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By G. H. Anderson

During an investigation of the copper-rich portion of the copper-silicon-iron system as part of an extensive research program on P.M.G. alloys, which was begun in 1937 in the research laboratory of the Phelps Dodge Corporation, it became apparent that the binary copper-silicon system below 7 per cent Si needed revision, therefore an investigation was undertaken. Both X-ray methods and microscopic technique were employed. Despite the many investigations that have the clarification of this system as their object, no two successive reports on the system have agreed in placement of the alpha-phase solubility limit, and uncertainty about the secondary phase first to appear has been voiced several times.3,4,5 It was therefore not surprising, when T. Isawa's6 and M. Okamoto's7 works were reported at the time we were finishing up our experimental work, to find essential disagreement between their results and ours. But when C. 5. Smith's latest work on the copper-silicon systema was published just as we were finishing our laboratory report, it was indeed gratifying to find that the latter's diagram and ours checked in all essential features, except the temperature of the peritectoid line above 830" C. Many attempts were made to retain the beta phase during the present investigation, but none was successful. Isawa,6 however, reports that he has been able to obtain X-ray photograms of the beta phase. The structure, he states, is body-centcrcd cubic with a lattiee constant of 2.848 ,&. at 7.2 per cent Si. Okamoto7 and Smith8 both use the designation Kappa for the hexagonal phase, and to avoid confusion this designation will he adhered to in the present paper. Preparation of Samples The alloys were melted in vacuum in a high-frequency furnace. from copper cathodes containing 99.98 per cent Cu, and specially refined silicon supplied by The Electro Metallurgical Co. This silicon is stated

Jan 1, 1940

By Willard L. Hunter

Four solvent extraction systems were studied to determine their efficiency jor extraction and separation of tantalum and columbium. Aqueous feed solutions of varying HF-HCl concentrations and metal content were contacted with equal volumes of cyclohexanone, 3-methyl-2-butanone, and 2-pentanone and solutions of varying HF-H2S04 concentrations were contacted with equal volumes of 2-pentanone. One multistage continuous test was made in a polyethylene pulse column using cy clohexanone as the organic phase. In each system studied, columbium and tantalum purities in excess of 95 pct with respect to each other were obtained in single-stage tests at low acidities in the feed solution. Separation factors ranging from 1700 to 2400 were obtained when rising HF-HCl mixtures in the aqueous phase. Best results were obtained when a solution of HF-H2S04 was used as the aqueous phase and 2-pentanone as the organic phase. A separation factor in excess of 6000 was obtained in one stage with aqueous solution concentrations of 2 _N HF and 2N H2S0,. When acid concentrations were increaszd to 52 HF and 10 _N H2S0,, 99.9 pct of the tantalum and 98.2 pct of the columbium initially present in the feed solution were transferred to the organic phase. The separation of columnbium and tantalum obtainable by means of the solvent extraction systems presented in this paper was found to corn -pare favorably with other systems, including the HF-H2SO4-methyl isobutyl ketone system currently used by most producers for the extraction and separation of these metals. TANTALUM and columbium are always found together in minerals of commercial significance, although the proportion of the two metals in ores varies within broad limits. Columbium is estimated to be 13 times more abundant than tantalum. Five methods generally employed for the separat:ion of these metals are: 1) fractional crystallization (the Marignac process),2 2) solvent separation, 3) fractional distillation of their chlorides, 4) ion exchange, and 5) selective reduction. Of these methods, the one currently used by industry to the greatest extent is that of solvent separation. One of the early technical developments in solvent separation of tantalum from columbium was reported by the Bureau of Mines: the HF-HC1-methyl isobutyl ketone system; data were presented for both laboratory and pilot-plant experimentation.3 Of twenty-eight organic solvents tested for their ability to extract tantalum from an HF-HC1 solution of columbium and tantalum, 3-pentanone (diethyl ke-tone), cyclohexanone, 2-pentanone, and 3-methyl-2-butanone were chosen for further study. Data on the HF-HC1-diethyl ketone system has been published4 and data describing the use of cy clohexanone, 2-pentanone, and 3-methyl-2-butanone as the organic phase are included in this report. RAW MATERIAL The source of tantalum and columbium oxides for this study was ('Geomines" tin slag from the Manono Smelter, Cie Geomines, Gelges, S.A., Congo. In order to extract the valuable Ta-Cb content, the slags were carbided, chlorinated, and the sublimate from chlo-rination was hydrolized and washed free of chloride with water. The washed material was air-dried and stored in a stoppered container. Throughout the paper, "feed material" refers to this mixture of hydrated oxides which was employed because of its high solubility in aqueous solutions. Typical analysis of the hydrated oxides is shown in Table I. I) HF-HC1-CYCLOHEXANONE SYSTEM Batch Separation. Effect of Acid Concentration. To determine the effect of varying the acid concentration upon the transfer of tantalum and columbium, a series of tests was made in which approximately 2.5 g of feed material was added to 25 ml solutions of 2, 4, 6, 8, and 10 N HF and 0 through 5 N HC1. Tantalum pentoxide concentration of the solu%ons was approximately 21 g per liter and columbium pentoxide was 14 g per liter. These starting solutions were shaken with equal volumes of cyclohexanone in 100 ml polyethylene bottles for 30 min. The phases were carefully separated in 125 ml glass separatory funnels. The time of contact of the solutions with the separatory funnels was kept at a minimum to reduce silica contamination. The measured phases were separated into 400 ml polyethylene beakers and the metal contents of each were precipitated by addition of an excess of ammonium hydroxide. Precipitate from each phase was filtered on ashless filter paper, ignited at 800" to 1000°C for 45 min, weighed, and analyzed by X-ray fluorescence.5 Data tabulated in Table I1 and illustrated in Fig. 1, show that maximum separation of tantalum from columbium for each HF concentration was obtained with no HCl present. The purest tantalum product was obtained with some HCl present. The highest separation factor was obtained at 2 N HF and

Jan 1, 1970

By E. S. Leaver

The authors of the paper above mentioned have sent the following statement to correct erroneous com-ments and implications in their article with the re-quest that it be published: "In the above mentioned article the reference to Bendelari jigs is used in the generic sense as applying to the mechanically driven diaphragm jigs. No attempt was made to dif-ferentiate between the various machines as installed on different dredges. The Natomas Company use a jig of their own manufacture. The Yuba Company installed the Placer jig as manufactured by the Pan American Engineer-ing Company. "The Pan American Engineering Company also manufactures the Crangle jig under their own name. This is a diaphragm jig in which the pulsations are produced by intermittent flow of water through the patent Crangle Valve and it has no driving mechanism."

Jan 1, 1937

By I. D. Shah

A roasting procedure for synthetic copper sulfide which results in a reproducible transient melting at 450° and 470°C is described in this paper. Possible causes for this melting phenomenon are suggested, and also a procedure to avoid it during the roasting. This melting is believed to be responsible for the occasional occurrence of stickiness in the commercial roasting of copper sulfide concentrate. Many workers have studied the roasting of copper sulfides, but only Lewis et al. have reported that fusion and, in some cases, fusion with "boiling" occurred in the neighborhood of 450°C. The DTA and TGA apparatus, the materials used in this investigation, and the discussion of chemical reactions occurring during the CuS roasting are described in an earlier publication. CuS, in the form of either pellet or powder, was heated at 280° to 330°C in a flowing air or 30% 02: 70% N2 atmosphere. Heating was continued at this temperature until no further weight increase due to sulfates formation was observed on the TGA thermogram. This is believed to be due to the inability of SO, to diffuse out from the sulfate layer on the surface of the pellet, thus halting the sulfate-forming reactions. Hereafter, the sample oxidized in this manner will be referred to as "partially oxidized CuS." X-ray analysis and microscopic examination of a polished cross section of a partially oxidized CuS pellet showed the following order of layers starting from the center of the pellet: Cu2S, Cu2O, CuO • CUSO4, and CuSO4. A bare thermocouple inserted in the center of a partially oxidized CuS pellet (2.5 g) did not indicate any noticeable change in the temperature of the pellet when melted at 450° to 470°C in an inert or oxidizing atmosphere. During the preparation of a partially oxidized pellet, the temperature of a CuS pellet increased by 50°C due to the exothermic formation of sulfates in air at 320°C. The endothermic peak associated with the melting phenomenon at 450°C is shown in [Fig. 1.] A CuS pellet (0.2 g) was oxidized in flowing air (150 cc per min) at 10°C per min up to 320°C and held there for about 2 hr. Then, N2 replaced air and heating was continued up to 500°C. The exothermic peak at 320°C represents an overall formation of sulfates, and an endothermic peak at 450° to 470°C represents the melting phenomenon. To observe the melting visually, a small (50 mg), partially oxidized CuS pellet was heated (in static air) in a hot-stage microscope. The melting phenomenon was observed on an auxiliary TV screen. The sample

Jan 1, 1976

By A. E. Back, S. F. Ravitz

When manganese ores are leached with sulphur dioxide, a large part of any zinc in the ore usually is extracted with the manganese.1 In the dithionate process,2,3 in which the manganese ore is leached with dilute sulphur dioxide gas and the manganese is subsequently precipitated by adding lime to the pregnant solution, the zinc is precipitated with the manganese. These facts suggested the possibility of applying the dithionate process to the recovery of zinc from oxidized ores or calcined sulphide concentrates. Accordingly, an investigation was begun, and preliminary tests were made on several ores and concentrates. The results were encouraging, but the need to devote available funds and personnel to other projects made it necessary to drop the work before the various factors involved in the process could be studied in detail. A brief mention of the process,4 however, elicited a number of inquiries, so that it appeared desirable to report the results of the preliminary work. The dithionate process as applied to zinc may be summarized briefly as follows: The ore or calcined concentrate is leached with dilute SO2 gas to dissolve the zinc, largely as zinc sulphate, in accordance with the following equation in which [ZnO] represents the "oxide zinc" in the ore or calcine: [ZnO] + SO2(g) + 1/2 O2 = Zn++SO4— [1] If, in the leaching step, the ore or calcine is suspended in calcium dithionate solution, the sulphate ion, as it is formed, is precipitated immediately as calcium sulphate, which is subsequently filtered off with the insoluble ore residue: Zn++SO4— + Ca++S2O6— = CaSO4(s) + Zn++S2O6- [2] An alternative procedure is to suspend the ore in water, filter off the insoluble ore residue after leaching, add calcium dithionate to the resulting zinc sulphate solution, and then filter off the calcium sulphate precipitate. In either event the zinc is obtained in the form of a pregnant solution of zinc dithionate, from which it can be precipitated as relatively pure zinc hydroxide by adding milk of lime: Zn++S2O6— + Ca(OH), = Zn(OH)2(s) + Ca++S2O6- [3] The zinc hydroxide precipitate is calcined to form the final zinc oxide product: Zn(OH)2 = ZnO(s) + H2O(g) [4] Theoretically the process is completely cyclic with respect to dithionate ion, the calcium dithionate required for reaction 2 being replaced by that formed in reaction 3. In practice, of course, small amounts of dithionate would be lost in the various filter cakes, depending on the economic washing limits. In the leaching step, however, some of the SO2 is oxidized to dithionale instead of sulphate, accord- ing to the following overall reaction: ZnO + 2SO2(g) + 1/2 O2(g) = Zn++S2O6-, [5] so that the dithionate losses are made up, at least in part. Ores and Concentrates Tested The analyses of the ores and concentrates tested are summarized in Table 1, which also includes the analyses of the calcines produced from the concentrates. The concentrates were roasted in a laboratory muffle furnace at the temperatures indicated until evolution of SO2 could no longer be detected; shallow clay dishes ordinarily were used; but calcines 2a and 2c were prepared in iron dishes, with the result that these two calcines were badly contaminated with iron scale. The "complex" concentrate, sample 1, consisted essentially of sphalerite grains containing inclusions of pyrite and other sulphide minerals too minute to be liberated by grinding within economic limits. In the " marmatitic" concentrate, sample 2, the iron was present largely as marmatite (that is, as iron sulphide in solid solution in zinc sulphide) rather than as pyrite inclusions. The zinc was present largely as the carbonate in the two "oxidized" ores, samples 3 and 4, and as the silicate in the "willemite" ore, sample 5. Leaching Tests The leaching cell was a 3-liter glass beaker placed in an electrically heated water bath manually controlled to maintain a temperature of 35°C in the leach slurry. The slurry was agitated by means of a flat paddle-type stirrer

Jan 1, 1950

By Å. Sterten, R. Tunold, J. Brun, K. Dalatun

The subjects of this study are two corrosion phenomena familiar to operators of aluminum plants employing Soderberg anodes of the vertical type, namely the sul-fide scale formation observed at the steel contact spikes and the corrosion wear observed at the lower edge of the gas collecting skirts. The causes of these corrosion phenomena are discussed on the basis of observations from practice and measurements on commercial cells as well as on laboratory experiments. It is concluded that the anode spike corrosion is mainly due to the attack by sulfur compounds contained in the anodically evolved gases and that this attack is greatly promoted by cracks and pores in the anode carbon embedding the lower spike ends. The corrosion of the gas collecting skirt is mainly attributed to anodic dissolution of the iron during occasional contacts with the melt. ANODE SPIKE CORROSION The visible result of this corrosion is the formation of a brittle loosely adherent scale on the lower ends of the mild steel spikes maintaining the electric contact to the carbon anode, Fig. 1. The thickness of the scale usually varies from 0.5 to 3 mm, but in some cases scale of about 10 mm thickness has been observed by the writers. During the spike pulling, usually repeated every 3 weeks, part of the scale often loosens and remains in the anode carbon, resulting in contamination of the aluminum with iron. This corrosion has been attributed to reaction of the iron with gaseous sulfur compounds originating from the anode paste materials.'-4 This paste usually consists of 70 pct precalcined petroleum coke with 1.2 to 1.4 pct S and 30 pct coal tar pitch containing 0.5 to 0.7 pct S. During operation the gasification of the anode materials takes place in two ways: First, gases are evolved by the gradual baking of the anode paste added on top of the anode. Due to the sealing effected by the upper semifluid zone of the paste, these gases, containing sulfur compounds from the pitch, are forced downward through pores and cracks in the rigid anode carbon. At the temperatures of the zone of the lower ends of the spikes, 700" to 900°C, the original components of the baking gas must undergo decomposition, similar to that taking place with the production of coal gas, resulting in a gas mixture consisting mainly of H2 and CH4 with small percentages of H2S as the main sulfur compound.5 Second, the anodic reaction results in a mixture of C02 and CO, the total quantity of which is about 15 times that of the baking gas. These anode gases must carry the entire sulfur content of the carbon gasified by the anodic reaction. Henry and Holliday6 have shown that COS is the main sulfurous component of the anode gases from Soderberg cells. Soderberg anodes usually exhibit a network of vertical cracks, frequently traversing from one spike to the next one and to the outside of the block and widening downward to the working face of the anode. In addition the anode carbon, particularly that below the spike ends, is rather porous, as indicated in Fig. 1. It therefore seems probable that the anode gases also may partly penetrate to the spike ends. As the anode is partly immersed in the melt, the static pressure of this must forward this penetration. I) SPIKE SCALE COMPOSITION A great number of freshly extracted spikes from industrial plants were inspected and samples of the scale collected for examination. Before analyzing the samples a thin iron oxide layer, formed by air oxidation of the outside of the red hot scale immediately after spike pulling, was removed. On the average the Fe/S ratio found by these analyses corresponded to I the composition FeSl. oo7. X-ray powder patterns showed no oxide and otherwise agreed with those found by Haraldsen7 for iron sulfide with 50.00 at. Pct

Jan 1, 1970

By F. R. Morral, K. T. Aust

The preferred orientation and earing characteristics of ZS aluminum were studied. An empirical correlation was obtained relating earing behavior and variation of mechanical properties for face-centered cubic metals. Strain-ratio mecsurements for ZS were found to be in good agreement with Hill's theory of plastic anisotropy. DIRECTIONALITY in sheet metals may be revealed by earing in a cupping test, or by the variation of mechanical properties of tensile specimens taken at different angles to the rolling direction. However, no simple correlation appears to exist between earing and variation in mechanical properties in metals and alloys studied.1-5 Aluminum of commercial purity (known as 2S, with a specified minimum of 99.0 pct Al) cast by the direct casting process was chosen in the present investigation since ears at 45° or 0°, 90° can readily be obtained. Preferred orientation is often the principal cause of uneven metal flow, such as earing, and variation in mechanical properties. Consequently, the preferred orientation characteristics of 2s aluminum were initially studied using both X-ray diffraction and micrographic techniques. The variation of mechanical properties with angle to the rolling direction was next determined in an effort to clarify the apparent lack of correlation of earing behavior with mechanical properties for face-centered cubic metals. Strain-ratio*5,8 measurements were finally conducted to determine first, if the maximum values of the strain ratio occurred at the earing positions, and second, if Hill's theory of plastic anisotropy7 is applicable to other face-centered cubic metals such as aluminum. Detailed treatments, working, and heat treating are not given in this paper since this study is not concerned with the methods to produce earing, but to correlate earing with other properties and characteristics of the material. The percentage of earing was measured using the formula specified by the x—Y Aluminum Association, -------- X 100 = pct earing, where x is the height of the ear and y is the height of the valley from the bottom of the cup. Preferred Orientation Characteristics of 2S X-ray Studies: Beck and coworkers8 have determined that the rolling texture of 2s after 95 pct cold reduction may be approximately described by four equivalent ideal orientations near (123) [121]. Also, the recrystallization or annealing texture of 2s aluminum consists of four components of the (123) [l21] type, retained from the rolling texture, and of a (100) [001] or cube texture component." Fig. 1 illustrates the typical appearance of deep drawn earing cups with the corresponding X-ray diffraction patterns. The X-ray patterns were taken with copper radiation using the structure-integrating method'" in which the specimen is scanned during the X-ray exposure. The rolling direction was vertical and the rolling plane was perpendicular to the horizontal X-ray beam. It was found that an increase in the retained rolling texture component resulted in a greater 45" ear height, while an increase in the recrystallization texture component (i.e., cube texture) caused greater 0°, 90° earing. X-ray diffraction studies of 2s cold rolled to final thickness followed by a final anneal have shown that the rolling texture, and also the retained component of the rolling texture after annealing, are increased with greater cold-rolling reduction. The preferred orientation characteristics of 2s sheets given an intermediate anneal between cold rolling followed by a final anneal were found to depend on the position of the intermediate anneal during cold rolling. For instance, three identical samples of 2s homogenized ingot, each 1 in. thick, were given the following cold-rolling and annealing treatments: 1—50 pct reduction, intermediate anneal, 96 pct reduction to 0.020 in. and a final anneal; 2—75 pct reduction, intermediate anneal, 92 pct reduction to 0.020 in. and a final anneal; and 3—90 pct reduction, intermediate anneal, 80 pct reduction to 0.020 in. and a final anneal. The intermediate and final annealing treatments consisted of a 70°F per hr heating rate, held 6 hr at 800°F and air cooled. Typical X-ray patterns after each of these treatments are shown in Fig. 2. It was evident that after treatment 1, the retained rolling texture component predominated; after treatment 2, the retained roll-

Jan 1, 1954

By J. P. Hirth, A. H. Clauer

A simple graphical method is presented for the determination of climb forces on dislocations exerted by uniaxial stresses. In conjunction with standard stereographic projections, the technique is applicable for any crystal system. Climb forces on glide dislocations in bcc crystals are determined as an application. In connection with a study1 of elevated temperature creep of molybdenum single crystals, a detailed analysis of the climb force on a dislocation was necessary. In general, such forces can be calculated for an arbitrary dislocation segment from the Peach-Koehler equation2 (or one of its variants3"6) where Fk are the components of the force per unit length of dislocation line, ?i of the unit sense vector tangent to the line, ojl of the stress tensor, bl of the dislocation Burgers vector, and is the Einstein permutation operator with components ?123 = ?231 = ?321 = —?132 = — ?321 = —?—213 = unity, and all other components zero. It is understood in the suffix notation of Eq. [I.] that repeated indices are to be summed over 1, 2, and 3. For the most general coordinate system, the matrix manipulations required to solve Eq. [I] are fairly complex. However, a simplification can be achieved by specifying a coordinate system based on the dislocation in question. The pertinent system involves cartesian coordinates xi, x;, xi, with attendant unit vectors e'i; ei being parallel to the Burgers vector b, and e'3; being parallel to the climb directions (b x ?), i.e., normal to the slip plane. In this coordnate system b = (bi, 0, 0) and ? = (?'1,?'2, 0). Hence, Eq. [I] yields for the climb force per unit length F3 = ?'2s'11b'1—?'1s'12b'1 [2] The first term on the right side of Eq. [2] is the climb force on the edge component of the dislocation, while the second is the (cross-slip) force normal to the glide plane acting on the screw component. Thus, the problem of determining the climb force reduces to that of resolving the stress components —il and s'12 in the xi coordinate system. Several authors7- l0 have shown, for the glide case, that such stress resolution can be performed most readily by graphical methods. Therefore, we present a similar method for the climb case. Of the suggested methods, we follow that of Hartley and Hirth,9 which can be performed directly with a standard stereographic projection for the crystal system in question. We treat the case of simple tension, or compression, this being the most common creep test in which climb is important. A set of cartesian coordinates (x1, x2, x3) is fixed with x3 parallel to the specimen axis. This set is connected to the x'i set by the factors defined in Table I. A perspective view of the coordinate systems relating them to the glide ellipse, is given in Fig. 1, while a stereographic projection of the coordinate systems is presented in Fig. 2. The transformation matrix connecting the xi and xi coordinates is x'i =aijXj [3] where an a11 a13 aij = a21 a22 a23 a31 a23 [4] s The corresponding transformation for the stresses is i [5] The only nonzero component of okl for the simple tension case is s33,. Hence, using Eqs. [4] and 151, we find that Eq. [2] becomes F3 = b'1s33 sin2 cos ?[cos ? sin a - sin ? cos a] [6] = m1b'1s033 sin a- rn2b's33 cos a where sin a = 5: and cos a = ?i, with a the angle measured from b to ?, i.e., a polar angle in the x1, x2 plane. The same equation with a sign change applies for compression. A graphical plot of the "Schmid" factor m1, relating to the climb force on the edge component, is presented in the stereographic projection of Fig. 3. In this projection the glide plane normal points toward the top of the page and the glide direction points normal to the plane of the page. To use the graph, one

Jan 1, 1970

By C. L. Warwick, A. B. Parsons

The papers and discussion appearing in this volume were presented at a symposium jointly sponsored by the American Society for Testing Materials and the American Institute of Mining and Metallurgical Engineers and held at Philadelphia, Pa., November 29, 30, and December 1, 1944. Late in January, 1944, Lieut. Col. C. H. Greenall, Officer-in-Charge of the Laboratory Division, Frankford Arsenal, Ordnance Department, U.S.A., suggested to the American Society for Testing Materials that it might be a propitious time to arrange for a symposium on stress-corrosion cracking at the Annual Meeting of the Society in June, and agreed to make available a number of papers relating to work being carried out at that Arsenal. The Society's Committee B-5 on Copper and Copper Alloys, Cast and Wrought expressed keen interest in the symposium, and other Society committees were contacted and agreed to cooperate. Since the subject is one of primary interest also to the American Institute of Mining and Metallurgical Engineers, it was arranged that the two organizations would jointly sponsor the symposium. A Symposium Committee was formed under the joint leadership, for A.S.T.M., of Carter S. Cole, then Chief of the Metals Branch, Conservation Division, War Production Board and now Engineer on the Headquarters Staff of the Society; and, for A.I.M.E., of E. A. Anderson, Chief of Metals Section, Research Division, New Jersey Zinc Co. (of Pa.) and chairman of the Program Committee of the Institute of Metals Division. The Symposium Committee was constituted as follows:

Jan 1, 1945

By Terkel Rosenqvist

THE desulphurization of molten iron and steel is a very complicated process. One way to arrive at a better understanding of this process is to break it down into several simpler chemical processes that can be studied individually in the laboratory. For a study of the different factors that influence the equilibrium distribution of sulphur between liquid metals and slags, several simpler equilibria may be investigated. One very important subject is the determination of the escaping tendency of sulphur in the liquid metal and its dependency on temperature and composition of the melt. Several papers in this field have recently been published.', ' Another subject is the study of the sulphur capacity of the slag. A molten slag is indeed complex, and even if sulphur distribution data for a large variety of molten slags may give empirical data about their desulphurizing power, the importance of the individual components is still not quite clear. It is accepted generally that lime is the most important desulphurizing component in the slag. The present investigation has as its purpose to study the desulphurizing power of lime in its standard state, and to provide a basis for thermodynamic calculations of the desulphurizing power of various lime-containing slags. The standard state of lime at steelmaking temperatures is solid calcium oxide, CaO. It can react with sulphur to form solid calcium sulphide, CaS. The relative stability of calcium oxide and calcium sulphide is expressed by the free energy of the reaction: 2Ca0 (s) + S1 (g) = 2CaS (s) + O2 (g) The existing free energy data for this reaction, listed by Kelley5 nd Osborn,' are uncertain to about 10 kcal and are of limited value for a calculation of equilibrium constants. Under the conditions prevailing in a melting furnace, the sulphur pressure may be expressed conveniently by the ratio H,S/H2 and the oxygen pressure by the ratio H,O/H, (or CO,/CO). The desulphurizing power of calcium oxide may, therefore, be studied by the reaction CaO + HIS = CaS + H2O. A study of this reaction may be complicated by certain side reactions: Water vapor and hydrogen sulphide may react. to form sulphur dioxide, and calcium sulphide may be oxidized to calcium sulphate. A thermodynamic calculation shows that these side reactions will be suppressed to insignificance if the equilibrium is studied in the presence of an excess of hydrogen. The apparatus used is shown in Fig. 1. About 10 g calcium oxide and 20 g calcium sulphide (laboratory qualities) were intimately mixed, and some water was added to make a thick paste. The paste was put into a thimble of zirconium silicate, which was placed within the constant temperature zone of a furnace, and capillary refractory tubes were attached in both ends. After the mixture had been heated in dry hydrogen at 1000°C for several hours all Ca(OH), and CaCO, had decomposed and CaSO, was reduced, so only CaO and CaS remained in the thimble forming a porous plug. The mixture was examined by X-ray diffraction after the initial reduction in dry hydrogen as well as after the subsequent experimental runs up to 1425 °C. It was shown that crystalline calcium oxide and calcium sulphide were always present together in about equal amounts. The unit cell edges were found to be 4.80A for CaO and 5.68A for CaS in good agreement with existing literature values." This shows that the mutual solid solubility is very small, and that the compounds are present in their standard states. Purified hydrogen was passed through water sat-urators kept at constant temperature in a thermostat bath. The amount of water vapor saturation was checked by means of a dew point method, not shown on Fig. 1. The gas mixture was passed through the capillary inlet into the furnace, where it was sifted through the porous plug of calcium oxide and calcium sulphide. The hydrogen sulphide present in the outgoing gas was absorbed in a zinc acetate solution and the hydrogen was collected over water. When one liter of hydrogen had been collected, the amount of hydrogen sulphide was determined by iodometric titration. As one molecule of H,O is used for the formation of each molecule of H,S, the equilibrium ratio H,S/H,O would be , where (H,O) is the molar concentration in the ingoing gas, and (H,S) the molar concentration in the outgoing gas. In the present work (H,S) was always very small compared to (H20). In order for the observed H,S/H20 ratio to represent the true equilibrium ratio the gas flow has to be: 1—Sufficiently slow to give a complete establishment of equilibrium, and 2—sufficiently fast to counteract thermal diffusion. Incomplete reaction would give a value decreasing with increasing flow rate, and thermal diffusion would give a value increasing with decreasing flow rate. When inlet and outlet tubes of about 2 sq mm cross-section were used, the observed gas ratio was independent of the flow rate between 15 and 125 cc per min, Fig. 2. In this range, therefore, the observed gas ratio represents true equilibrium.* For the rest of the in-

Jan 1, 1952

By Lyndon L. Foley

The proper spacing of oil wells is a problem of vital importance to the oil industry. Conservation demands a maximum recovery, while economic considerations attach primary importance to profitable extraction of oil. Formulas have been evolved,1-4 which attempt to solve the spacing problem mathematically, using reservoir characteristics as factors. Comparisons have been made of fields producing from supposedly similar sands, which have been drilled to varying densitiesl5,6 and also of different leases in the same field. 5-8 Intermediate wells have been drilled in old, depleted fields to test the efficiency of drainage by the original operation.9-11 Intermediate wells have been drilled also to obtain cores of the depleted sand for analysis, from which to learn the degree of efficiency of extraction. The problem is being attacked by many students, and along several different lines. Unfortunately, the fields now depleted were produced generally without restriction, and conclusions drawn from them may not apply to fields operated under restricted production. Reservoir conditions in a restricted field, especially reservoir pressures and pressure gradients within the reservoir, are quite different from those in a field produced to capacity. Many people have held the idea that the closer the spacing, the greater the recovery. This general statement is not always true, as other factors than well spacing enter into the recovery problem. Cutler,5 who has been quoted frequently by advocates of close spacing, limited his conclusions to gas-drive fields, and to the particular fields and spac-ings that he investigated. Hill and Sutton7 analyzed four tracts in the Powell field; the tract most closely drilled had the third largest recovery per acre-foot, and the tract with the widest spacing had the second largest recovery per acre-foot. A search of the references to well spacing in the publications of the U. S. Bureau of Mines5,13-19 discloses repeated condemnations of the wastefulness of the practice of town-lot drilling.

Jan 1, 1938

By J. Gross

G. H. Clevenger, Chairman of the Milling Methods Committee of the Institute, made the following introductory remarks: "Several years ago, a number of us felt that the time was ripe for a fundamental investigation of crushing and grinding. At the outset, it appeared that the first thing we needed was a yardstick whereby grinding efficiency could be measured. "The U. S. Bureau of Mines has been most generous in cooperating with the Milling Methods Committee in getting this work well under way. John Gross, who served the committee for a year as secretary, was placed in charge of the investigation and has been ably assisted by S. R. Zimmerley. "Messrs. Gross and Zimmerley will now present the Progress Report covering their work to date." PROGRESS REPORT J. Gross, Salt Lake City, Utah. -This has not been a simple problem, and when I tell you that we can give you a verbal report of what we have done during three years in five or ten minutes you will realize that we have necessarily done a great deal of work that does not appear in the final results. We have been in a virtually new field. Therefore, we have had to feel our way at each step so as to be fairly sure we were correct. The first thing that was realized in this problem was that unless we had some method of measuring what happened when we crushed an ore, we might as well quit and not go further. We realized that if we tried to measure the results of crushing or grinding by the ordinary sizing test we would not get anywhere. We could sieve down to say 200 mesh fairly accurately; we could even go one sieve size further with a little less accuracy, and then we were absolutely lost, unless we went into microscopic measurement. Even this latter method would not have been satisfactory. The question was: what were we going to measure? If we were going to measure size, we might measure one particular dimension of a particle and be off 200 per cent. or more. If we tried to measure surface, we would have something tangible; the surface of a particle, no matter how irregular, is something definite, far more definite than any linear dimension would be. Moreover, if we could measure surface, we would immediately have a method of checking the Rittinger law as to whether it was or was not correct. So the first thing that we sought was a method whereby we could measure surface on crushed particles or, we will say, on particles; it would not make any difference whether crushed or not. It then became a question of what kind of particles we were going to use in obtaining a method for surface measurement. We realized that it would be rather a com-

Jan 1, 1928

By F. Louis Grammer

The difference between blast-furnace gas and ordinary producer-gas is chiefly that blast-furnace ga,s is higher in CO2 and lower in hydrocprbons and hydrogen, as is shown in Table I. Table I.— Volume-Composition of Producer-Gas and Blast-Furnace Gas. Producer-Gas. Blast-Furnace Gas. Per Cent. Per Cent. (Volume.) (Volume.) CO......22 to 33 25.5 CO2........ 6 to 3 13.0 H...13 to 7 1.5 CH4........ 3 to 1.5 0.0 N..54 to 60 60.0 If natural gas, water-gas, or by-product coke-oven gas be available, blast-furnace gas can be made to resemble very closely in calorific value the average producer-gas by merely mixing in one of these products, which are richer in hydrocarbons and hydrogen. Blast-furnace gas, however, possesses a virtue which is sometimes absent in producer-gas, especially with coals very high in sulphur—namely, that blast-furnace gas is relatively free from sulphur. In some regions the sulphur in coal is an acute problem, and it is highly desirable to find uses for the mine-products which are specially high in sulphur-content. An advantage would be gained if the coal were coked before charging into the gas-producers, but this adds to the expense. Sulphur may be partly eliminated from coal by washing it, the success depending upon the form in which the sulphur occurs, and upon the efficiency of the washer, but it cannot always be removed to the extent desired at most American plants. It is not practicable to remove all the sulphur from producer-gas by washing it, though Mr. Reinhardt, of Germany,

Jan 1, 1909

By H. C. Chao, L. H. Van Vlack

Small MnS inclusions with known crystallographic orientations were placed inside powder compacts of low-carbon steel. After the metal was axially campressed with negligible end friction, the deformstions for the metal and the inclusions were compared. The MnS inclusions deformed more when the [100] direction was aligned with the compression axis than when the [111] direction was parallel to this axis. The deformations of the inclusions in the two principal radial directions were equal for each of the above orientations. Inclusions with [110] compression alignments did not deform with radial symmetry. The relative deformation of the inclusion and metal was closely dependent upon the relatiue hardness of the two phases. The relative deformation of the two phases was not sensitive to the rate of deformation. RECENT studies by the authors1.&apos; suggested that the plastic deformation of MnS in steel would probably be highly sensitive to the orientation of the inclusions and to the temperature of the metal. This paper reports an investigation of these factors upon MnS behavior within steel. Manganese sulfide (MnS) possesses an NaCl-type structure and typically has extensive (l10) {110} slip as a separate (noninclusion) crystal.&apos; A secondary slip system, ( l 10) { l l l}, has also been observed where the major slip system is restricted. In general, MnS inclusions must be rated as a highly deformable second phase.3 The amount of sulfide deformation varies, however, with several composition and processing factors. Some of these have been only partially assigned. For example, it is known that minor amounts (<0.01 pct) of silicon within free-machining steels will increase the amount of MnS deformation,4 but the mechanism of the added deformation can only be surmised at the present. Manganese sulfide and steel have sufficiently comparable deformation characteristics so that slip which is started in steel may be continued through the sulfide inclusions and back into the steel if the crystal orientations are favorable.5 A more detailed discussion of previous work on the plastic deformation of NaC1-type crystals and on the plastic deformation of inclusions within a metal is given in Chao&apos;s work.6 EXPERIMENTAL PROCEDURE The manganese sulfide which was used in this study was prepared by previously described methods.&apos; Single crystals of MnS, both as cleavage cubes and as spheres, were oriented within steel powder compacts so that the desired crystal directions were parallel to the direction of axial compression. A four-stage hydrostatic compaction procedure was used and involved the following steps. In the first stage part of the powder was placed in a metal die 1 in. in diameter with a thick (1 in. OD, 5/8 in. ID) rubber liner which had one end plugged. The steel powder was hand-rammed, making it as dense as possible before placing a carefully sized MnS crystal (either as a sphere or as a cube) near the center. The crystal was oriented with the chosen direction vertical; viz., [001], [011], or [111], with the aid of a X10 microscope. A pair of tungsten wire threads 0.020 in. in diameter was inserted along the side of this (&apos;core compact" to locate the desired plane after the compression tests. After the crystal was positioned in the center of the die, more powder was added and carefully rammed by hand. The die was then capped with a rubber plug of the same hardness and thickness as that of the liner. The whole assembly as shown in Fig. 1 was compacted by a ram load of 54,000 lb (about 70,000 psi). In the second stage a smaller, 3/4-in, rubber-lined die was used to give a stress of approximately 120,000 psi. The above process was repeated with the initial compact serving as a core for a larger compact. The final product after sintering was a cylinder 1 cm long and 1 cm in diameter, having a density of 7.54 g per cu cm. This was close to the theoretical density since the metal contained a non-metallic phase. There was no evidence of MnS deformation during the hydrostatic compaction or subsequent sintering. Elevated-temperature hardness data were obtained by procedures previously described.2 Compression tests for inclusion deformation utilized the cylinders which were described above. The critical problem in these tests was the lubri-

Jan 1, 1965

By W. C. Bleimeister

The rock salt deposits of the U. S. occur in five major basins: 1) the Eastern Basin (which includes parts of Michigan, Ohio, Pennsylvania, New York, West Virginia, and Ontario) ; 2) the Gulf Coast Basin (the northern rim of which includes parts of Alabama, Mississippi, Louisiana, Arkansas, and Texas, while its southern rim includes parts of the Isthmus of Tehuampetic and the Yucatan Peninsula of Mexico), 3) the Southwestern Salt Basin (which includes parts of Kansas, Colorado, Oklahoma, Texas, and New Mexico), 4) the Williston Basin (which under- lies parts of North Dakota, South Dakota, Montana, and Saskatchewan), and 5) the Green River Basin (whose periphery underlies parts of Wyoming, Utah, and Colorado). The greatest salt production in the U. S. is obtained from the Eastern Basin,-particularly from the areas of Michigan, Ohio, and Ontario. In 1958 the salt production from these areas amounted to 8.7 million tons or 35.3 pct of the combined total produced in the U. S. and Canada. Of the 24,286,000 tons produced by Canada and the United States, 6,443,000 tons were in the form of rock salt from dry mining, while 17,843,000 tons were produced from wells in the form of artificial brine, the result of dissolving the rock salt formation by fresh water. Three important factors lie behind the location and development of this salt mining area:

Jan 5, 1961

By S. C. Sircar, B. K. Dhindaw

KINETICS and mechanism of sulfation reactions have been studied extensively. Qualitative and quantitative data are reported for the kinetics and mechanism of sulfation reaction of copper oxide pellets;&apos;&apos;2 yet it remains a controversial Subject. The mechanism and kinetics of the sulfation of zinc oxide powder (B.D.H. analar quality) were studied by gravimetric technique. A spring of high-carbon steel wire with accuracy well within 0.2 mg was used for measuring weight change. The sample was taken in a Pyrex glass bucket, hanging freely inside a Pyrex glass tube reactor, with a gold wire attached to the spring, the reactor being surrounded by a movable furnace (double wound type) with a temperature control within *5"C. Flow of oxygen, 60 cu cm per min, and sulfur dioxide, 45 cu cm per min, were measured through calibrated flowmeters and the mixture dried over PZ05 before entering the reactor. The experimental results are presented in Figs. 1 and 2. The plot of weight change of zinc oxide, giving the amount of zinc sulfate formed, was found to be linear over the temperature range of 500" to 700° C and up to a conversion of approximately 60 pct, as beyond it they divert from linearity.&apos; The rate of sulfation was found to be affected by total flow rate of gases at a fixed ratio of SO2/O2, up to 60 cu cm per min, and was found to be linearly dependent on quantity of total powdered sample taken. The plot of rate of sulfation of zinc oxide, 70 mg, with varying amounts of vanadium pentoxide, at a constant ratio of gases as mentioned above, and at a constant temperature of 625"C, was also found to be linear as seen in Fig. 2. Extrapolation to zero V2O5 addition gives the same rate of sulfation, as in case of pure zinc oxide, 70 mg. Vz05 was prepared by heating ammonium vanadate at 600°C for 15 min in a muffle furnace. The available literature points out that the theory of sulfation that has received the maximum support postulates the following mechanism for sulfation of ZnS: 2ZnS + 30, = 2Zn0 + 2SOZ; the ZnO formed acts as a catalyst for the reaction, 2 S02 + O2 = 2S03. This SO3 reacts with ZnO to give zinc sulfate. It is suggested here that the above widely supported mechanism is operative in the sulfation reaction and that the rate-controlling step in the sulfation reaction of ZnO is the formation of SO3 from SO, and 0,. The authors are carrying on further work in this line, and the details will be published later. The authors wish to acknowledge the facilities provided and the help extended by Prof. P. R. Dhar, Head, Department of Metallurgical Engineering, and the permission granted to publish this note by the Director, I. I. T., Kharagpur.

Jan 1, 1969

By AIME AIME

CONVENTION plans for the A.I.M.E. Regional Meeting to be held on the Minnesota Iron Range Aug. 12 to 1.5 are being completed to give the visiting member?s from all parts of the country a wide variety of technical and recreational features. High spots of the social events are the official dinner, dance, luncheon and golf: and of course, special provision is being, made for the entertainment of visiting ladies. The technical sessions will fit in with scheduled trips to view all phases of mining, beneficiation and transportation of iron ore. Men from the West and Southwest will find much of interest in the detail of the mining operations. A trip to the steel plant of the

Jan 1, 1941

By C. L. Mantell

When considered from the viewpoint of world annual output, tin is one of the rarest metals. Its annual production is exceeded by that of iron, copper, lead, zinc. aluminum, magnesium; probably by that of the ferroalloys of manganese, chromium, and silicon, and possibly by the world production of sodium; while it exceeds only the output of such common metals as nickel, antimony, mercury, tungsten, molybdenum, silver, gold and the precious metals. Cassiterite, or tinstone, until the second World War was the only mineral that was an important source of tin. This mineral is frequently called "tin ore." The use of the term should be restricted to the ore containing the mineral, and not, in addition, to the concentrates of the mineral obtained from the ore or from stanniferous alluvial deposits. There is a general misuse of the term "tin ore" in the Malay Peninsula and adjacent localities, particularly where most of the mineral is obtained from secondary stanniferous deposits. The term "cassit-erite" is exclusively used in mineralogical, geological, and other scientific writings; it is unfortunate that it is not mole frequently used when referring to the occurrence of the mineral on an economic scale. Ores "Tinstone" js a very convenient old English term deserving a more frequent use. Tinstone is a dioxide of tin, or stannic oxide. When chemically Pure, as in the very rare transparent variety, it has a metallic content of 78.6 per cent tin. Frequently, however, the crystals and grains contain appreciable amounts of impurities, chiefly iron and tantalum. The impurities here referred to are those actually in the mineral itself. Tinstone usually has a deep brown or black color with an adamantine luster. Several other colored varieties are known, among them red ruby tin, yellow rosin tin, and yellow wax tin. the names jn each case being descriptive of the mineral&apos;s appearance. Ainalite is a variety of cassiterite con-taining almost g per cent of tantalum pentoxide. Sparable tin, tooth tin, and needle tin, as a result of their acute dite-tragonal pyramidal crystalline form, re-ceive their names from their crystallo-graphic appearance. Wood tin is a compact variety of cassiterite composed of radiating fibers resembling dry wood. Toad&apos;s-eye tin is a similar variety on a smaller scale in which the fibers appear to resemble the eye of a toad. Stream tin is water-worn tin-stone. Float tin is sometimes employed to describe the cassiterite occurring in soil derived from the weathered surface of a mineralized area. Stannite is an ore of tin of lesser impor-tance than cassiterite. It is a sulphide of tin, copper, and iron, sometimes known as tin pyrites or bell-metal ore. Its chemical com-position is sometimes expressed by the for-mula Cu,S.FeS.SnSz with zinc usually present in varying quantities. The tin con-tent varies from 22 to 27 per cent and there is about 29 per cent of copper, 13 per cent of iron, and 30 Per cent of sulphur.

Jan 1, 1944

By V. Kondic, W. M. Rumball

HUNT and Jackson&apos; have recently suggested that binary eutectics might be classified according to the entropy of fusion of the component phases. Such a classification was in fact proposed2 some time ago by the authors and the purpose of this communication is to present some supporting evidence. As pointed out by Hunt and Jackson&apos; typical metals of low entropy of fusion, ?S, tend to form normal eutectics (using Scheil&apos;s terminology3). Where one phase is metallic (low ?S) and the other nonmetallic (high AS) then anomalous structures are formed. These anomalous structures may either be regular (e.g., Bi-Sn)or irregular (e.g., A1-Si). Finally Hunt and Jackson suggest that where both phases are of high ?S regular faceted structures are formed. In the proposed classification2 the ratio of the entropies of fusion is taken as the quantitative parameter, i.e., ?Sa/?SB = E, where e > 1.0. An extensive literature survey has been carried out2,4 for reliable photomicrographs of eutectic structures. The structures, numbering 151, have been classified into two classes, viz., normal (i.e., lamellar and rodlike) and anomalous. The latter group has been arbitrarily subdivided into regular and irregular anomalous eutectics. Details and sources of the structures are given elsewhere.2,4 In the absence of better thermodynamic data the value of E has been determined for each of the systems making the following assumptions and approximations: l) The entropy of fusion of solid solutions may be calculated additively;5 ii) Inter metallic compounds are treated similarly but with the addition of the following term for fully ordered structures.5 -4.573(Ni lOg10 Nl + N2 log10 N2) where Nl and N2 are the atomic fractions. It might be argued that the entropies of fusion are meaningless as the eutectic phases form at some temperature (T,°K) below their true melting point (Ta or Tb). As AS, = Lo/To if La the latent heat of fusion of a, is assumed independent of temperature, the actual entropy change AS& is given by La/Te. The ratio of entropies of fusion then becomes Fig. 1 shows E&apos; plotted against cumulative totals for normal and anomalous systems from which it is seen that normal eutectics predominate at low values of E&apos;; e.g., some 90 pct of normal systems lie below E&apos; = 1.5 with 65 pct of anomalous systems above this value. Fig. 1 also shows that the regular anomalous structures occur mostly at low values of c&apos; possibly suggesting growth more akin to normal than anomalous growth. The choice of c as a classification parameter was based on nucleation concepts, it being observed that metals widely separated in Sundquist and Mondolfo&apos;s8 nucleation series were anomalous and those close together normal. The basis of these classification views is really outside the intended scope of this communication but it is hoped to explain them fully in the near future.7 Having arrived at this entropy of fusion classification by nucleation considerations it is interesting to note that Hunt and Jackson&apos;s ideas are based on growth considerations. Elsewhere one of the authors4 has shown that an even better classification of the metallographic structures reviewed is obtained from purely conventional heterogeneous nucleation concepts. It has been found that systems having a value of the parameter +b less than 0.28 are almost certainly normal and above this anomalous: V, is the gram atomic volume. This parameter is nevertheless closely related to c and c&apos;, these being an approximation of p. The parameter is also related to the other classification3,8,9 which have met with any success in the past. All these classifications are to be reviewed shortly by the authorsU7 A significant thing about the classifications based on

Jan 1, 1968

By J. B. Cohen

In an investigation of the microstructure of shock-loaded metal specimens1 C. S. Smith found that there was little distortion of grain boundaries or change in external dimensions, although the hardness corresponded to that produced by severe rolling. Furthermore, the hardness of individual grains was independent of the deformation structures present; grains with no markings were as hard as those heavily populated with markings. These markings which resemble twins do not appear in copper until high shock-loads of the order of 200 kb are applied. It appears that due to the high velocity of a shock wave and its sharp front, heavy deformation takes place on a submicroscopic scale. In this work an attempt was made to examine this deformation, particularly at a load below which macroecopic markings appear. The broadening of 200 and 400 diffraction peaks was examined from the face of a poly crystalline sample normal to the direction of the shock wave. These peaks were chosen because the elastic constants in copper are such that residual strains would be greatest in the <100> directions; hence, the strain-broadening contribution to these peaks would be large in comparison to other sets. Also the broadening due to stacking faults and microtwins would be greatest with this set.2 A complete Fourier analysis of these peaks was made, as described by warren,&apos; using the Stokes correction with peaks from an annealed sample.3 With the aid of the U. S. Army Ordnance Corps EDVAC computer and a curve matching program, sixty sine and sixty cosine coefficients were obtained for each peak. The last of these was about 1.5 pct or less of the largest coefficient, indicating good representation. Transverse sections from a 1.5 in. rod of as-drawn electrolytic tough-pitch copper were reduced in cross-sectional area about 50 pct by cross rolling. Disks 1.5 in. in diam were recrystallized at 275 °C for 30 min and then at 300°C for the same time. The faces were polished and deeply etched in nitric acid to remove 25 mils. The grain size was 0.02 mm. Laue back-reflection patterns indicated the sample was well recrystallized. The two diffraction peaks from this sample were recorded on a Norelco diffractometer using filtered copper radiation. Peak intensities were kept less than 600 cps to assure linearity of the counter and circuitry. This sample was then placed between two rods of the same material and diameter. The top rod was 2.25 in. long and the bottom rod 1 in. long. (In

Jan 1, 1961