zone of the blast-furnace. In the United States this dust usually contains 20 per cent. of coke and more than 40 per cent. of iron. Estimating coke to be worth $3.25 per ton, and iron-ore 7 cents per unit, a ton of flue-dust, unless made available, represents a loss to the furnace-man of $3.50. This accounts for the first efforts to recharge the flue-dust into the furnace, either by moistening it clown with an excess of water, or mixing it with clay to form balls of pulp, or treating it with lime-water. These methods, however, have been practically discarded, as they failed to produce the desired economies. To recover, in the blast-furnace, all the values represented by the material contained in the flue-dust, the following con¬ditions should be complied with

May 1, 1912

By R. E. Smallman

R. E. Smallman (University of Birmingham, England)—Recently, Tedmon and Ferriss11 have determined the yield stress parameters oi and ky for tantalum by measuring the lower yield stress as a function of grain size 2d and fitting the results to a relationship of the form They report that although ky , which is taken to be a measure of the dislocation locking strength, is small (- 2 to 4 x 106 cgs units) a substantial yield drop is nevertheless observed in a normal tensile test. Niobium gives a similar result,12-14 as pointed out in the original work by Adams et a1.,12 and in order to check this apparent anomaly the yield-stress parameters of electron beam-melted niobium have recently been reanalyzed15 by the Luders strain technique. In this method the strain hardening part of the stress-strain curve is extrapolated to zero plastic strain; the intercept on the preyield portion of the curve is taken to give oi, whilst the difference between oi and the lower yield stress gives kyd-1/2. The results indicate that ky increases with increasing grain size and hence, a plot of vs d-112 yields an apparent ky, which is lower than the true value. A similar effect could account for the small ky found in the relatively pure tantalum used by Tedmon and Ferriss. The variation of ky with grain size shows that dislocations are more strongly locked in coarse-grained specimens than in fine-grained samples. In niobium, this may be attributed to the fact that the dislocation density in the fine-grained material is higher than that found in the coarse-grained samples which are given a sufficiently prolonged anneal to remove any residual substructure and, since the metal contains only a small amount of interstitual impurity, a variation in locking occurs. By contrast, application of both the grain size analysis and the Luders strain method to yield-stress data from commercially pure vanadium containing a large amount of interstitial impurity gives consistent values of oi and ky, with ky independent of grain size and temperature. Electron microscope observations show minor variations in dislocation density from grain size to grain size, but in any case in this material the dislocations are heavily locked with precipitate. On yielding new dislocations are generated and, as a consequence, the importance of any differences in dislocation density between the various specimens of different grain size is considerably reduced. It is perhaps significant that Adams and lannucci,16 working with a grade of tantalum containing a higher interstitial content than that used by Tedmon and Ferriss, prepared the specimens of different grain size by annealing in the temperature range 1500" to 2000° C to minimize any differences in dislocation structure, and found that ky had a value of 1.04 x 107 cgs units, independent of testing temperature. Such behavior is consistent with the dislocations being locked by carbide precipitates so that the generation of free dislocations is an athermal process. The recent work of Gilbert et al.17 also shows that in tantalum there is no significant variation of ky with grain size provided it contains 150 ppm of oxygen. In this case, however, the dislocations are not locked by precipitate and ky is temperature dependent. C. S. Tedmon and D. P. Ferriss (authors&apos; reply)— We would like to thank Dr. Smallman for his interesting comments and discussion to our paper, "The Dependence of Yield Stress on Grain Size for Tantalum and a 10 pct W-90 pct Ta Alloy".18 It was suggested that perhaps the relatively small values obtained by us for ky of tantalum could be attributed to the same cause that accounts for the apparently small values of ky that result when it is determined by the Luders Strain technique. Since our values were obtained by plotting the lower yield stress vs the reciprocal of the square root of the grain size, it is not clear how this could be the case. The values of ky in this experiment have been calculated, using the Luders strain technique. With this method, values for ky on the order of 2 x 105 to 5 x lo6 cgs units were obtained. In spite of this rather large variation, the magnitudes are still small, and there appeared to be no good correlation between ky and the grain size or the yield stress, probably because of the difficulty in accurately extrapolating the work-hardening portion of the curve back to zero plastic strain. As was shown in the original data,18 there was little work hardening in any of the curves, at any temperature. In his discussion, Dr. Smallman also points out how ky has been observed to increase with increasing grain size, when determined by the Luders strain technique. There are at least two possible explanations for this. In the first case, if it is assumed that the bulk of the interstitial impurities are concentrated at the grain boundaries, then, of course, the available grain boundary area would decrease with increasing grain size, thus presenting less area for the interstitials, which would then presumably increase the concentration within the grains, thereby increasing the locking of the dislocations. In the second case, the increase in ky with increasing grain size would be attributed to the nature of the grain boundary itself. One of the several ways of deriving the Hall-Petch equation19 is based on the stress concentration arising from a pile-up of dislocations at the boundary. The ability of the stress concentration to unlock a source in a neighboring grain would depend on the strength of the grain boundary. As is well-known, the nature and struc-

Jan 1, 1963

By Arthur Knapp

NEARLY every technical man is called upon at some time in his life to deliver a paper before a technical audience or to write a technical paper for publication. It is not necessary to be an accomplished public speaker to deliver a technical paper successfully; it is only necessary to bear in mind a few obvious points. Poorly presented papers are nearly always the result of a few errors of deportment, but they are vital and at the same time so obvious that they are often overlooked. Experience is a good teacher but a little study and careful preparation will moderate the suffering of both an amateur speaker and his audience. An author&apos;s paper suffers, through poor delivery, in direct proportion to the suffering of his audience. The best test of a paper is to have it read aloud to the author. This will bring out faults not apparent when read silently by the author. Some of the things which will probably need correcting are improper phrasing, noneuphonic sequences, and long or involved sentences. Frequently authors write their papers in short crisp sentences for oral presentation, then rewrite them for publication.

Jan 1, 1942

By C. F. Allen, G. B. Walker

K. F. TROMP*—In dealing with the question of the most suitable kind of solid media for heavy density suspension processes Walker and Allen point out that the particle size of the solid media should not be taken too fine, as the viscosity increases with the area of the solid media and a low viscosity is essential lor high tonnage and accurate separation. A coarser particle size of the solid media will, in their opinion, of necessity give rise to a differential density in the bath (higher gravity at the bottom of the bath than at the top) but they advocate acceptance of the differential density rather than a higher viscosity. Though I fully agree with the choice the authors have made, I cannot subscribe to their view that only by accepting a differential density in the bath a coarse particle size of the solid media can be used. There certainly is another alternative: stronger agitation. Applying sufficiently strong vertical currents, a uniform gravity can be obtained quite well in a suspension of a coarse solid media. Of course, this is not a very attractive solution, for it means a degradation of the true gravity separation and a step backwards to hydraulic classification, which makes the washing dependent on size and shape of the particles. However, to a greater or lesser extent, this is what actually takes place in all the heavy density suspension processes relying on a uniform gravity in the bath. The so-called "stable" suspension processes make no exception. They all "stabilize" their suspensions by introducing or creating vertical currents, be it upwards or downwards or both, be it by hydraulic or by mechanical means. In fact, there is no such thing as a "stable" suspension in gravity separation, as the very reason for the use of suspensions in this field is the property that the solid media is able to settle and so facilitate the recovery. I have been enlarging on this point because the characteristics of the various processes can only be well understood and viewed from the same angle (from Bar-voys up to Chance) when the fact is recognized that mechanical or hydraulic agitation is a condition sine qua non for obtaining a uniform density from top to bottom in a suspension. Is a Cone-slraped Vessel Essenlial? Of the two alternatives for getting a low viscosity Walker and Allen have preferred correctly the sacrifice of uniform gravity in the bath instead of increasing further their vertical current arid agitation. The resulting differential density of the bath brings the problem of bow to prevent accumulation of intermediate gravity products in the bath, an accumulation which, if not prevented, would ultimately plug their cone. According to the authors an open-top cone combined with a downdraft current of the bath liquid would he the only suitable way to cope with such suspensions and they assume as a fact that "in any vessel other than a cone, such a differential density could not be tolerated." My experience is quilt: different. In my process, which has been in successful operation for more than a decade, differ-ential density of the suspension is applied ranging from values below 0.1 up to differentials above 0.5, according to the prevailing requirements of the individual plant. In this process, which is charac-terized by the use of horizontal currents in a suspension of differential density, the form of the vessel is of secondary importance and different types are in operation. It so happens that none of these are in the, form of a cone. The fact that 24 washboxes on my process have been installed and 12 others are under construction may constitute sufficient proof against the opinion that only a cone-shaped separator would be suited for differential density separation. Horizontal Currents in Differentia1 Den-sity Sepparation I myself have some doubts as to the suitability of a cone with downdraft for dealing with differential density (or, for that matter, any other washbox relying on vertical currents for removing the intermediate gravity products). It ap-pears to me that it is restricted to feed of small size only and even then with watch-fulness. If we take, for example, a piece of 2 in., the draft necessary to pull such a piece down to a zone wherein the den-sity of the suspension is, say, 0.03 higher, is quite considerable. For a suspension of, say, 1.6 sp gr the downdraft will have to be in the region of 3 in. per second. Unfortunately. most of the differential in density is in the part immediately below the reach of the top current which transports the floats. Consequently, we need the downdraft where we like it least: in the upper part of the cone. This entails the risk that light float particles are carried away with the downward current. This current of, say again, 3 in. per second would carry particles up to 1.3 sp gr and 3/8 in. size into the 1.6 gravity zone. This is prohibitive. It is also prohibitive because a downdraft of 3 in. per second in the upper part of the cone would require a tremendous circulation of medium. IIalf way up a 20 ft diam cone, a downdraft of 3 in. per second would correspond with 8500 gpm. With the downward current following the way of least resistance, the strength of the downdraft will not be exactly the same at different places of a cross area. If, as I anticipate, the center of the cone is favored, the strength of the downdraft will fall below the critical value near the

Jan 1, 1950

By A. H. Larson, R. J. McClincy

Laboratory-scale decopperizing experiments with multiple sulfur addifions were conducted at 330°C on ternary Pb-Cu alloys containing, as the third elenlent, Sn, Ag, As, Sb, Bi, Zn, and Au, common impurities in lead blast-furnace bullion. For silver and tin, an increased rate and extent of &apos;cofifier removal was obsert3ed. The elements As, Sb, Zn, Au, and Bi had no effect or less effect as compared to sulfur additions with no i)npurily additions. THE production of primary lead in the blast furnace yields an impure lead frequently containing such impurities as copper. antimony. arsenic. tin, gold, silver iron, oxygen. and sulfur. By cooling this lead to a temperature near its melting point. most of the iron, sulfur, and oxygen and part of the other impurities are removed in the form of a dross. With incipient solidification of the lead, the copper concentration wil have been reduced to 0.02 to 0.05 pct. depending upon the concentration of the other impurities. according to Davey.&apos; Since copper interferes with the treatment of silver after the desilverizing process, it is desirable to decrease the copper content of the lead still fur-ther before the lead is desilvered. The decopperizing of the lead is accomplished by stirring a small quantity. approximately 0.1 pct. of elemental sulfur into the lead at a temperature near its melting point, 330" to 360°C. The copper is removed as a copper sulfide which constitutes a small fraction of a voluminous dross consisting mostly of lead sulfide and entrained metallic lead. The residual copper concentration following the decopperizing operation is frequently as low as 0.001 to 0.005 pct. Thi fact has aroused considerable interest because the equilibrium copper concentration of lead in contact with solid PbS and solid Cu2S is at least an order of magnitude greater, 0.05 pct Cu at 330C. 1, 2 Most investigators have suggested that various impurities in the lead bullion are responsible for the very low copper concentrations frequently encountered in practice. There is little agreement, however? as to which of the impurities are helpful and which are not.3"11 Also. few investigators have sought to explain the mechanisms responsible for the removal of copper to very low concentrations. Willis and Blanks9 have proposed that a nonstoichiometric copper-deficient cuprous sulfide forms in place of the supposed Cu2S. Being copper-deficient, this sulfide phase would possess a low copper activity, and the diffusion of copper dissolved in the liquid lead into this phase would be greatly facilitated. Pin and wagner2 have investigated the removal of copper from liquid lead by studying the effect of impurity-doped lead sulfide on the decopperizing of pure Pb-Cu alloys. Samples of the doped PbS were held in contact with copper-saturated lead for 1 week at 33&apos;7°C. They reported a beneficial effect on decopperizing with bismuth and antimony and no effect with tin or silver. which is directly opposite to the results observed in practice and those reported by Davey 3 and this studv. The purpose of this paper is to describe the effects of certain additive elements on the extent to which copper can be removed fro111 liquid lead by successive additions of sulfur. The impurity elements were added individually to prepared Pb-Cu alloys. The resulting ternary alloys as well as a binary Pb-Cu alloy were then decopperized with repeated additions of sulfur. EXPERIMENTAL Materials. Granulated test lead with a purity of 99.999 pct and the additive elements Cu. Ag. Sb. Bi. Zn. Sn. and Au with purities of 99.99 pct were American Smelting and Refining Co. research-grade materials. The major impurities in the lead were 1 ppm each of iron and copper. all others being less than 1 ppm. The arsenic used was a technical-grade arsenic of 98+ pct purity. Reagent-grade flowers of sulfur were melted under argon to provide small pieces free of fines. Apparatus. The decopperizing experiments were carried out in a 25-mm-OD by 375-mm-long Pyrex tube sealed at one end. The tube was mounted vertically in a resistance-heated. hinge-type tube furnace controlled to within ±lcC. Temperature measurement was accomplished by means of a standardized chromel-alumel thermocouple sealed into the base of a silica. paddle-type stirring rod. All decopperizing experiments were carried out under an argon atmosphere. Procedure. A Pb-Cu starting alloy containing 0.05 pet Cu was prepared under carbon and poured into cold tap water to produce shot. The ternary alloys were prepared by melting together 100 g of the starting alloy and a sufficient amount of the impurity element to yield the desired concentration. The resulting alloy was then homogenized in a Pyrex tube at 450C with continuous stirring. The furnace temperature was then lowered to the operating temperature of 330°C. When thermal equilibrium had been obtained at the operating temperature, individual additions of 0.2 pct (0.2 g) of solid sulfur were added to the melt and stirred in. Stirring was continued for a period of 3 min. discontinued for 5 min. and resumed for the remaining 2 min of a 10-min cycle. This cycle was repeated for as many sulfur additions as desired. When the decopperizing experiment had been completed the lead bullion was quenched and samples of the bullion and dross phases were taken for analysis. Results. The results obtained in the decopperizing

Jan 1, 1970

By LESTER H. WOOLSEY

IT IS doubtful whether anything new can be said upon this subject and, therefore, it is with considerable hesitation that l prepare this paper. On account of my recent connection with the Department of State as its law- officer, perhaps I will be pardoned if I discuss this subject briefly from the point of view of that branch of the Government. , I cannot, of course, speak for that department, since I have-no connection with it at-present, but it may be helpful if I set down certain* of the factors which in my opinion and from my experience have a bearing on this international problem. With oil-burning ships, motor vessels, motor vehicles, airplanes-all consuming, astonishing quantities of petroleum or petroleum products; with the United States having within its borders only 12 per cent. of the petroleum resources of the world, and consuming about 70 per cent. of the world&apos;s output; with the more enterprising nations wide awake to the necessity ; of obtaining and preserving as large a share as possible of the oil resources extant ; the &apos;problem is peculiarly and in a, very practical sense an international one.

Jan 1, 1921

By John F. Elliott, Martin Weinstein

The solubility of hydrogen in liquid Pure iron and in a number of liquid binary iron alloys has been measured in a Sieverts&apos;-type apparatus. Sieverts&apos; law is obeyed in all alloys studied UP to 1 atm Pressure H2. The alloying elements, Al, B, C, Co, Cu, Ge, P, S, and Sn, decrease the solubility while Cb, Cr, Mn, and Ni increase it. ALTHOUGH the importance of hydrogen in steel-making has long been recognized, there has been, until recently, very little information in the literature on the effect of alloying elements on the solubility of hydrogen in iron. There is also disagreement on the solubility of the gas in liquid pure iron. Because of the earlier work in this laboratory,1,2 with the Sieverts method to determine the solubility of nitrogen in liquid iron alloys, it was decided to apply the equipment and procedure to hydrogen. The literature that is pertinent to this study (Ftefs. 1 through 15) will be discussed later when the results of this investigation are considered rather than at this point to avoid repetition. The reaction under study is:

Jan 1, 1963

By O. W. Ellis

IN VIEW of the extensive use of the brasses and bronzes in engineering practice it is indeed surprising that so little scientific work has been done on the oxides in these alloys. Recognition of the ill effects of these compounds is universally accorded, both in foundry procedure and in such specifications as cover the melting of these useful mixtures. In the latter, limits, for example, are placed on the proportions of scrap metal in the charge, and clauses governing the use of fluxes are included, the addition of fluxes to the charge being made with the object of avoiding oxidation during the process of melting and of subducing and separating non-metallic matter from the melt.1 In the manufacture of brass it occasionally, though rarely, occurs that the virgin metals, copper and zinc, alone are used. In the event of virgin metals being employed, that of the higher melting point is always melted first. If it is melted under suitable conditions, little or no oxidation will occur. Whether oxidation has occurred or not, however, can best be determined by microscopic examination of a sample taken from the melt.

Jan 1, 1930

By Georges Ranque

THIS paper by. Georges Ranque is an attempt to reason out the conditions of formation and stability of inclusions and to establish an operative process capable of eliminating or minimizing them. The paper consists of two parts; one in which the author considers the laws of physical chemistry and physics that apply to liquid slag-metal systems and one in which he describes a practice based on the theoretical considerations. Such a paper is bound to contain controversial points and it is therefore your speaker&apos;s intention to present Mr. Ranque&apos;s side of the picture to the best of his ability without regard for any thoughts or criticisms of his own. RISING VELOCITIES OF INCLUSIONS Most nonmetallic inclusions have densities ranging from about 1.7 to s, while liquid steel has a density of about 7. Under these conditions, ferrostatic pressure should cause the inclusions, being lighter, to rise from the bath and be eliminated. This rising tendency is expressed by Stokes&apos; law, from which may be calculated the rising velocity of a spherical particle in a liquid medium whose density and viscosity are known.

Jan 1, 1947

By E. P. Mathewson

W. W. (Pop) Adams&apos; contribution is short and to the point. "I went to Pueblo about March, 1881, and started working for the Pueblo Smelting & Refining Co. in the sampling room, going from there to the assay and chemical departments, and ending in charge of the lead refinery on the night shift. I was there about two years, returning to California in 1883. The plant was under the management of Mather and Geist, plus Charley Ballard; Herman Schlapp was the metallurgist. Herman Schlapp&apos;s name remains prominently in my recollection of Pueblo for his never-failing kindness and charity to the young men who were trying to learn the Pb. game. John Allen, Walter Koehler, George Westcott, Charley Molson, and Alfred de Ropp were there with me. Schlapp and Koehler I have never seen since I left Pueblo; the others I have met often. , Alfred de Ropp I have often met-here, as he was manager of the Selby Smelting & Lead Co. of San Francisco, for years. &apos;He tells me that I gave him the rudiments. of American profanity at Pueblo; an apt pupil, for since then, by his own efforts, he has added largely to his vocabulary..

Jan 12, 1923

By Nelson Severinghaus

Consolidated Quarries Corp. must conduct operations for an average sales price of $1.25 per ton, about the same price at which stone was sold 25 years ago when the dollar was worth twice what it is now. To maintain success in a competitive business, the company must solve the usual production problems, which will be discussed here in the order they occur in the flowsheet. Most crushed stone quarries must first remove soil and decomposed rock. Although seasons of lower sales sometimes release regular quarry excavating and haulage equipment for this purpose, overburden is often too extensive for this method to be used. In such cases, special equipment must be assigned or the job may be contracted. As many earthmoving contractors have heavy duty specialized equipment well suited to overburden removal, they should be seriously considered before the job is undertaken. Pockets of clay beneath peaks of hard stone sometimes defy mechanized removal. Pick and shovel laborers at this point might be supplanted by more ingenious use of pumps and water or some other mechanical means. Where overburden is extremely heavy or difficult to remove, underground mining must be considered.

Mar 1, 1956

By T. E. Brower

SHORTLY after our previous paper on this subject was printed,1 we located a source of uncertainty in the results arising from the unexpected fact that hydrogen slowly reduces silica at 1100° C. in presence of iron, even when the iron nowhere touches the silica, although in absence of iron it does not do so. This difficulty has been completely obviated by the use of a small high-frequency furnace, which enables us to heat the specimen for analysis without raising the temperature of the silica vessel to a point at which any appreciable reduction of the silica glass occurs. We now describe this modification of procedure, and present some typical results obtained with it. For any given specimen of steel, the results are entirely definite and. reproducible, but their precise significance is still open to some question, as indeed is true of all methods of oxygen determination so far developed. Further elucidation of this question is to be expected from the cooperative comparison of the several methods now under way under the auspices of the Bureau of Standards at the instance of the Iron and Steel Division of the A.I.M.E.

Jan 1, 1934

By Ross Blake

THE St. Joseph Lead Co. became interested in the Tri-State district in 1921 through acquisition of prospecting and development rights on approximately 20,000 acres of land extending northeastward from the Picher mining field, mostly in Kansas. Prospecting with churn drills was begun the same year and the first commercial ore body was found on the Jarrett tract at the then northern limit of the main Picher field. Operations at this mine began in 1923. Kansas Explorations, Inc., was organized in 1923 by the St. Joseph Lead Co. as a wholly-owned subsidiary to carry on its operations in the Tri-State District from an office at Joplin, Mo. As prospecting of the original acreage proved largely negative, attention was subsequently turned to other sections of the district. Numerous additional leases were acquired and prospected. The Company was successful in developing into operating properties the Jarrett, Isherwood, Robinson, Sullivan and Lotson tracts and, in more recent years, the Ritz, Snapp, Buckingham, Gibson and Jasper mines.

Jan 1, 1947

By R. B. Mears, E. H. Dix, R. H. Brown

In the Institute of Metals Division Lecture of 1940, E. H. Dix, J~.(I)~ discussed the acceleration of the rate of corrosion by high constant stresses. Experimental data obtained on this subject at the Aluminum Research Laboratories and a theory compatible with these facts were included in that lecture. The present paper contains data and ideas subsequently developed at the Aluminum Research Laboratories. Dix pointed out that in order for acceleration of corrosion by constant stress to occur: (1) There must exist in the alloy a susceptibility to selective corrosion along more or less continuous paths, as for in- stance at grain boundaries, and (2) There must exist a condition of high stress acting in a direction tending to pull the metal apart along these continuous paths. Acceleration of corrosion rate by stress is not confined to any one alloy system. Table I lists some of the metals and alloys and some of the corroding environments in which acceleration of corrosion by stress was sufficient to cause cracking (stress- corrosion cracking). It will be noted that the variety of alloys in which stress-corrosion cracking has occurred is rather large. Also, it will be observed that

Jan 1, 1945

By K. C. Li

TITANIUM, because of its high strength, weight ratio, and high melting point, became the metal of the jet age. Zirconium, because of its low neutron cross section and high corrosion resistance to hot water, became an integral part of the nuclear era. Since both metals are very similar chemically, more or less the same process is used for their production. The sudden rise of this new industry of titanium and zirconium seems to have caused some uncertainty as to the future development of processes for making these metals. Although zirconium was discovered in 1789 and titanium in 1790, it was not until 1945, six years after W. J. Kroll patented his process, that F, S. Wartman of the U. S. Bureau of Mines perfected the technique of producing ductile titanium, and it was not until 1946 that Kroll perfected the technique of producing ductile zirconium in commercial quantities. Since then the Government and private industry have spent millions of dollars each year constantly searching for new methods, with the hope that some day titanium can be produced as economically as aluminum and magnesium or as cheaply as stainless steel.

Jan 11, 1957

By Persifor Frazer

The structure of the Caucasus as made out by the Russian geologists and represented in Pamphlet XXII. of the Livret Guide, by Loewinson-Lessing, is an overturned anticlillal from Lars to Passanour; a distorted synclinal with monoclinal dip from Passanour to near the summit of the anti-Caucasus chain south of Tiflis; and another overturned auticlinal from this point to Erivan on the Armenian border. Ng—Neogene; Pg— Paleogene ; J,—Upper Junssic (and Cretaceolls in other parts of the chain) ; J1—Lias ; Pz—Paleozoic schists plicated ; yd—massives of granite and diabase; —porphyries ; —trachytes, andesites; Pz, Mz, plicated sedimentary deposits of Armenia. The details of the geology, however, are extremely complicated—so much so, indeed, as to make it impossible for any set of men to have worked them out in a single season. For &apos;the structure as a whole it will be necessary, therefore, to rely implicitly upon the labors of the Russian Geological Survey, and it is intended here only to state these, and to record a few notes supplementing them. Setting out from Wladikavkaz in the early morning, we pass

Jan 1, 1899

By V. F. Hollister

The South American porphyry copper province includes most of the Andean orogen, excepting the tin province of Bolivia. Porphyry coppers have not been commonly found within the tin belt, though copper-tin deposits in San Rafael, Peru and copper-tungsten deposits in Llamuco, Chile, are known. James" shows the depth to the Moho to appreciably thicken where the tin belt exists, but disclaims any genetic significance from this coincidence. Table I gives some of the most important porphyry copper deposits explored to date, but does not include Antaminas, Peru, nor a number of other large skarn deposits associated with intrusions. V. D. Perry* points out that porphyry copper mineralization does occur to some extent in the intrusion associated with many of these skarn deposits. Also not included are the suspected possible porphyry type occurrences, such as, for example, quartz-sericitically altered lower levels of the Casapalca mine, Peru. Many deposits not well enough documented to list in a summary paper are also excluded. Should anyone care to tabulate all other large copper de- posits in South America, they could demonstrate that a tremendous potential awaits the proper political and economic climate in the future.

Jan 8, 1973