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By Mark Sheppard

DuRing the winter of 1937, the writer visited a West Virginia stone quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide. The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the overburden was dug by hand labor, dumped to the quarry floor and transported to the dump in quarry cars. Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semicircle around the top of the quarry about 200 ft. from the face and about 50 ft. above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river. As quarrying progressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp., 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft. above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting bays into which the motors are raised whenever the river reaches flood stage. An 8-in. suction line, equipped with a strainer and a foot valve, takes the water from a concrete sump at the edge of the river. The pipe line from the pump consists of 1750 ft. of 8-in. spiral steel flanged pipe and

Jan 1, 1938

By Mark Sheppard

DuRing the winter of 1937, the writer visited a West Virginia stone quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide. The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the overburden was dug by hand labor, dumped to the quarry floor and transported to the dump in quarry cars. Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semicircle around the top of the quarry about 200 ft. from the face and about 50 ft. above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river. As quarrying progressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp., 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft. above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting bays into which the motors are raised whenever the river reaches flood stage. An 8-in. suction line, equipped with a strainer and a foot valve, takes the water from a concrete sump at the edge of the river. The pipe line from the pump consists of 1750 ft. of 8-in. spiral steel flanged pipe and

Jan 1, 1938

By V. O. Struebing, K. A. Johnson, F. H. Ellinger, C. C. Land

The system Pu-Ce-(Pu,Ce)Co2 has been investigated by thermal, micrographic, and X-ray diffraction methods. The principal characteristics of this system were found to be: 1) a "eutectic" valley extending across the phase diagram from the 6 Pu/Pu3 Co eutectic at 408°C to the 7 Ce/Ce24Co11 eutectic at 42Z°C, the maximum temperature in the valley being 443°C; 2) three ternary inter mediate phases, each having an appreciable mutual solid solubility for plutonium and cerium; 3) eleven four-phase equilibria comprising one eutectic, four peritectics, and six Class II reactions; 4) a pseudobinary section between the ternary phase A and the y Ce solid solution, and another such section between the 6 Pu solid solution and the 77 phase (Pu3 CO) containing cerium in solid solution. ALLOYS of plutonium that are completely molten at temperatures in the vicinity of 500°C are of interest for use in the liquid state as fast reactor fuels. Since exploratory work1,2 on Pu-Ce-Co alloys had indicated that a "eutectic" valley extended across the ternary system between the Pu-co[6 Pu/n (Pu3co)] and the Ce-CO[? Ce/? (ce24co11)] binary eutectics, thus allowing alloys of any plutonium concentration to be selected within the temperature limitation, the present investigation was begun for the purpose of accurately defining the "eutectic" valley. However, when the existence of three ternary intermediate phases and the consequent complexity of the liquid-solid equilibria were discovered, the investigation was expanded to cover the complete system, except for the corner between the isomorphous series PuCo2-CeCo2 and pure cobalt. 1) EXPERIMENTAL The purities of the alloy components are given in Table I. Alloy buttons, each weighing about 5 g, were prepared by arc melting the accurately weighed components (±0.1 mg) under a purified He-Ar atmosphere. Weight loss as a result of melting was usually less than 1 part in 1000; therefore, the buttons were not chemically analyzed. One half of each button was re-melted by induction heating in vacuo in a magnesia crucible and furnace-cooled, so as to produce a somewhat coarser microstructure than that of the rapidly cooled arc-melted button. About 20 g of each alloy to be examined by thermal analysis were remelted in vacuo in a tantalum crucible of the inverted thermocouple well type. The thermal, X-ray diffraction, heat treating, and micrographic methods used in this investigation were essentially the same as have been described in earlier reports.3,4 Two etchants for these alloys were found to be satisfactory: A) 8 (parts by volume) orthophosphoric acid, 5 ethanol, and 5 glycol; B) 7 tetraphosphoric acid, 36 water, and 57 2-ethoxyethanol. Etchant A was used at 5 to 6 v for 15 to 20 sec, and etchant B was used at 10 v for 10 sec. 2) THE LIMITING BINARY SYSTEMS 2.1) Plutonium-Cerium. The phase relationships of this system have been investigated by Ellinger, Land, and Cramer,5 and by Selle and Etter.6 The over-all agreement of their phase diagrams is good, but there are some discrepancies in the positions of some phase

Jan 1, 1967

By Robert C. Freas, William H. Lane, David J. Krohn

Maysville, Kentucky, is the location of Dravo Lime company's new 9 070 mt(10,000 tons) per day limestone mine and 2 727 mt (3000 tons) per day calcination plant. This complex has been designed as a unit, and capitalizes on recent technological advances. Geologically the Dravo facility is.located on the eastern flank of the Cincinnati Arch, with the high calcium limestone being found at a depth of approximately 274 m (900 ft). A systems approach has been employed in designing and equipping the mine. This allows for maximum production while maintaining flexibility and versatility. All hydraulic drilling, large LHD equipment, and an extendable mine conveyor system are utilized in a five-part mining cycle. A modular concept modifies the room and pillar system to optimize ventilation. The limestone is calcined in three 5.18 x 61.87 m (17 x 203 ft) KVS rotary kilns with a design capacity of 909 mt (1000 tons) per day of lime per kiln. The kilns are coal fired and equipped with preheaters. Lime storage is provided by three 57 m (186 ft) high concrete silos with a combined capacity of 40 909 mt.

Jan 1, 1976

By Harry M. Mikami, Harold A. Heiligman

The term chromite is often used to cover all chrome ores and concentrates which are the designations actually employed by most industrial users and producers. Chromite is, of course, the predominant mineral in chrome ores which are aggregates of chromite and gangue minerals, most commonly magnesian silicates. Most chrome ores now used are petrographically chromitites which is a rock, rather than an individual mineral term. The title of this chapter, following past usage, is meant in the broad sense. Chromite is the only primary commercial source of chromium. This book is intended to deal only with nonmetallic minerals and rocks. Chromite uses are both metallic (ferroalloys) and nonmetallic (refractories and chemicals), but any discussion of mineralogy, geology, and mining is common for all types. A number of the major commercial ores are interchangeable, and their uses are determined by economic considerations. Mineralogy COMPOSITION [Chromite is a term embracing a wide range of chromium-bearing spinel*-group minerals the compositions of which can be expressed as (Mg,Fe2+)0.(Cr,AI,Fe3+ )203.t This can be further generalized to R2+O.R23+03 and the molecular RO: R,03 ratio is generally close to one. The exact composition may be arbitrarily expressed as a solid solution of six end-members, MgAI2O4, MgCr2O4, MgFe2O4 FeCr2O4, FeA12O4, and Fe2+.Fe23+04 with the proportions of each stated or plotted in the six-cornered prism of Johnston and Stevens61 or this scheme may be simplified to comprise only three endmembers making up a composition triangle, Fig 1. [ ] FIG 1-Triaxial composition diagram showing molecular compositions of chromites listed in Table 1. The RO.Fe2O3 corner includes Ti02 norms listed in the table. Note that the high-alumina chromites (1, 2, 3, 9), form a cluster well apart from the high-Cr2O3 chromites. There is a tendency for high-iron chromites to be higher in Fe2O3 + TiO2 but it is not a consistent or systematic relationship. The MgO/RO ratio must also be known to characterize the chromite chemically. Simplest and most convenient method is Thayer's' in which a chromite such as (Mg.65Fe.352+)O.(Cr.40Al.50Fe.103+)203 would be written as (Mge,) Cr eAlo. The molecular pro- portions of ferrous and ferric oxides can be seen at a glance as the divalent and trivalent oxides]

Jan 1, 1960

By J. F. Breedis

The morphology and crystallography of marten -site formed during quenclzing were examined by transmission electron microscopy in alloys whose compositions lie between Fe-19 wt pct Cr-11 wt pet Ni and Fe-33 wt pet Ni. Depending upon composition, four types of transformation products were observed in this portion of the Fe-Cr-Ni system. Three of these products are denoted as lath, plate, and surface martensite. The hcp structure found in association with martensite it? certain higlz-clzromium stairzless steels does not influence the transformation crystallogyaphy and cannot he Considered to act as a transitional structure in the transformation from austenite to martensite. DISTINCT changes in the characteristics of mar-tensitic transformations are produced by the partial replacement of nickel by chromium in austenitic iron alloys. Large, discrete martensite plates form in Fe-30 pct Ni, while aggregates of small crystals form in sheets parallel with (111)A austenite planes in stainless steels containing in the vicinity of 18 pct Cr and 8 pct Ni. These transformation structures are termed plate and lath martensite, respectively. In addition, an hcp structure (E) has been observed in certain high-chromium stainless steels in association with martensite. It has been suggested that the E structure acts as an intermediate step in the nucleation of martensitel-' and influences the morphology and crystallography of the final transformation product. However, this structure may be the consequence of the deformation of austenite arising from the shape distortion accompanying the transformation to bcc martensite.' An hcp stacking of atoms is produced in an fcc lattice by the introduction of an intrinsic stacking fault on alternate (111)A planes. At present, the arguments behind each explanation for the E structure are based on circumstantial evidence. This investigation seeks to determine the importance of the hcp structure to the martensitic transformation, and to examine the crystallography and morphology of the structures produced in bulk samples by quenching in alloys whose compositions lie between Fe-19 wt pct Cr-11 wt pct Ni and Fe-33 wt pct Ni. These aspects of the martensitic transformations occurring in these alloys have been studied by electron metallography and selected-area diffraction of foils examined in transmission in the electron microscope. EXPERIMENTAL METHODS The alloys have been prepared from thoroughly mixed powders of pure iron (99.6+ pct, 0.06 max pct C, 0.15 max pct N), chromium (99.9+ pct), and nickel (99.9+ pct). After sintering in an Ar-H2 atmosphere, the mixtures were melted by induction heating in recrystallized alumina boats under purified argon. The ingots were cold-rolled to 0.4-mm sheets (100 pct reduction in thickness) after the removal of at least 1.5 mm from their surfaces by grinding. Three-millimeter discs were punched from the cold-rolled sheet and annealed at 1150°C for approximately 24 hr in evacuated quartz capsules. The average grain size was 40 µ Thin foils for examination in transmission in the electron microscope were prepared from discs quenched to liquid-nitrogen temperature. A depression was first jet-machined into the center of a disc and a thin area (less than approximately 3000A in thickness) was produced at the base of the depression by electropolishing. Electron-microscope studies utilized the RCA EMU-3 and the Philips EM-100B microscopes operating at 100 kv. X-ray examination of the polycrystalline alloys was performed either in a diffractometer or in a

Jan 1, 1964

By Eugene A. Stephenson, Leon J. Pepperberg

The natural gas industry is essentially a byproduct of the oil industry. When first discovered the gas was usually regarded as a nuisance, and even when found immediately associated with oil, or suspected to be so, it was customary to blow it freely into the air, in the hope that gas would shortly be replaced by a flow of oil. The early methods of production,' transportation, and utilization were neither economical nor efficient; the chief value of gas was based on its use as a fuel, with no regard to (1) its effect. upon the oil with which it was associated, (2) its importance as a lifting agent, or (3) to its possibilities for the manufacture of byproducts. The spread of scientific research among major gas and oil companies, manufacturers, and academic groups, has gradually brought recognition of the fact that many of the technical problems bearing on production, transportation and utilization are common to both the gas and the oil industry. A few of the problems that have received serious study, and from which highly gratifying results have been obtained, are: (1) gas measurement, (2) oil well and gas well capacity tests, (3) elimination of field waste, (4) design of pipe lines and compressor stations, (5) extraction of constituents which might interfere with the normal flow of gas through the line, (6) control of reservoir pressure, (7) estimation of reserves of both oil and gas, (8) control of water encroachment, (9) measurement of factors which influence fluid movement through sands, (10) solubility of gases in oils, (11) effect of dissolved gases upon the properties of the solution. These investigations have demonstrated that almost every problem relating to oil production also involves some phase of natural gas engineering. During the past 15 years the use of experimental methods for the discovery and measurement of obscure, influencing factors, which cannot be determined purely by hydrodynamic or thermodynamic analyses, have contributed much towards the progress made in solving the problems of the oil and gas industries. Most of the experimental work on

Jan 1, 1934

By J. W. Hickman

INTRODUCTION MOST of the electron diffraction studies of oxide films which form on metals and alloys have been carried out by oxidizing the specimens in an auxiliary furnace, cooling down to room temperature and then inserting the sample in the camera and taking a diffraction photograph. Prior to the investigations at the Westinghouse Research Laboratories, no systematic study of the oxidation of any metal or alloy had been made over a wide temperature range with photographs taken at elevated temperatures except the investigation of iron by Jackson and Quarrell.1 In order to interpret the electron diffraction patterns of the oxides which may form on alloys, it seemed advisable to make a comprehensive study of the oxidation of the purest metals available in the bulk form. Thus far studies have been completed on iron,2 cobalt,2 nickel,2 chromium ,2 Copper2, tungsten,3 molybdenum,3 beryllium ,4 uranium,4 thorium4 titanium5 and zirconium5 while investigations are in progress on tantalum and columbium, The next step in the study is concerned with the investigation of the oxides which form on binary alloys. Some of the alloys consist of a single, phase while others contain more than one phase. The identification of the oxides on the surface of a binary alloy may be more difficult than on the metals since oxides containing both metals as well as those of each metal separately may form. Furthermore, it is possible that solid solutions of the various oxides may occur. In these cases the electron diffraction method may not be able to identify the oxidation products unequivocally, Examples of solid solutions which may present difficulties are MoO2 and W02, Fe3O4 and FeO-Cr2O3 since the diffraction patterns are so similar. Since electron diffraction can determine only the structure of the oxide, the chemical composition must be inferred from the diffraction data based upon X ray diffraction investigations of specimens of known composition. It is possible in some cases to decide whether a given solid phase reaction can occur on the surface of an alloy by a study of available thermodynamic data. Table I lists the equations for the formation of most of the oxides from the elements together with values of log10 KR for the several temperatures. [OF°R = -RT In KR 0 and-~75T = logi°KR] [?F°R] and KR are the free energy and equilibrium constant of the reaction. Calculations may be made using the values given in Table I to determine whether a given solid phase reaction is possible thermodynamically. For example, a solid phase reaction should occur between aluminum and chromium' oxide to form

Jan 1, 1948

By James Gayley

(Presented at the Washington meeting, May 3, 1905, and simultaneously sent to the Iron and Steel Institute, for presentation at the meeting of that Society in London, May 11, 1906.) It is to be regretted that the data respecting the use of dry-air blast, which were presented to the Iron and Steel Institute at its meeting in the United States, in October, 1904, should have been restricted to the period from August 25th to September 9th inclusive, and from September 17th to 30th inclusive. In order to present the paper at all it was found necessary to limit the record of operations to the period above stated. It appears from the reported discussion that some of those who took part felt themselves to be at a disadvantage, in considering the economies secured through the use of dry-air blast, by reason of the fact that the data presented covered so short a period. It is the purpose of this communication to present in detail the record of operations of the Isabella Furnaces from November, 1904, to March, 1905, inclusive, as shown by the furnace-records. The month of October, 1904, is not included, as the furnace using dry-air blast was stopped several times for repairs; and while the record was a very good one considering these interruptions, yet it was not a continuous one. The record for Furnace No. 1 using dry-air blast, and the comparison with Furnace No. 3 using natural air for the month of November, 1904, is set forth in Table I. By a reference to my former paper it will be noted that the month of November is the beginning in this country of the winter period, when the atmosphere decreases rapidly in humidity. The month of October being the transition month in the autumn and April in the spring, and furnaces using natu-ral air approach more closely, from November on, to the conditions obtained by the use of dry-air blast. The data given in

Jan 1, 1906

By C. C. Harris

The characteristics of some common size distribution equations are critically discussed. A generalized form of several well-known size distribution equations is obtained from a differential equation describing statistical distributions. The equation contains three parameters and can describe the major features of size distributions in the fine and the coarse size regions. A graphical method for its implementation is provided. The application of this and other equations to sets of data are compared both for the quality of fit and from a comminution kinetics viewpoint. If a narrow size range of a brittle material is broken sufficiently to obliterate the feed sizes, but not so severely that excessive secondary breakage occurs, a plot of cumulative fraction by weight undersize (Y) vs. sieve size (X) on a log log grid (sometimes called the Gates-Gaudin-Schuhmann plot)1-3 gives a straight line of slope ~ 1 over much of its range. On closer inspection, deviations in the extreme coarse range covering perhaps 10 to 30% of the total sample may be apparent (frequently, a change in slope from ~ 1 to a value different from unity) together with over-all slight irregular departures from a smooth line. One model of breakage postulates that the same fracture pattern persists throughout all size ranges.4 The fracture pattern is characterized by the slope of the line on a log log grid. Accordingly, a distribution of sizes broken in the manner specified earlier is expected to produce a size distribution having the same slope as that of a broken narrow size range. Additionally, the slight irregularities mentioned above should even out in the summation process, giving a smooth straight line of slope ~1 on a log log plot. This idealized state of affairs does not, however, describe the distribution of a multi-event process such as that for a tumbling mill product. On the average, these curves (plotted again log log) tend to have a slope in the fine region different from and usually less than unity,5 while the coarse size region curves with increasing slope for cases of mild reduction, and with decreasing slope when moderate or severe size reduction has occurred. In addition, there are usually slight irregular deviations from a smooth curve. Most curves display two distinct regions — fine and coarse — and a few curves show one or more intermediate regions. Feed size studies6,7 show an effect which theory is required to explain. For the same material, mill loading and other operating conditions, and the same time of grinding, plotting the dimensionless ratios Y vs. X/(feed size) does not reduce the size distribution data to the same curve. Y vs. X does not correlate the coarse region, but it can provide a crude correlation of the fine region which improves as size diminishes and as grinding time increases. The size distribution in the coarse region is somewhat more dependent on the feed size than is the distribution in the fine region wherein the degree of dependence diminishes as time proceeds; the distribution of sizes in the fine region are determined largely by the nature of the material and the comminution conditions. A comprehensive model of comminution must recognize that several different patterns or modes of breakage can occur in a mill; 5,8,9 that there can be some selection in that some size ranges are broken more than others; and that, while some particles may be the product of a single fracture event, or may even remain unbroken, others result from multiple rebreak-age. Thus, breakage in a tumbling mill is more complex than the Schuhmann4 model admits in that several different types of comminution micro-events* occur rather than just one type, and these events, though different in size scale, are of the same overall pattern as that which is visualized by the model. The concept of an average comminution micro-event has, therefore, a mathematical rather than a physical connotation, at least for tumbling milling. Equations with which to describe particle size distributions have been sought for over half a century.10-13 No equation presently in general use was first derived from an analysis of the statistical mechanics of breakage; whatever theoretical basis the existing

Jan 1, 1969

By J. R. Stone, J. T. Roy

ASARC09s continuous tapper for lead blast furnace is described. Its use throughout the company's plants has resulted in higher production rates, lower labor costs, and better working conditions. It has also permitted wide latitude in blast pressures, as well as in charge and slag compositions. The practice of tapping a lead blast furnace has remained essentially unchanged for many years. In principle, it is a batch operation discharging the products from the continuous smelting of lead-bearing materials. As such, it imposes certain restrictions on the proper functioning of the smelting operation. First, it calls for a highly developed sense of timing, and a considerable amount of physical effort on the part of the operator. Dirty and hot working conditions, as well as the hazards of flying metal and slag, usually prevail. These conditions have resulted in higher labor costs, both in rate and in amount. Labor turnover has also been high. Secondly, rigid control of mechanical and metallurgical factors must be maintained. Frequent oxygen lancing of the tap hole is needed to permit tapping, thereby increasing operating costs. Rapid increases in smelting rate, plugged tap holes, or forgetfulness on the part of the operator may permit the molten material to rise above the tuyeies and freeze the furnace. Also, the operator will often blow out the tap hole after draining, filling the area with fume. Composition of charge also vitally affects the tapping. The type of lead-bearing material being smelted will naturally affect the smelting rate, and thus the frequency of tapping. The amounts and types of fluxing materials, which determine the ultimate fluidity of the slag, must be regidly controlled. Finally, the batch tapping of a continuous smelting operation of this type appears to be wrong in principle. The molten material drains into the crucible at the bottom of the furnace where it starts cooling immediately and continues to cool until tapped out. While keeping the lead molten is no particular problem, maintaining a fluid slag at all times is often difficult, and sometimes impossible. These factors all point to the desirability of a method for the continuous tapping of a lead blast furnace. The expected advantages would be more desirable working conditions, lower labor costs, increased production, and a much wider latitude in both blast pressure and charge composition. CONTINUOUS TAPPING Historical. A number of devices, e.g., mechanical tappers, siphons, and lead wells, have been tried in the past in attempting to effect continuous tapping. None proved to be entirely satisfactory for tapping lead and slag simultaneously. It was not until 1955, however, that ASARCO designed and put into operation a successful tapping unit at its East Helena, Mont. lead smelter. This unit is described in detail in U.S. Patent NO. 2,890,951. Continuous Tapping Unit—See Figs. 1 and 2. Fig. 1 is a schematic drawing of the unit in place, showing front and side views. The component parts are identified in the caption. Fig. 2 is an actual photograph, front view, of the tapping unit attached to the base of the blast furnace. Operation. The operation of the tapping unit is quite simple. Essentially it consists of a narrow tapping bay surrounding the furnace tap hole, with a deep notch in the front wall. When the blast furnace is started this notch is open down to the level of the tap hole. Starting the furnace may be accomplished in several ways but the most satisfactory method is to plug the tap hole with clay, permitting metal and slag to build up inside the furnace. Before the liquid reaches the tuyeres, the plug is removed and the furnace is drained to the settler. The blast is then stopped momentarily while the notch in the front wall of the tapping bay is dammed with chrome brick and plastic chrome ore. The blast is again started, and the height of the dam is adjusted so that the liquid in the bay will counterbalance the internal pressure of the furnace. When the desired level is reached the slag and metal will overflow to the settler. It should be noted that the molten material ,is not siphoned from the furnace and that the metal 'acts as a liquid valve controlling the flow of slag through the tap hole. During normal operations the tapping bay is filled primarily with lead, covered with a thin layer of slag. Thus, the heavier lead plays the major role in compensating for variations in furnace pressures. Due to the high specific gravity of lead, changes in the liquid level in the tapping bay will never exceed a few inches. For this reason it is seldom necessary to change the height of the dam due to minor variations in operating conditions. Some turbulence of the liquid in the bay is essential at all times. This indicates that the level of the

Jan 1, 1963

Jan 1, 1971

By J. T. Norton, A. L. Mowry

The monocarbides of the A subgroup elements in the fourth and fifth group of the periodic table in addition to being hard and refractory are of special interest in that they are isomorphous in crystalline structure. They are cubic with a sodium chloride type structure in which the metal atoms are essentially close packed in a face-centered cubic arrangement with the carbon atoms placed in the interstices between. Interstitial structures of this close packed type were first investigated systematically by Egg1 and he gave the rule for their formation, stating that the radius ratio of the nonmetal to the metal atom should not exceed the value of 0.59. The carbides of interest are those of titanium and zirconium of the fourth group and vanadium, columbium and tantalum of the fifth group. Table 1 shows the radius ratio using the Goldschmidt radii for 12 coordination for the metal atoms and the diamond radius for the carbon atom. It will be noted that while there is considerable variation in the size of the metal atom, in all cases the ratio is smaller than the limit of 0.59 placed by Hägg. It has been known for some time that these cubic carbides are soluble in one another, at least to some extent or, in other words, the metal atoms can be replaced, one by another without destroying the stability of the structure. Since the stability of these close packed interstitial substances appears to depend more upon geometry than upon the exact chemical nature of the atoms involved, it is of interest to examine the possibilities of replacement in these carbides in some detail. Hume-Rothery2 has pointed out the importance of the difference in size of solute and solvent atom as a factor in limiting the solubility in simple binary solid solutions. Largely on an empirical basis, he states that if the difference in size between solvent and solute atom is more than 14-15 pct of the solvent atom, the range of solubility is very restricted. The atom size was based on the distance of closest approach in the elements involved. While there is some question as to how one should calculate the size of the metal atom in the carbide structures, reference to Table 1 will show that zirconium is the largest and vanadium the smallest of the group and that the difference is about 15 pct. The Ti-Zr difference is about 9 pct and the others are smaller. Thus one would predict that if the size factor controls the solubility, all of the pairs except VC-ZrC would have wide or complete solubility whereas this latter pair is on the border line and might have restricted solubility. The purpose of the present investi- gation was to examine the solubility of the several pairs of carbides by heating them together until equilibrium was established and then examining the product by X rays. Previous Work Agte3 and his associates prepared various transition metal carbides and determined the melting points of binary mixtures. He concluded from the shapes of the melting point curves that there was extensive solubility in the case of the cubic carbides. Umanskii and his colleagues made an investigation of a number of pairs of the cubic carbides, using X rays and plotted lattice parameter vs. composition curves for the systems TaC-Tic, CbC-Tic, TaC-ZrC and CbC-ZrC. All pairs showed a continuous series of solid solutions. The first two pairs gave a linear relation while the latter two showed a negative deviation from Vegard's law. Kiefer and Nowotny, in a paper which became available after the present work was well advanced, investigated the binary pairs of the five cubic carbides by means of X rays. Relatively few points were obtained and results indicated that in some cases, at least, equilibrium was not reached at the temperatures used. The results indicated that solubility in the VC-ZrC system was not complete. All of the results of previous investigations indicated the desirability of a more detailed study. Materials The raw materials used were mono-carbides of titanium, zirconium, vanadium, columbium and tantalum and were the purest which could readily be obtained commercially. Spectrographic qualitative analysis showed that the CbC and TaC contained less than 1 pct

Jan 1, 1950

Executive A R LEDOUX, Chairman ARTHUR S DWIGHT CHARLES F RAND E P MATHEWSON J V W REYNDERS Membership E DEGOLYER, Chairman W Y WESTERVELT, Vice-chairman WILLIAM H BASSETT H G MOULTON C MINOT WELD Finance WALTER H ALDRIDGE, Chairman GEORGE D BARRON GALEN H CLEVENGER Library GEORGE C STONE,' Chairman JOHN H JANEWAY ALEXANDER C HUMPHREYS, ROBERT LINTON FREDERICK F SHARPLESS

Jan 1, 1923

By G. Gabra

The kinetics of acidified thiourea leaching of pyrite concentrate was investigated at temperature ranging form 20 °C to a 60 °C. The effect of various parameters on the dissolution rates are discussed. The gold extraction was found to be zero and first orders with respect to Fe (111) and thiourea concentration, respectively. The order of the reaction with respect to sulfuric acid was found to be zero. The activation energy was determined from the Arrhenius plot as 2,039 kcal mole -1 (for a freshly prepared acidic solution containing Fe (111) as the oxidant and at 800 rpm stirring speed. This relatively low activation energy value suggests that the gold extraction is controlled by a-physical phenomenon. Under optimum conditions, more than 90% of gold has been dissolved in about 30 min. and 95% in 60 min. Attempt to recover gold from thiourea solution using active carbon is also discussed.

Jan 1, 1984

By D. L. Harris

Chemical upgrading testwork on Stillwater chomite concentrate shows that reduction roasting and leaching (sulfuric acid and/or ferric sulfate) can provide upgraded feed to ferrochromium furnaces for making upgraded ferrochromium. Two-stage reduction smelting in electric arc furnaces can make upgraded ferrochromium also. Sulfating roasts and leaching (acid and/or water) results using gaseous SO2—SO3 or solid ammonium sulfate were not satisfactory for upgrading chromite. Chemical upgrading of domestic Stillwater (Montana) chromite has been considered at various times1'2' to provide a material with higher chromium to iron ratio, more like foreign high grade ore, for the production of standard grade ferrochromium. Physical bene-ficiation of Stillwater chromite ore can provide a ratio, Cr/Fe, no higher than the ratio in the chromite mineral, which is approximately 1.7. For comparison the analyses of a foreign metallurgical grade chromite ore, a high grade electrostatic process Stillwater concentrate and a stockpile concentrate from the Stillwater ore are noted in Table I. Single stage smelting of the high grade foreign ore with a ratio, Cr/Fe, of 3.37 can provide a ferrochromium of over 65 pct chromium content. Similar smelting of the Stillwater concentrate with a ratio of 1.7 can provide a ferrochromium of 50 to 55 pct chromium content.' It has been demonstrated on a pilot plant scale that chemically upgraded Stillwater chromite having a ratio, Cr/Fe, of 3.0 also can be smelted to provide a ferrochromium containing over 65 pct chromium.5 Two major considerations for an upgrading process are the ratio, Cr/Fe, in the product and the chromium recovery. A minimum ratio, Cr/Fe, of 3.0 is required in the product but for economic reasons a higher ratio is desirable for blending purposes as long as the recovery is satisfactory. Data from different kinds of chemical upgrading testwork are presented in this report. Some of the testwork was conducted to make a product which was not intended to be feed for ferrochrome smelting, but the data are considered pertinent to chemical upgrading. Results reported by others are noted also for comparison. WORK REPORTED BY OTHERS The U. S. Bureau of Mines has contributed much to the literature on chemical upgrading of chromite and a summary is given in Table 11. Reduction roasting and acid leaching were reported by Boericke and Bangert,1 by Lloyd and Gorst, et a1,2 and by Boericke3 to demonstrate very good success on a laboratory scale and only fair success in pilot plant tests for Stillwater (Benbow) concentrates. This work included basic studies on thermodynamic properties to explain the reduction and leaching results. Canadian workers, Downes and Morgan,9 have reported experimental success in upgrading a low grade Bird River chromite, similar to Stillwater material by the same process. Russian workers1' have noted results on reduction tests of high grade chromite which may apply to low grade ores and which help in understanding the reduction process. The use of flux additives in the reduction process was an innovation of their work. The Canadian workers9 also reported other kinds of experiments, including autoclave leaching of the chromite in a limited amount of sulfuric acid. An insoluble basic chromium sulfate was formed and separated from soluble iron, magnesium, and aluminum sulfates to provide an upgraded material. U.S. Bureau of Mines workers, Lloyd, Rosenbzum, Home and Davis,' reported on autoclave leaching of Benbow concentrates as a means of obtaining chromium alum for electrolysis. This last work required selective crystallization of alum to separate iron from chromium. Chloridizing roasting of Benbow and Mouat concentrates has been reported by Hunter and Banning6 and others8 to show that the iron chloride could be volatilized from the chromite to provide upgraded residues.

Jan 1, 1964

By E. H. Timmerman, A. F. van Everdingen, J. J. McMahon

The prevent paper contains a method which combines the material balance equation' with the water influx equation' to obtain reliable values for the active oil originally in place and a quantitative evaluation of the cumulative water influx. The method is illustrated by an application to a reservoir without original gas cap. In the absence of an original gas cap. results may be obtained using only field production and pressure data, PVT analyses. and a minimum of subsurface information. CHARACTERISTICS OF RESERVOIR AND RESERVOIR FLUIDS Production in the field under review is obtained from the top of the Wilcox formation of Eocene age, at a depth of approximately 8,100 ft subsea. The structural map, Fig. 1, shows that the accumulation, half elliptical in shape with the long axis in east-west direction, is trapped to the north and on the upthrown side of a normal fault. A number of east-west faults with additional minor faulting in random directions are found in the general area, and seismic information and production performance indicate in this particular case the existence of a second fault a short distance downdip, with a strike more or less parallel to the fault controlling the accumulation. It is estimated that 1,830 acres were originally underlain by oil and that the maximum thickness of the original oil column, about 37 ft, was only slightly greater than the maximum thickness of the oil-bearing sand. Maximum net sand thickness was estimated at approximately 26 ft. From logging and subsurface information the gross sand volume has been placed at 37.400 acre-ft. which reduces to about 27,500 acre-ft after probable non-productive intervals have been deducted. The following table presents laboratory analyse; data for the Wilcox sand cores. Average Average Samples Considered Permeability Porosity All Samples 236 md ..19.9percent Samples with k>l md 256 md 20.5 per cent Samples with k>8 md 275 md 20.9 per cent Interstitial water was estimated at 15 per cent. The initial gas-oil ratio was approximately 900 cu ft/bbl. and the initial shrinkage factor was determined by the laboratory to

Jan 1, 1953

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Jan 1, 1938

By J. W. Downey, M. V. Nevitt

DURING the course of the present investigation of alloy phases involving transition metals, phases were found in the systems Os-Ta, Os-W, and Ir-Ta. The (Os, W) phase was also recently reported by Raub.1 Pertinent information is given in Table 1. The approximate composition range of sta- bility was determined metallographically on arc-melted buttons that were annealed 72 hr at 1200°C and water quenched. X-ray powder patterns were obtained with a 114.6 mm diam powder camera using filtered copper radiation. The composition for (Os, W) given by Raub (66 atomic pct W) is in agreement with the composition range reported here. However, the lattice constants reported by Raub are: a0= 9.913±0.007, c/a= 0.515. The reason for the discrepancy between the lattice constants reported by Raub and those given here is not known. Acknowledgments The authors wish to thank M. H. Mueller and H. W. Knott for assistance with the X-ray diffraction work, and R. A. Morris for the preparation of the alloys. This paper is based upon work performed at the Argonne National Laboratory under the auspices of the Atomic Energy Commission.

Jan 1, 1958

By Robert A. Haney, Stephen J. Gigliotti, James L. Banfield

A series of investigations were made by the Mine Safety and Health Administration to evaluate the performance of various face ventilation systems as part of a program to reduce frictional ignitions of methane and to develop face ventilation systems that could be used with the line brattice more than 10 feet from the face. Tests were conducted in a full-scale laboratory model of a face area. Systems tested included unassisted blowing and exhaust line brattice, diffuser fan and spray fan systems. Systems were evaluated at air quantities ranging from 1.4 to 7.0 m3/s (3,000 to 15,000 cfm), and brattice-to-face distances from 1.5 to 9 meters (5 to 30 feet.). Tests were designed so that systems could be evaluated for the location of the high-point methane concentration in the face area, methane dilution capacity, and face ventilation index. The methane dilution capacity is the total amount of methane a system can dilute to 1.0 volume-percent. The face ventilation index is the efficiency at which the available face air volume is being used.

Jan 1, 1982