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By David W. James, Howard Jones, George M. Leak

Conditions are described for producing, by primary recrystallization, a matrix suitable for the growth of large grains near (110)[001] orientation in silicon iron strip by secondary recrystallizaliun in a steep temperature gradient. The orientation distribution of these large grains is expressed in terms of rotational deviations about the cross-rolling direction, the rolling direction, and the normal to the sheet, the deviational spread increasing in that order. With the aid of cowplenientary published data on the orientation dependence of growth rate, it is shown that this observation is consistent with the oriented-growth theory of recrystallization lextures. It is conclutled that growth-rate and orientation-distribution data obtained in a steep thermal gradient should be used with caution to account for isothermally Produced recrystallization textures. SEVERAL authors have reported methods of growing large grains by re crystallization of a small-grained matrix in silicon iron 1- B and pure a cr The present study was a preliminary in the growth of single crystals and bicrystals for surface relaxation," grain boundary mobility, and grain boundary diffusion studies. The method was to control the growth of a seed crystal into a suitable primary re crystallized matrix by feeding through a steep temperature gradient. The driving energy for growth derived from the grain boundary energy released as the seed crystals grew into the matrix. Thus, stability of the matrix against normal grain growth was considered to be essential for success. It was known that the manganese sulfide dispersion present in commercial silicon iron performs this function during secondary recrystallization to the (110)[001.] texture.12 Hence commercial, rather than high-purity, material was used throughout. The paper describes the growth conditions for grains large enough to be used as seed crystals for further growth into single crystals. The orientation distribution of the seed crystals is analyzed and its significance for the theory of recrystallization textures is discussed. EXPERIMENTAL PROCEDURE Strip material was supplied by the Steel Co. of Wales, Ltd. The chemical analysis in weight percent was Si, 2.90; C, 0.015; Mn, 0.059; P, 0.011; S, 0.027; Ni, 0.032; 0, 0.009; Fe, balance. A gradient furnace of similar design to one described previously4 was loaned from B.I.S.R.A. It consisted essentially of a vertical water-cooled copper slot projecting downwards into the hot zone of a molybdenum furnace. Hydrogen was passed through the furnace to protect both heating element and specimen from oxidation. Strip specimens up to 8 cm wide and 0.2 cm thick were sealed into the furnace at the mouth of the copper slot. A coating of light oil on the strip surface maintained the seal during translation of a specimen. The maximum temperature gradient in the region just below the copper slot was 500°C per cm over 1 cm, with the hottest point controlled at 1175°C. Several large grains would usually grow by secondary recrystallization from the primary matrix when a specimen was immersed in the hot zone for about 30 min. A back-reflection X-ray camera was constructed to facilitate rapid and accurate orientation determinations of the large grains produced. It was possible to reproduce a standard geometry, with regard to strip and camera, without the tedium of careful alignment on each occasion. Specimens, typically 4 cm wide and 75 cm long, were cut with the longitudinal axis parallel to the rolling direction of the original strip. The surfaces were cleaned by immersion alternately in a hot aqueous solution containing 2 pct hydrofluoric acid plus 10 pct sulfuric acid and in cold 10 pct nitric acid. The nitric acid etch was just sufficient to reveal the grain structure. Rolling and annealing treatments to prepare the matrix (discussed below) were followed by growth of seed crystals in the gradient furnace. The matrix was transformed to a single crystal by growth of a selected seed crystal connected to the matrix by a thin neck. 4,5 Growth was promoted by controlled feeding into the gradient furnace. Several single crystals of controlled orientation were grown successfully from seed crystals by twisting the interconnecting neck in a reorien-tation jig.4 EXPERIMENTAL RESULTS AND DISCUSSION Growth Conditions. A suitable matrix for growth of large grains was prepared starting from primary re-crystallized strip 1.9 mm thick. This was cold-rolled in two stages each being followed by a recrystallization anneal at 800°C for a few minutes. Such treatment gave the required growth matrix only if the two cold-reduction stages were each performed in several passes and in the following ranges: the first, 30 to 70 pct; the second, 10 to 50 pct. Immersion in the temperature gradient otherwise resulted in an equiaxed

Jan 1, 1967

By R. J. Wysor

MoRe than two years ago, in making efficiency tests on our hot-blast stoves, I was surprised to discover a marked difference in temperature as indicated by a pyrometer inserted near a stove on blast, and the regular recording instrument situated near the bustle pipe. A special hole was then provided at the junction of the hot-blast main and the bustle pipe, and portable pyrometers with bare fire ends were used, both at this point and in the stove necks. I had occasion to make a number of such tests on various stoves on the same, furnace, the difference in temperature as shown by the two pyrometers varying chiefly with the temperature of the blast, and the distance of the stove from the furnace. However, this difference was always great, often averaging above 100" F., for the entire period of a stove on blast. Later, experiments were made with similar results on another group of stoves. The pyrometers used were properly standardized and checked against each other before and after each test. However, there was still a slight doubt as to the accuracy of the results obtained, as subsequently explained. But I have obtained corroboration of the magnitude of heat losses from blast-furnace hot-blast mains from two independent sources. I have learned from H. A. Brassert, Superintendent of Blast Furnaces, South Works, Illinois Steel Co., that he has recently done some experimental work similar to my own, and found temperature drops of the same order of magnitude. Also, C.,R. Lyle, of the Armstrong Cork Co., has made some interesting theoretical calculations, based on the thermal conductivity of fire brick, for the particular conditions of several of our tests; and his theoretical interpolations, all facts being considered, check our actual observations fairly well. I understand that Mr. Lyle expects to publish the results of his investigations. Method of Conducting the Tests The location of the pyrometer fire ends is shown in Fig. I, a plan of Bethlehem Steel Co.'s "G" blast furnace, hot-blast stoves, and main. Bare fire ends were used at both positions "A" and "B," being inserted

Jan 1, 1916

By Robert D. Pike

In November 1939, on behalf of the Harbison-Walker Refractories Co., the author undertook the study of the problem of utilizing the dolomite of northwestern Ohio for the manufacture of calcined magnesia suitable for use in refractories. Grateful acknowledgment is made of the unfailing assistance and encouragement throughout this work from the executive and technical staff of Harbison-Walker Refractories Company. The dolomite of the Niagara formation that is exposed or near the surface of a wide area in northwestern Ohio1 constitutes a vast supply of almost pure dolomite. It offers a virtually unlimited supply of cheap raw material of uniform dependable analysis for the manufacture of magnesia. A typical analysis of dolomite from the vicinity of Luckey, Ohio, follows: Stone Completely Calcined CaO(Sro)a............. 30.35 57.806 Mgo................... 21.55 41.050 SiO2. ................... 0.30 0.572 R2O3................... 0.30 0.572 Ignition loss............. 47.50 a SrO about 0.06 pct. Brines containing magnesium chloride or sea water can be reacted with calcined dolomite to produce magnesia.2'3 In such reactions the MgCl2 of the brine supplies some of the magnesia and the amount of dolomite required is thereby reduced by an equivalent amount, but brines of commercial importance containing magnesium chloride are some distance away from the purest grades of the dolomite of northwestern Ohio. It was decided therefore to concentrate upon the development of a suitable process for separating the lime and magnesia of dolomite, depending upon the dolomite itself as the sole source of magnesia. This would make it possible to place a plant at the dolomite quarry, thus eliminating the cost of transporting the stone and also rendering the calcined magnesia available at a most convenient production point, considering the cost of delivery throughout the steel industry. The selection of a process depending solely upon the dolomite as a source of magnesia also necessarily entails production of the lime content of the dolomite as a by-product. It was recognized that if such a lime by-product could be produced in a suitable form, a plant location in northwestern Ohio would also be advantageous for marketing the lime. The problem, therefore, was the discovery and development of a low-cost process for separating the magnesia and lime of the dolomite of northwestern Ohio and producing each in a form relatively free from the other, and meeting the respective requirements of established markets. A satisfactory solution of this problem has been found, and it is the

Jan 1, 1948

By Robert D. Pike

In November 1939, on behalf of the Harbison-Walker Refractories Co., the author undertook the study of the problem of utilizing the dolomite of northwestern Ohio for the manufacture of calcined magnesia suitable for use in refractories. Grateful acknowledgment is made of the unfailing assistance and encouragement throughout this work from the executive and technical staff of Harbison-Walker Refractories Company. The dolomite of the Niagara formation that is exposed or near the surface of a wide area in northwestern Ohio1 constitutes a vast supply of almost pure dolomite. It offers a virtually unlimited supply of cheap raw material of uniform dependable analysis for the manufacture of magnesia. A typical analysis of dolomite from the vicinity of Luckey, Ohio, follows: Stone Completely Calcined CaO(Sro)a............. 30.35 57.806 Mgo................... 21.55 41.050 SiO2. ................... 0.30 0.572 R2O3................... 0.30 0.572 Ignition loss............. 47.50 a SrO about 0.06 pct. Brines containing magnesium chloride or sea water can be reacted with calcined dolomite to produce magnesia.2'3 In such reactions the MgCl2 of the brine supplies some of the magnesia and the amount of dolomite required is thereby reduced by an equivalent amount, but brines of commercial importance containing magnesium chloride are some distance away from the purest grades of the dolomite of northwestern Ohio. It was decided therefore to concentrate upon the development of a suitable process for separating the lime and magnesia of dolomite, depending upon the dolomite itself as the sole source of magnesia. This would make it possible to place a plant at the dolomite quarry, thus eliminating the cost of transporting the stone and also rendering the calcined magnesia available at a most convenient production point, considering the cost of delivery throughout the steel industry. The selection of a process depending solely upon the dolomite as a source of magnesia also necessarily entails production of the lime content of the dolomite as a by-product. It was recognized that if such a lime by-product could be produced in a suitable form, a plant location in northwestern Ohio would also be advantageous for marketing the lime. The problem, therefore, was the discovery and development of a low-cost process for separating the magnesia and lime of the dolomite of northwestern Ohio and producing each in a form relatively free from the other, and meeting the respective requirements of established markets. A satisfactory solution of this problem has been found, and it is the

Jan 1, 1948

By Vince Poxleitner, John Delaney

Introduction In the mining industry, comminution typically begins in the mine with a blast of explosive to break rock so that it can be handled by the avail- able equipment. Though the breaking of the rock is essential for material handling in the mine and plant, this is not the only reason for ore reduction in the mineral industry. Comminution is primarily required to liberate minerals from ore so that they can be concentrated into valuable products or to prepare minerals, such as talc, for market as mined. The size of the equipment used for comminution ranges from small to massive. Regardless of equipment size, the power required to reduce the rock is a major expense. Years ago, mineral deposits could be mined more selectively and required less grinding to produce saleable products. However, as richer deposits were depleted, it has become necessary to mine low-grade deposits, with the valuables dispersed throughout the rock. Comminution equipment evolved in physical size and capacities to meet the new mine and process requirements. A modern, modestly sized mining operation, with a capacity of 25.000 TPD may use crushers two stories high weighing over a million pounds, SAG mills 36 feet in diameter driven by dual motors putting out 10.000 horse- power, and ball mills over 16 feet in diameter . Because of the increasing size and weight of modern equipment, the builder is faced with the problems of erecting massive foundations and field assembling equipment on these foundations. As equipment costs and plant construction costs increase, builders are finding that owners are asking for more detailed progress and cost data. Comminution construction projects are generally classified as grass roots plants or expansions or modifications to existing plants. It is estimated that 60-702 of the jobs involve existing plants. Today ' s costs for construction of small, new comminution plants can reach tens of millions of dollars; larger plants can cost hundreds of millions of dollars. Modifications, on the other hand, usually run from a few thousand to several million dollars. Though many problems are common to the construction of both new plants

Jan 1, 1982

By Stanley J. LeFond, Langtry E. Lynd

Elemental titanium has become famous as a space age metal, because of its high strength/ weight ratio and resistance to corrosion. However, the major use is in the form of titanium dioxide pigment, which because of its whiteness, high refractive index, and resulting light-scattering ability, is unequaled for whitening paints, paper, rubber, plastics, and other materials. A relatively minor use is in welding rod coatings, in the form of the mineral, rutile. United States consumption of titanium dioxide for these applications in 1973 was as follows: pigments, 891,000 st; titanium metal, 40,000 st (=24,000 st metal) ; and welding rod coatings, 10,000 st. The only commercially important titanium ore minerals at the present time are ilmenite, and its alteration products, and rutile. Titanium was discovered by Gregor in 1790, as a white oxide which he recovered from menaccanite, a variety of ilmenite occurring as a black sand near Falmouth, Cornwall. Barksdale (1966) states that the fundamental chemical reactions on which the present day titanium industry is based were known before 1800, although it was not until 1918 that these pigments were available commercially on the American market. Pings (1972) outlines the early history of the titanium industry in the United States, referring to the work of Guise (1964) and others: "Mining of titanium minerals in the United States began sometime between 1880 and 1900, in Chester County, Pennsylvania. Small quantities of rutile were also produced during the time in North Carolina and Georgia. In 1901 rutile was mined from a deposit near Roseland, Va., and was used in making titanium chemicals and for coloring ceramics. Ilmenite in the deposit was produced as a separate item in 1913, and rutile and ilmenite were obtained from this deposit through 1921. The mining of titanium bearing beach sands began in 1916 near Mineral City and Pablo Beach, Fla., for the purpose of making titanium tetrachloride to be used in tracer bullets, flares, and smokescreens. The titanium pigment industry was founded in 1918. By 1928 a large part of the domestic production of rutile, ilmenite, and zircon came from Florida. However production ceased in 1929 in Florida with the mining of newly discovered deposits in Virginia which were mined until 1968. The discovery of new deposits in Florida, and the development of better mining and concentrating methods for lowgrade sands, led to a return of activities in this area. The large deposits near Tahawus, N.Y., were brought into production in 1942 by National Lead Co. (now N L Industries Inc.), and by 1949 this company was the lead¬ing producer of ilmenite in the world. In 1948 the extensive deposits of ilmenite and ilmenite-bearing iron ores were discovered in eastern Quebec." The beginning of the modern titanium metal industry was in 1948, when E. I. du Pont de Nemours & Co. produced the first metal. U.S. Bureau of Mines reports, which gave details of the Kroll process, together with the attractive properties of the metal for

Jan 1, 1975

By J. Chipman, N. J. Grant, H. L. Bishop

In a recent review1 of the iron-oxide activity of simple open-hearth type slags containing lime, magnesia, silica, and iron oxide, it was established that activity values were lacking in the range of slag composition most interesting to open-hearth operators. That is, laboratory experimental results were not available for simple slag compositions having basicity ratios ranging from about 1.5 to 4.0 and iron-oxide concentrations ranging from about 10 to 30 pct. However, in order to estimate iron-oxide activities for the range of slag compositions of interest to open-hearth operators, activity values were calculated using ionic concepts of slags by following the treatment of Flood and Grjotheim.2 A study of these data indicated that the desired slag compositions were probably not obtained due to the limited solubility of the basic constituents lime and magnesia. Thus, it was decided to make additional heats equilibrating molten carbon-free iron with slags containing MnO, P2O5, and A12O3 in percentages commonly found in open-hearth refining slags, and to examine the liquid slag field over a greater composition range than that of the simple slags studied previously. The ranges of the constituents used in the present study are summarized as follows: Metal Pct 0 0.070-0.201 PctMn 0.005-0.142 Slag Pet CaO 24.45-41.90 Pct MgO 3.23- 9.21 Pct MnO 1.13- 9.62 Pct FeO 20.08-48.7 Pet Fe2o3 2.23-10.14 Pct SiO, 3.51-28.50 Pct P2O5 0.23- 3.65 Pct A12O3 0.43- 3.38 The experimental procedure consisted of equilibrating the slag and metal in an induction furnace provided with a carbon-arc slag heater as described previously by Winkler and chipman.3 The data obtained were used to establish both the iron-oxide activity values and the activity coefficients for FeO and MnO in the open-hearth type slags. RESULTS AND DISCUSSION I) Iron-Oxide Activity of Open-Hearth Type Slags—The iron-oxide activities of various iron-oxide contents of the open-hearth type slags, as calculated from the oxygen content of the metal by the method of Taylor and Chipman, 4 are plotted against the molar ratio of bases to acids in Fig. 1. The data represent slag-metal studies carried out in the temperature range 1530" to 1700°C. For the purpose of this study, the constituents CaO, MgO, and MnO are assumed equal on a molar basis in their effect on the iron-oxide activity; and, PO2,5 is calculated as an equivalent to SiO2 and A12O3 on a molar basis. The total iron content of the slag is calculated to FeO and is represented by the symbol (FeO) The points presented in Fig. 1 represent the results of this investigation for iron-oxide slags containing between 1 and 8 mol pct MnO and less than about 2 mol pct each of A12o3 and PO2,5 The curved lines shown in Fig. 1 represent the results of the summary1 of the iron-oxide activities of simple slags of the system (CaO + MgO) - SiOz - (FeO),. The solid portions of the curved lines represent experimental data, and the broken portions represent iron-oxide activities calculated from ionic concepts of slags. The iron-oxide values of the open-hearth type slags of this study are in good agreement with the values of the simple iron-oxide slag system. Only a slight modification is indicated. The results of this study suggest that the line for 20 mol pct (FeO), should be lowered slightly in the basicity range 1.5 to about 3.

Jan 1, 1959

By Stanley J. LeFond, Langtry E. Lynd

Elemental titanium has become famous as a space age metal, because of its high strength/ weight ratio and resistance to corrosion. However, the major use is in the form of titanium dioxide pigment, which because of its whiteness, high refractive index, and resulting light-scattering ability, is unequaled for whitening paints, paper, rubber, plastics, and other materials. A relatively minor use is in welding rod coatings, in the form of the mineral rutile. United States consumption of titanium dioxide in concentrates for these applications in 1982 was as follows: pigments, 630,000 st; titanium metal, 31,000 st; and welding rod coatings, 5,000 st. The only commercially important titanium ore minerals at the present time are ilmenite and its alteration products, and rutile. Titanium was discovered by Gregor in 1790, as a white oxide which he recovered from menaccanite, a variety of ilmenite occurring as a black sand near Falmouth, Cornwall. Barksdale (1966) stated that the fundamental chemical reactions on which the present day titanium industry is based were known before 1800, although it was not until 1918 that these pigments were available commercially on the American market. Pings (1972) outlined the early history of the titanium industry in the United States, referring to the work of Guise (1964) and others: Mining of titanium minerals in the United States began sometime between 1880 and 1900, in Chester County, Pennsylvania. Small quantities of rutile were also produced during that time in North Carolina and Georgia. In 1901 rutile was mined from a deposit near Roseland, VA, and was used in making titanium chemicals and for coloring ceramics. Ilmenite in the deposit was produced as a separate item in 1913, and rutile and ilmenite were obtained from this deposit through 1921. The mining of titanium bearing beach sands began in 1916 near Mineral City and Pablo Beach, FL, for the purpose of making titanium tetrachloride to be used in tracer bullets, flares, and smokescreens. The titanium pigment industry was founded in 1918. By 1928 a large part of the domestic production of rutile, ilmenite, and zircon came from Florida. However production ceased in 1929 in Florida with the mining of newly discovered deposits in Virginia which were mined until 1968. The discovery of new deposits in Florida, and the development of better mining and concentrating methods for low-grade sands, led to a return of activities in this area. The large deposits near Tahawus, NY. were brought into production in 1942 by National Lead Co. (now NL Industries, Inc.), and by 1949 this company was the leading producer of ilmenite in the world. In 1948 the extensive deposits of ilmenite and ilmenite-bearing iron ores were discovered in eastern Quebec. The large Trail Ridge sand deposit in Florida was discovered by E. I. du Pont de Nemours & Co., Inc. geologists in cooperation with the US

Jan 1, 1983

By O. J. Huber, P. D. Frost, H. A. Robinson

URANIUM is a (ß-stabilizing, eutectoid-forming element when alloyed with titanium. Udy and Boulger1 have reported that an alloy of 48.6 pet uranium in titanium retains ß-phase when quenched from the ß-region, while compositions of approximately 30 pet uranium transform to martensite during such quenching. Since the eutectoid reaction is rather sluggish, titanium-uranium alloys should be capable of being strengthened by the ß — a trans-

Jan 1, 1959

By R. C. Mahon

During the early history of the Lake Superior district, tamarack timber was used almost exclusively in the wines. It had the strength and lasting qualities necessary for a satisfactory mining timber. But about fifteen years ago the tamarack trees were attacked by a blight which killed nearly all of them so today no tamarack timber is available. Hemlock, spruce, and poplar have neither the necessary strength nor the lasting qualities for mining timber, but a large supply of maple and birch is available. These have ample strength but, when placed underground, fail by decay in about three years. This means that, instead of replacing the timber sets once in fifteen years, it must be done five or more times in the same space of time which involves not only incraesed timber costs but more important, an increased labor cost.

Jan 1, 1949

By Robert A. Rapp, Ronald L. Pastorek

Solid-state electrochemical measurements by three alternative experimental procedures were made with the cell FeO, Fe3O4 |Zro.85Cao.15O1.85 |Cu| Zr0.85CaO.15O1.85 | FeO, Fe304 to establish the solubility and diffusivity of oxygen in solid copper in the temperature range 800" to 1030°C. The solubility of oxygen in solid copper and the diflusivity of oxygen in solid copper Dgu = 1.7 X 10-2 exp(-16,000/RT) Cm2/sec were determined and confirmed in alternative experiments. The enthalpy of solution of oxygen in solid copper equals —10,000 cal per mole; the partial excess entropy of the oxygen atoms in the Cu-O dilute solution is approximately the same as that found for interstitial atoms in other metals. The diffusivity of oxygen in solid copper is consistent with that expected for an interstitial atom. RELIABLE values for the saturation solubility N(s) and diffusivity DO of oxygen in solid copper have not been unambiguously established in the literature. Following three early determinations by others,1"3 Rhines and Mathewson4 reported that the solubility of oxygen in solid copper increased from 0.007 at. pct 0 at 600°C to about 0.015 pct at 1050°C. Phillips and skinner,, using essentiially the same analytical procedure, reported that the solid solubility increases from 0.0018 at. pct 0 at 550°C to about 0.0075 pct at 1050OC. The only previous value for the diffusivity of oxygen in solid copper was reported by Ransley.6 Ransley deoxidized Cu-Cu2O alloys in an atmosphere of carbon monoxide gas to yield a solubility-diffusivity product. He used the solubility data of Rhines and Mathewson to calculate the diffusivity values. Another method for obtaining the solubility-diffusivity product (N(s) DO) is by measuring the widths of internal-oxidation zones in copper alloys as reported by Verfurth and Rapp.7 However, the calculated N(S)Do products depend upon the alloy content of the specimen, so that the internal oxidation of copper alloys does not follow ideal internal oxidation kinetics. As a result, unequivocal values for the N(s) DO product were not obtained by this procedure. A solid-state coulometric titration technique similar to that employed in this work was introduced by C. Wagner8 to study the dependence on silver activity of the Ag/S ratio in silver sulfide in the temperature range of 160" to 300°C. Similar experiments have been carried out by C. Wagner and co-workers9-11 to study the stoichiometry range of silver and copper tellurides, cuprous sulfide, and cuprous selenide. Numerous authors have carried out electrochemical measurements with a solid oxygen-ion-conducting electrolyte to determine the solubility and/or diffusivity of dissolved oxygen in several liquid metals.12-l6 Rickert and Steiner17,18 have used solid-state electrochemical measurements to determine the diffusivity of oxygen in solid silver from 760" to 900°C. Two different cell geometries were used. In the cell of linear geometry Fe, FeO | ZrO2 + (CaO) | Ag + [0 (dissolved)] [1] oxygen diffused from the interior of the silver electrode to the silver/electrolyte interface where the oxygen activity had been lowered from a fixed initial value to practically zero by the application of voltage to the cell. The diffusivity of oxygen in solid silver was determined from the solution of the diffusion equation and the time dependence of the cell current. However, this determination of the diffusion coefficient depended upon a knowledge of the solubility of oxygen in solid silver. A cylindrical geometry was used for the cell Pt, O2(Po2 = 0.21 atm) | ZrO2 + (CaO) | Ag + [0 (dissolved)] [II] which also allowed the diffusivity of oxygen in solid silver to be determined. These values were in agreement with other available data.l9 Recently, Raleigh20,21 used a method involving the measurement of diffusion-limited currents in a cell involving the AgBr solid electrolyte to determine the diffusion coefficient of silver in Ag-Au alloys at 400°C. Diffusivity values on the order of l0-14 sq cm per sec were measured in the alloy composition range 10 to 60 at. pct Ag in a single experiment. From numerous electrical conductivity and galvanic cell measurements,9'22"26 the solid solution Zr0.85 Ca0.15 O1.85 has been established as an electrolyte with predominant oxygen ion conduction over a wide range of intermediate and high oxygen activities. For interrelating the thermodynamics and the kinetics of the dissolution of oxygen in solid copper in this investigation, a galvanic cell was constructed with FeO-Fe3O4 as the reversible reference electrode, the Zr0.85Ca0.15 O1.85 electrolyte, and a pure copper specimen under-saturated in oxygen as the other electrode. THEORETICAL ANALYSIS Three variations of a high-temperature electrochemical technique were used in this study to provide two determinations each of the solubility and diffusivity

Jan 1, 1970

By M. L. Rudee, R. A. Huggins

Samples of iron containing nitrogen or carbon have been given treatments similar to those used in cold-work peak (CWP) measurements and examined by transmission electron microscopy. It was observed that the unusual and nonreproducible behavior of the carbon CWP can be explained by a strong tendency for carbon to form grain boundary precipitates at temperatures below those used for CWP measurements. These precipitates dissolved at the temperatures used in the CWP measurements. There was no evidence for nitrogen precipitation at grain boundaries. There was no indication of precipitation along dislocations in either carburized or nitrided samples given treatments similar to those of CWP measurements. Although it is possible that subelectron-microscopic clustering had occurred, this observation supports the theories of the CWP that are based on continuous atmospheres rather than on individual precipitates. In an earlier paper,' the present authors developed a new distribution function to predict the occupation of sites for interstitial impurity atoms around a dislocation. When this distribution was applied to the case of carbon and nitrogen in iron, it predicted that, if the temperature dependence of the concentration of solute atoms in the matrix was controlled by the presence of equilibrium carbide or nitride precipitates, the tendency for nitrogen to segregate to dislocations would be greater than that for carbon even though their binding energies to dislocations are identical. The cold-work internal-friction peak (CWP) is considered by most authors to be produced by the interaction of interstitial impurities with dislocations. Many investigators have studied the CWP in iron containing carbon and nitrogen and have observed a significant difference between its behavior in the two cases. In this paper a series of experiments will be reported that were initiated to determine whether the application of the new distribution function would explain the observed differences between the carbon and nitrogen CWP. Although it was found that the distribution function, pev se, did not explain the differences, the differences became clear, and some insight into the mechanism of the CWP was realized. Before reporting the present experiments, the literature pertaining to the differences between the carbon and nitrogen CWP in iron and the various mechanisms proposed for the CWP will be reviewed. LITERATURE REVIEW Snoek2 first observed the CWP in iron specimens containing nitrogen, but also reported a weak and unreliable peak in carburized samples. Later, Ke3 established that the CWP height was proportional to the degree of deformation. The presence of nitrogen alone would produce a peak of the same size as found in a sample containing both nitrogen and carbon, and KG concluded that the CWP was caused by nitrogen. In a discussion of G's paper it was reported that Dijkstra had investigated the CWP in samples that contained only carbon. He found it to be much smaller than the nitrogen peak and "unstable". KG et al.4 charged specimens of iron with both carbon and nitrogen. They observed that the carbon CWP was much smaller than that observed in nitrided samples, but that aging at 300°C caused the carbon peak to increase. A similar treatment produced no change in a nitrogen peak. Annealing at higher temperatures caused the height of the CWP in both the nitrogen and carbon samples to decrease. This behavior was also observed by Kamber et al. 5 who found that the activation energy for the annealing of the CWP was identical with the activation energy for the self-diffusion of iron. They concluded that the annihilation of dislocations by climb caused the reduction in the CWP height. Kamber et al. studied the "unstable" carbon peak in detail. They measured both the Snoek and CWP during various aging treatments. In carburized samples, aging at 100°C caused the Snoek peak to disappear, although the CWP peak remained small. However, a subsequent treatment for 5 hr at 240°C caused the CWP to reach a maximum. They proposed that an additional location for the carbon, other than whatever site produced the CWP, is present. In nitrided samples the CWP was completely developed as soon as a measurement was made; additional sites are not present. No explanation of either the additional site or the difference in the behavior of carbon and nitrogen was offered. petarra5 performed a systematic study of the effect of composition on the CWP. Using three kinds of "pure" iron, he showed that there was a residual CWP when the carbon and nitrogen concentrations had been reduced to less than that detectable by Snoek-peak measurements. He observed the same general annealing behavior and composition dependence as previous investigators, with the following exceptions. On first measuring the carbon CWP, it was found to be identical with the residual peak, and essentially independent of the carbon content. If the CWP was measured a second time in the same sample, it increased in size, but was still only about one-fourth the size of a CWP in a sample of the same iron nitrided to the same composition. On the other hand, a series of annealing experiments showed that the nitrogen CWP was not al-

Jan 1, 1967

By J. P. Morgan, J. R. Dawson

A review of Jones & Laughlin's Vesta coal preparation plant flowsheet in 1959 indicated that, of the three sizes making up the company's metallurgical product, the -48-mesh fraction offered the greatest potential for improvement. Efforts to improve quality and yield were directed toward the use of froth flotation. After laboratory and plant tests were completed, it was decided that treatment of the entire -@-mesh fraction of the coal mix by flotation was a significant step forward in providing better coal quality and recovery. Reduction of ash in the metallurgical coal from 6.75 to 6.25% provided material savings at the coal washer, coke plant and blast furnace. The Vesta preparation plant of the Jones & Laughlin Steel Corp. is located on the Monongahela River, eight miles south of Brownsville, Pa. This plant is capable of cleaning 2400 tph of a mixture of Pittsburgh seam high volatile coals from the Vesta and Shannopin mines to produce a metallurgical coal of 6.75% ash and a steam coal containing 14% ash. Heavy media vessels are utilized to upgrade the coarse coal while launders and tables handle the -1/4-in. fraction of the raw coal feed. Cleaned -1/4-in. coal goes to boot classifiers where the 48 mesh material settles and is removed by bucket elevators. This 1/4x48 mesh material is then sent to centrifugal dryers for dewatering. As designed, the -48 mesh fraction overflows the boot as a 7 to 8% solids slurry and is recovered by cyclones and a thickener and then dewatered by filtration. The products of the heavy media circuit, the launder-table-boot-dryer circuit and the cyclone-thickener-filter circuit, are combined to constitute the shipped metallurgical coal. The Vesta-Shannopin mix is combined at Jones & Laughlin's Pittsburgh and Aliquippa Works with purchased low and medium volatile coals for the production of metallurgical coke. A review of the preparation plant flowsheet in 1959 indicated that of the three sizes making up the metallurgical product, the -48-mesh fraction, containing 16.7% ash, offered the greatest potential for improvement of coal quality. This fraction was considered low, averaging 55%, while the rejects averaged only 33% ash. It was evident that while the boot overflow fraction represented only 6% of the product mix, a reduction in ash to 6 or 7% would reduce the ash of the metallurgical coal by 0.5%. The original plant flowsheet handling the boot overflow (Fig. 1) consisted of ten 14-in. cyclones treating 70% of the overflow and producing a 48x150-mesh underflow and a -150-mesh overflow. The cyclone overflow was combined with the remaining 30% of the boot overflow to constitute the feed to a 150-ft thickener. The cyclone underflow and the 48x325-mesh thickener underflow were fed to vacuum disk filters for dewatering prior to combination with the coarse metallurgical coal. The thickener overflow stream containing the -325-mesh material was split, 40% being recycled directly to the plant water system and 60% being flocculated and clarified in an 85-ft thickener. The underflow of this thickener was pumped to waste. Efforts to improve quality and yield on the -48-mesh coal were directed toward the use of froth flotation early in the investigation. A review of the fine coal flowsheet indicated that three streams should be considered for treatment by froth flotation. These were the -325-mesh thickener overflow, the -150-mesh cyclone overflow and the -48-mesh boot overflow. Flotation of the thickener overflow was considered basically to improve recovery by upgrading the reject stream so that it could be fed to the filters along with the thickener and cyclone underflows. The overflow from the thickener contained 3.7% solids and had an ash analysis of 33%. Flotation of the cyclone overflow and boot overflow were considered to upgrade coal quality as well as yield. Treatment of the cyclone overflow, a 5.4% solids slurry averaging 27% ash, would eliminate the coal thickener and provide a feed of flotation concentrate and cyclone underflow to the filters. Treatment of the boot overflow would eliminate both the cyclones and thickener and feed only flotation concentrates to the filters. Flotation feed in this

Jan 1, 1962

By G. V. Smith, C. O. Tarr, R. F. Miller

Metallurgists have long recognized that the Fe3C type of carbide is not a stable phase in steel and that, given sufficient time, it will decompose with formation of graphite, at least at temperatures below about 205o°F.l This slow graphitization has never been of much concern to users of hypoeutectoid steels, for in such steels it occurs only during prolonged exposure at elevated temperature. In I935, Kinzel and Moore2 reported complete graphitization of the carbide of a 0. I5 per cent carbon steel that had been in service for about 26,000 hr. at a temperature of about IIOO° to 12oo°F. In discussing this paper, both Mathews and Sauveur reported graphitization of slightly hypereutectoid steels. In an extensive investigation of graphitization of high-purity iron-carbon alloys, Wells1 produced graphite at 7oo°C. (I29o°F.) in an alloy containing as little as 0.13 per cent carbon, and showed that graphite may precipitate either directly from austenite or from decomposition of Fe3C, also that the rate of graphitization tends to increase with increase of temperature, at least up to a certain point, and with the presence of graphite nuclei. Jenkins, Mellor and Jenkinson,3 studying the structural changes in carbon steels ranging from 0.17 to 1.14 per cent carbon during stress-rupture tests at elevated temperature, observed graphite in an aluminum-killed 0.40 per cent carbon steel after fracture in 4340 hr. under a stress of 17,900 lb. per sq. in. at 840°F. and in many of the steels after stress-rupture test at 12oo°F. Unstressed samples of 0.60, 0.86 and I.I4 per cent carbon steels graphitized almost completely in 360 hr. at I200°F., while samples of 0.90 and 1.10 per cent carbon steels were somewhat graphitized, and steels containing 0.24, 0.40 and 0.57 per cent carbon showed no sign of graphite. Their results indicate that graphite precipitates the more readily the higher the carbon content, though this tendency is influenced to some extent by deoxidation practice, and that graphitization is accelerated by plastic deformation. In a recent study of the graphitization of 0.4 and 0.6 per cent carbon steel strip, made by M. A. Hughes,4 it was found that at I2oo°F. graphite appeared after 72 hr. in aluminum-killed but not in silicon-killed steel, and that the rate of graphitization was increased by slow cooling from I6oo°F. and by cold-working prior to annealing. In creep tests of normalized 0.15 per cent carbon steel killed with 2.5 Ib. of aluminum per ton, at the U. S. Steel Corporation laboratory at Kearny, spheroidization and some graphitization were found after 3000 hr. at Iooo°F. under a stress as low as 1500 lb. per sq. inch. .

Jan 1, 1944

By C. O. Tarr, G. V. Smith, R. F. Miller

Metallurgists have long recognized that the Fe3C type of carbide is not a stable phase in steel and that, given sufficient time, it will decompose with formation of graphite, at least at temperatures below about 205o°F.l This slow graphitization has never been of much concern to users of hypoeutectoid steels, for in such steels it occurs only during prolonged exposure at elevated temperature. In I935, Kinzel and Moore2 reported complete graphitization of the carbide of a 0. I5 per cent carbon steel that had been in service for about 26,000 hr. at a temperature of about IIOO° to 12oo°F. In discussing this paper, both Mathews and Sauveur reported graphitization of slightly hypereutectoid steels. In an extensive investigation of graphitization of high-purity iron-carbon alloys, Wells1 produced graphite at 7oo°C. (I29o°F.) in an alloy containing as little as 0.13 per cent carbon, and showed that graphite may precipitate either directly from austenite or from decomposition of Fe3C, also that the rate of graphitization tends to increase with increase of temperature, at least up to a certain point, and with the presence of graphite nuclei. Jenkins, Mellor and Jenkinson,3 studying the structural changes in carbon steels ranging from 0.17 to 1.14 per cent carbon during stress-rupture tests at elevated temperature, observed graphite in an aluminum-killed 0.40 per cent carbon steel after fracture in 4340 hr. under a stress of 17,900 lb. per sq. in. at 840°F. and in many of the steels after stress-rupture test at 12oo°F. Unstressed samples of 0.60, 0.86 and I.I4 per cent carbon steels graphitized almost completely in 360 hr. at I200°F., while samples of 0.90 and 1.10 per cent carbon steels were somewhat graphitized, and steels containing 0.24, 0.40 and 0.57 per cent carbon showed no sign of graphite. Their results indicate that graphite precipitates the more readily the higher the carbon content, though this tendency is influenced to some extent by deoxidation practice, and that graphitization is accelerated by plastic deformation. In a recent study of the graphitization of 0.4 and 0.6 per cent carbon steel strip, made by M. A. Hughes,4 it was found that at I2oo°F. graphite appeared after 72 hr. in aluminum-killed but not in silicon-killed steel, and that the rate of graphitization was increased by slow cooling from I6oo°F. and by cold-working prior to annealing. In creep tests of normalized 0.15 per cent carbon steel killed with 2.5 Ib. of aluminum per ton, at the U. S. Steel Corporation laboratory at Kearny, spheroidization and some graphitization were found after 3000 hr. at Iooo°F. under a stress as low as 1500 lb. per sq. inch. .

Jan 1, 1944

By H. B. Pulsifier

The following report on the structure of copper is the result of work done in the laboratory of the Rome Wire Co. early in 1925. Previous work had indicated to the author that excellent results might be expected if a suitable technique in surfacing and etching could be developed. The resources of the Rome Wire Co. enabled this metal to be studied more intensively than previously, although none of the materials were investigated exhaustively. The material presented is hardly more than a preliminary summary of the most obvious character of the metal; a much longer research would have been necessary to solve many of the pressing problems connected with the casting and working of copper. The Rome Wire Co. has generously released the paper; to Dr. E. H. Darby, technical supervisor of the company, are due particular thanks for his appreciation of the importance of metallographic study of copper and for his effective support in the prosecution of the work and in making its publication possible. Dendrites, Cells and Granules The microscope reveals at least three orders of structural units in copper, as in other metals. The largest structural units are the dendrites —spine-shaped crystals building up the masses of metal when it solidifies from a melt or is crystallized from solutions by chemicals or electricity. The nucleus cells, commonly called "grains," are the largest and most conspicuous units of recrystallized solid metal; the limitations of their growth usually confine their sizes to microscopic proportions, although in "single crystals" they may become as large as the piece of metal itself. The granules are the microscopic fine-structural blocks or laminations that build up the nucleus cells; they are revealed when metal is etched with suitable chemicals. The granules are commonly visible at 500 diameters magnification and are plainly visible at 1000 diameters; magnification of the granular surfaces up to 6000 to 9000 diameters shows no finer units, not because of lack of optical resolution but because the chemical attack has left an unfeatured smooth surface.

Jan 1, 1926

By S. A. Ford

I WOULD like to call the attention of our Bessemer steel manufacturers to a few facts in regard to the action of the manganese in. the Spiegel with the oxide of iron in the blown iron. The oxygen is united with the iron as a magnetic oxide, as Gautier asserts, and this oxide is reduced to a protoxide by one-atom of the manganese uniting with one atom of the oxygen of the magnetic oxide of iron. This oxide of manganese and the protoxide of iron thus formed unite with the silica which has been formed by the oxidation of the silicon in the pig iron. Mr. Snelus and others

Jan 1, 1881

POSITIONS VACANT (Under this heading will be published notes sent to the Secretary of the Institute by members or other persons.) Two metallurgical chemists for permanent positions in Siberia; one man to have charge of laboratory dealing with the products of a mill working silver-lead-zinc ores, the other to have charge of laboratory in connection with smelting these products. Established camp; living conditions and climate good. Married men preferred. No. 57. An electro-metallurgist for small zinc plant now under construction. Moderate salary to start but likelihood of substantial increase if plant proves successful. No. 58. Assistant engineer in office of consulting engineer for large mining and smelting company. Man desired with technical education and broad experience in mechanical line in connection with mining and smelting operations, capable of taking charge of office and drafting room. No. 59.

Jan 1, 1916

Despite the uncertain political situation in Katanga, copper and cobalt mining and metallurgical operations of Union Miniere du Haut-Katanga were normal at the time of this writing. In fact operations continued normal after Congo independence on June 30, 1960, with copper and zinc production for that year being slightly higher than in 1959. United Nations military operations from September to December 1961 resulted in death, shutdowns and damage to company persons and installations. Nevertheless, results for that year were only slightly off normal production. Moise Tshombe delared Katanga a separate state after Congo independence. Although there has been tribal unrest, the new State has enjoyed relative calm and normal industry by comparison with other parts of the Congo as a result of the firm hand of Moise Tshombe's government and the continued presence of Belgian officials and technicians.

Jan 12, 1962

By V. Koump, T. F. Perzak, R. G. Olsson

Jan 1, 1965