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By John H. Walthall, Raymond L. Copson, Travis P. Hignett

In October 1942, the War Production Board requested the Tennessee Valley Authority to undertake investigations to determine the feasibility of producing alumina suitable for reduction in aluminum cells, from clays and other domestic materials, by a lime-sintering, soda-ash leaching process. The work was undertaken by the chemical engineering staff of the Authority under the general direction of the War Metallurgy Committee of the National Academy of Sciences. The investigation included small-scale tests to determine the most suitable conditions for the process, and construction and operation of a pilot plant to obtain data for evaluation of the process and for design of a plant. The results of this investigation are described in the present paper. Sources of Alumina Alumina for the commercial production of aluminum is obtained from bauxite by the Bayer process. The tremendous demand for aluminum for wartime uses is causing rapid exhaustion of the high-grade bauxite deposits in the United States. On the other hand, there is an abundance of aluminous materials such as clay and lower grades of bauxite, which are not suitable for treatment by the Bayer process because of their high silica content. The Tennessee Valley Authority has conducted research and pilot-plant work for more than 5 years on methods for the production of alumina from clay and low-grade bauxite. Particular attention has been given to the high-grade white kaolin found in Carroll County, Tennessee. Exploratory drilling has demonstrated that at least 4 1/2 million tons of kaolin containing in excess of 34 per cent AlzOI is available in this area by strip-mining methods. It has been estimated that hundreds of millions of tons of clay of similar quality occur in the Tennessee Valley region. The Process The process is illustrated in the flowsheet (Fig. I), and consists essentially of the following steps: I. Heating a mixture of clay and limestone to form a sinter containing calcium aluminate and dicalcium silicate. z. Leaching this sinter with a dilute sodium carbonate solution, to form a sodium aluminate solution and a calcium carbonate, calcium silicate residue, which is discarded. 3. Treating the sodium aluminate solution with carbon dioxide to precipitate aluminum trihydrate and regenerate the sodium carbonate solution, which is recycled. 4. Heating the aluminum trihydrate to produce alumina. The process was considered promising because of the abundance and cheapness of the two raw materials, clay and limestone; and because, as compared with an acid process, smaller amounts of critical

Jan 1, 1944

By James R. Holden, Frederick E. Wang

Through both single-crystal and powder X-ray diffraction methods, ten A6B23-type compounds have been confirmed to exist between lanthanides (A) (plus scandium and yttrium) and manganese (B); A = Y, Nd, Sm, Gd, Tb, Dy, Ho, Er, Tm, and Lu. The formation of a compound of this type is shown to he extremely atomic size-sensitive; hence it can be classified as a "size-factorH compound. The "enveloping effect", a geometrical consideration observed in its crystal structure, is proposed as the reason for the A6B23-type compound being size-sensitive. The approximate ideal geometrical ratio of the radii R/r is 1.31 while experimentally A6B23-type compounds have a radius ratio lying within the range 1.2 to 1.4. FLORIO t al.' characterized the structure of Th6MnZ3 as fcc, space group Fm3m, with 116 atoms in the unit cell. Since then, a number of isotypic binary compounds, and recently Gd n,,' have been confirmed to exist. The fact that strontium and barium form A6Bz3-type compounds with magnesium strongly suggested the possible existence of Ba6Liz3. However, investigation3 showed the compound Ba6LiZ3 to be absent. Since both strontium and barium are group 11-a elements and are therefore "open metals",6 the nonexistence of Ba6LiZ3 can hardly be explained satisfactorily by valence-electron considerations. On the other hand, the consistent atomic-radius ratio, (R/r),* observed for the known A6Bz3-type compounds,3 strongly suggests that the formation of compounds of this type is atomic size-sensitive. Therefore, one is tempted to explain the nonexistence of Ba6LiZ3 entirely on the basis of the atomic-size difference between strontium and barium. However, this approach is not entirely without objection. Atoms are not rigid spheres and are known to vary in size within certain limits.7 Since the atomic-radius difference between strontium and barium (0.07 to 0.09A) is within these limits, it is reasonable to assume that the size difference would have a negligible effect on the formation of Ba6LiZ3. This view is further supported by the fact that the radius ratio, R/r, in other known "size-factor" compounds is observed to range widely—for example, from 1.08 to 1.45 for ABz-type compounds (C15, MgCuz type)' and from 1.37 to 1.58 for AB5-type compounds (D2d, CaZn5 type).g The present investigation was undertaken in order to find a more satisfactory explanation for the non-existence of Ba6LiZ3 and, consequently, a better understanding of the nature of the A6Bz3-type compound. The primary objectives are to confirm the previous conclusion3 that the A6B23-type compound is indeed a "size-factor" compound and subsequently to determine the atomic-radius ratio range in which the A6Bz3-type compound can exist. In order to achieve these objectives, stoichiometric A6Bz3 alloys, where manganese (B) was alloyed with various lanthanide elements (A), were selected for investigation. The atomic-radius ratios of lanthanide elements with manganese range from 1.26 for Lu/~n to 1.46 for Eu/Mn. This radius ratio range includes and exceeds the range of all previously reported A6Bz3-type compounds—1.32 for Th/Mn' through 1.38 for Sr/Li. Furthermore, the atomic-size difference between successive elements of the lanthanide series in order of atomic number) is of the order of 0.01A (europium and ytterbium are exceptions). The series of lanthanon-manganese alloy systems is ideally suited to a precise determination of the limits of allowable atomic-radius ratio for A6Bz3-type compound formation. EXPERIMENTAL PROCEDURE The lanthanide metals, in ingot form, supplied by Michigan Chemical Corp. (St. Louis, Mich.) and Nuclear Corp. of America (Burbank, Calif.), were guaranteed by the suppliers to be at least 99.9 wt pct pure (traces of silicon, calcium, and other minor constituents present on occasion, not to be more than 0.05 wt pct) as shown by spectrographic analysis. Manganese metal, in polycrystalline form, was redistilled from the commercial, chemically pure grade and was analyzed to be at least 99.95 wt pct pure. In all cases, the atom ratio between the two elements in each charge was A (rare-earth meta1):B (manganese) = 6:23 and a constant weight, 3 g, of

Jan 1, 1965

By Roscoe C. Clark

A study of the effect of various corrosive waters on five different types of cements indicated that those cements containing less than 5 per cent tricalcium aluminate were the most resistant to corrosion. INTRODUCTION Oil field experience indicates that some formation waters are excessively corrosive to some cements. A study has been made to determine the most corrosive waters and the relative resistance of various types of cement to these corrosive waters. PROCEDURE This study of the corrosion of cement was limited to an evaluation of the effect of various corrosive waters on five different types of cements. A determination was made of the types of cements which offer the greatest resistance to corrosion and of the chemical contents of waters found to be the most destructive to cements. The laboratory tests were conducted by immersing small cylinders of 20 samples of set cement in various corrosive waters. The initial mixing of most of the cement slurries was done using Tulsa tap water as the mixing water; however, a few of the slurries were mixed using a sulfate water as the mixing water. When the 20 samples had been immersed for a period of nine months, they were removed from their corrosive baths, inspected and photographed. The field tests were conducted by placing ten cement specimens, each contained in a screen wire thimble, in the gas anchors of insert pumps in wells in six different fields. The specimens were observed whenever it was necessary to pull the pumps and were returned to the 1aboratory when deterioration had reached a serious state. EXPERIMENTAL RESULTS In both the field and laboratory tests the resistance of the different cements to corrosion was determined by the condition of the samples and the number of samples of each that were in good condition at the end of the immersion period. Table I lists the various cements tested in the order of their resistance to corrosion with the most resistant at the top and the least resistance at the bottom. An analysis of Table I reveals that the upper half of the table predominates in slow set and sulphate resisting cement.: (low 3CAO.A12O3 content) while in the lower half the port-land and high early strength cements are in the majority. Of the last two the high early strength cements appear to offer less resistance to sulfate water corrosion. An investigation of the composition of the cements listed in Table I indicated that as a general rule the cements containing less than five per cent tricalcium aluminate will exhibit the greatest resistance to corrosion. The destructive action of various corrosive waters was compared by listing the waters in the order of those containing the largest number of samples that failed completely. These ratings are presented in Table II with the most corrosive waters listed first and the least corrosive last. Figs. 1 and 2 show the end views of cement samples which had been immersed in the baths of five per cent MgSO, water and the five per cent Na2SO4 water respectively in the laboratory. These .pictures are representative of the results obtained and present a pictorial evaluation of the corrosive resistance of various types of cements.

Jan 1, 1950

By T. A. Wright

The sampling of waste materials containing copper, lead and tin has taken on a new significance within recent years, and is of increasing importance, on account of the entry of some of the copper refineries into this field and the tendency towards consolidations among the smelters. In the past, such material was sold on the basis either of a rough or grab sample or of casual inspection of the lot. The two principal sources of this waste are the processes of fabrication and the scrapping of automobile, electrical, plumbing, railroad and other equipment. In both cases much material can be sorted, salvaged and sold by classification and grade to be melted "as is" or with the addition of new or virgin metal. The increasing care in classifying exercised in the large manufacturing plants tends to increase the proportion of this usable scrap; on the contrary, the increasing use of more complex and special non-ferrous alloys tends to make sorting and grading more difficult in turnings, borings, clippings, etc., and consequently is making sampling of more moment. In tin, lead, aluminum, and zinc furnace by-products such as ashes, drosses, skimmings, etc., the trade practice of evaluation has not varied much. For instance, tin drosses, battery lead, and aluminum skimmings are ordinarily sold on what is called "metallic button" or "metallic recovery" or ('metallic yield." Copper and lead waste, however, is being sold more and more on actual copper and lead contents based on sample and chemical analyses. This paper will be devoted chiefly to the sampling of materials suitable only for refining and recovery of one or more of the metals present. The methods described are often based on the fact that but one metal, such as copper, is to be paid for. The loss of volatile constituents in melting does not appreciably affect the final copper result, as the "after-melting" assay can be readily calculated to the "before-melting" basis, but if smelter efficiency continues to increase, so that in time zinc, for instance, may be recovered on a paying basis, the method of sampling dirty brass, for example, will have to be changed also. As a matter of fact, we may expect that improved refining processes will make possible the partial recovery of lead and tin which are now lost.

Jan 1, 1928

By V. Balasubarmanian, Rajendra Kumar

ISOTHERMAL formation of martensite in a copper-aluminum-nickel alloy was previously reported by Hull and Garwood.' In the present work an attempt has been made to investigate some of the character- istics of the formation of isothermal martensite in binary copper-aluminum alloys containing 14.00 and 14.50 pct Al, respectively. Chill-cast alloys were prepared from OFHC copper and aluminum of greater than 99.5 pct purity. They were homogenized for 24 hr at 950° C. Polished samples were soaked at 1000°C and quenched in water to retain the p phase untransformed; martensite did not form during the quench. The surface was then repolished for isothermal microscopic studies at room temperature. The salient features of the martensitic transformation in copper-aluminum alloys observed in this investigation are: a) The transformation progresses (i) by nuclea-tion at new sites and (ii) by simultaneous slow isothermal growth of existing martensite plates. Slow isothermal growth of martensite has previously been observed by Holden2 in an uranium-chromium alloy and by Hull and Garwood in copper-aluminum-nickel alloy. b) Martensite plates nucleate preferentially at the pin holes formed during the casting of the specimen, Fig. 1. c) Some of the martensite plates pass right through other plates lying in their path, Fig. 1. d) Martensite plates cross grain boundaries without any change in the direction of growth. e) An irregular network of markings, Fig. 2, appears simultaneously with the formation of martensite. It was less prominent in the case of 14.5 pct alloy. The markings may arise due to spontaneous deformation under the lattice strains of the martensitic transformation. Crossing of martensite plates was first observed by Kulin and cohen3 when a high-nickel steel was cooled to 77°K. They suggested that the rapidly advancing plates atquire considerable kinetic energy and strike the obstructing plate with a momentum so as to nucleate the process on the other side of the plate. This, however, fails to explain the present observation where the slow isothermally growing martensite plates cross each other. Such plates would not possess large momentum to bodily displace the obstructing plate. Fig. 1 shows the martensite plates crossing each other without body displacements. This should be contrasted with Fig. 3 of Kulin and Cohen's paper where crossing of plates is accompanied by shearing or body displacements of the plates. The present observations can, how-

Jan 1, 1961

By Jesse O. Betterton

In the Parkes process of lead refining, after desilverization has been completed by means of zinc additions, there will remain in the lead from 0.5 to 0.6 per cent zinc. At this stage in the refining practically all impurities are reduced to refined lead specifications except zinc and antimony. All excess of copper and nearly all the arsenic goes out with the silver. Silver remaining should not exceed 0.1 oz. per ton. It thus remains only to remove zinc and antimony to produce refined lead. The older practice was to complete refining by means of oxidation processes in reverberatory furnaces. Essentially this consisted of heating the lead to 1200' to 1300" I?. and allowing it to come in contact with a considerable excess of air passing through the furnace, or accomplish the same result by blowing the lead with steam by means of pipes introduced for that purpose, at the same time allowing air to enter the furnace. Usually some procedure involving alternate oxidation and reduction with fine coal or coke was employed in order to reduce the weight of dross or skim finally removed from the furnace. This practice had many elements of weakness in that the main flour of lead, after the preceding expensive operations had been performed, and the whole refining process nearly completed, was subjected to an operation at a comparatively high temperature, which involved much lead loss if the furnaces were not on bags, and the production of 3 to 5 per cent of the weight of the lead as dross or skim, which had to be resmelted and the lead either refined or produced as antimonial lead. In addition, the operating expense of the process was a considerable item. The process of dezincing with chlorine was developed as a substitute for the foregoing described oxidation process, and finally resulted in such advantages as nearly equal operating costs, a practical elimination of the lead loss, and the direct production of a very valuable byproduct, zinc chloride, that may be sold as such or retreated for its zinc and chlorine contents, both of which can be reused in the refining processes. Essentially this process consists of circulating the lead by means of a pump in a kettle through a closed reaction cylinder, gaseous chlorine entering the reaction cylinder with the stream of lead from the pump. Molten zinc chloride is thus produced and is discharged together with

Jan 1, 1937

By Will Mitchell, T. G. Kirkland, C. L. Sollenberger

R. S. Handy (Santa Rosa, Calif.)—It would be interesting to learn the comparative results of treating the Korean scheelite ore described in this paper according to the following procedure: I—Follow the procedure described in the paper to the recovery of the sulphides by flotation. 2—Dilute the residue from the sulphide flotation strongly with water and disperse the contained colloids by adding a mixture of 75 pct sodium silicate and 25 pct quebracho to dispersion point. 3— Settle the dispersed pulp and decant the supernatant colloids. 4—-Dilute, settle, and decant the pulp until the settled residue is crystalline. Not over 10 pct of the tungsten minerals should follow the colloids. 5—Treat the settled residue by flotation, having established the proper pH. Oleic acid homogenized with cresylic acid or kerosene is more selective for scheelite than is oleic acid alone. (The recovery of the contained tungsten minerals should be practically complete in this operation.) 6—Deflocculate the tungsten concentrate from step 5 with the same dispersion agents as in step 2 and treat the deflocculated pulp by gravity. The minerals that tend to float with scheelite are generally of low specific gravity and the separation is efficient. The heavier contaminants, such as pyrrhotite and magnetite can be removed magnetically, leaving a practically pure scheelite. E. J. Pryor (Imperial College of Science and Technology, London)—It occurred to me that instead of using temperature as a differentiating control, more precision might be obtained with a reagent of a type we developed here for the flotation of high grade fluorspar in cold pulps. This reagent, which is manufactured on a modest scale in England, is based on oleic acid, which has been modified by our patented process to ensure complete stable dispersion at any desired concentration in water. It thus becomes possible to utilise unimolecular layers with good selectivity. In the fluorspar flotation for which we developed it, it is possible to hold a tight control on depressing quebracho for the calcite without substantial loss of flu-orite, and recovery exceeding 99 pct is not uncommon in the mill where it is in use, even under wintry conditions. No attempt has been made yet to extend the use of the reagent, as we have not completely overcome

Jan 1, 1952

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947

By T. H. Alden

In several recent studies of fatigue fracture in single-phase metals, tests have been made using the condition of essentially constant plastic strain amplitude.1-4 This procedure departs from the more common practice of holding the amplitude of stress constant. From these studies have emerged two important generalizations about the behavior of metals during cyclic straining. First, measurements of the stress amplitude, required to impose the fixed plastic strain amplitude, have shown that the rate of cyclic strain hardening in annealed metals is rapid in the early cycles but decays quickly to a very low value.1"5 This behavior is displayed particularly for a range of strain amplitudes below about 2 pct and at low and high temperature. Frequently, the hardening rate becomes so low that one can determine a "saturation" stress. This stress increases with strain amplitude. In addition, an experimental relation between fracture life and plastic strain amplitude was obtained by Coffin and Taternelli,' namely N 1/2 ep = constant. For several metal:;, this equation has been shown to hold down to a strain amplitude of 0.1 pct. Whether it also holds in the "low amplitude" range (lives of more than about 105 cycles) has not been established. The importance of the saturation stress level (a value derived from strain hardening measurements) to the question of fracture life (the usual quantity of interest of fatigue testing) lies in the observation that cracks start to form when the cyclic hardening rate is essentially zero, and not before.4"5 This point suggests immediately a means of relating tests at constant stress and constant strain amplitudes. When making comparison, the stress in a constant stress test is to be identified with the saturation stress in a constant strain test. If these ideas are correct, then it should be possible, using the relation between fracture life and strain amplitude and experimental curves of saturation stress versus strain amplitude, to predict the fracture life as a function of stress, i.e. predict the S -N curve, for low stress fatigue. Saturation stress amplitude-plastic strain amplitude data, obtained in tension-compression, exist only for a few metals and for relatively high amplitudes of strain. Comparable S -N curves (push-pull loading) were known by the author only for OFHC copper. Values for this metal taken from Ref. 2 are shown in Fig. 1. The strain amplitudes, the lowest

Jan 1, 1962

By G. P. Walters, W. C. Fuller

THE production of thin films of beryllium became essential early in the irradiation program at Harwell in order to study the fundamentals of irradiation damage in this metal. In common with other metals, electron-microacope specimens of beryllium can be produced by different techniques. The techniques used to produce thin-film specimens varied according to the starting material. Most of the starting material was -0.060 in. thick. Large-area specimens (1 in. long by 0.5 in. wide) were chemically thinned to -0.005 in. thick in a solution containing 85 pct orthophosphoric acid, 10 pct chromium trioxide, and 5 pct sulfuric acid at 70°C prior to electropolishing. Small-area specimens, Fig. 1, had to be mechanically thinned prior to electropolishing because chemical thinning considerably reduces the area as well as the thickness. Mechanical thinning tends to produce a worked layer about 0.002 in. deep and hence this method cannot be used to thin bulk specimens below about 0.005 in. thick. Mechanical thinning is carried out by mounting the specimen on perspex then carefully polishing on wet grinding papers. Polishing is continued until about half the initial thickness is removed; the reverse side is then polished in a similar manner. After the bulk specimens have been mechanically or chemically thinned to -0.005 in. they are elec-tropolished in the solution shown in able I at 50" to 80°C. This solution should be prepared by adding the chromium trioxide to the orthophosphoric acid and then the glycerol last. The glycerol should be added slowly because a vigorous effervescence can occur. Electropolishing is carried out by holding the specimen in stainless-steel tweezers and gently agitating in the solution. Masking or "window" techniques are not required for beryllium, good microscope specimens being obtained from the edges of elec-tropolished foils. A stainless-steel beaker should be used, acting as the cathode and solution container. A closed circuit voltage of 30 v is used for polishing. Too low a voltage produces a matt surface and pitting occurs when it is too high. The solution temperature is important. If it is too low, the solution is viscous and bad polishing results and, if too high, etching of the specimen occurs. After electropolishing, the foils are thoroughly

Jan 1, 1963

By B. E. Lindsly

The withdrawal of gas and/or oil in appreciable quantity from a natural oil reservoir causes the pressure within the reservoir to diminish, and if the oil is completely saturated with gas, this dissolved gas begins to evolve from solution. The liberation of gas from solution in oil causes the oil to change in several ways, the principal changes being: (1) reduction in volume or "shrinkage"; (2) increase in specific gravity (decrease in degrees A.P.I.); (3) decrease of energy in the oil; (4) increased viscosity; (5) increased surface tension. As the writer's investigations have been concerned more with the first three items than with changes in viscosity and surface tension the following discussion relates particularly to the volumetric shrinkage of the oil, to the increased specific gravity, and to the decreased energy in the oil, each caused by the liberalion of dissolved gas. One of the most surprising facts that have developed in the study of high-pressure well-head and "bottom-hole" samples is the great reduction in volume that occurs in some oils when the naturally dissolved gas is liberated. High-pressure samples taken from well heads in the Oklahoma City, Kettleman Hills and Ventura, Calif., fields decreased in volume from 11.2 to 40.5 per cent when the dissolved gas was liberated in reducing the pressure on the sample to atmospheric. The amount of shrinkage an oil undergoes when gas is liberated depends upon the quantity of gas liberated and also upon its composition. A given volume of the heavier gases, propane and butane for example, will occupy more space in the oil than an equal volume of methane. The reason for this probably is that the molecules of the heavier gases are larger than those of the lighter gases. When gases are not in solution, and particularly when they are at low pressures, the distance between the gas molecules is so much greater than the dimensions of the molecules themselves that the mere size of the molecules has little effect upon the space occupied by the gas, Avogadro's hypothesis, which in effect states that all gases under the same conditions have the same number of molecules per unit volume, relates to this phenomenon. The fact that the heavier gases occupy more space than the lighter gases in the oil, and

Jan 1, 1934

By William Hutton Blauvelt

The type of gas producer in which the ashes are fluxed and run off as slag was among the very earliest made. Ebelmen built the first one in 1840 at Audincourt, France, only a year after the installation of the first gas producer of which we have record. Charcoal was used as the fuel, with blast-furnace slag and clay as flux. My interest in this type of producer began when as a boy I saw the fluxing producer at Chester, N. J., which was invented by W. J. Taylor, and described by him in 1881.' This producer was invented independently by Mr. Taylor to meet the difficulties he experienced in making producer gas for roasting sulphurous iron ores. It will be of interest to review briefly his description of his producer. It was built like a small blast furnace, having a hearth 24 in. in diameter and 24 in. high. The bosh angle was 25' from the vertical; the diameter at the bosh 4 ft. and at the top 3 ft.; the total height 12 ft. The producer contained one water-cooled tuyere 12 in. above the bottom, with a 1.5-in. nozzle. Depth of the coal above the tuyere, 6 ft. Mr. Taylor mixed the coal with from 30 to 40 per cent. of basic blast-furnace slag to flux the ash. Occasionally some limestone was also used, but never limestone alone, as the use of the furnace cinder gave a larger volume of slag and made it easier to maintain a proper fluidity. The producer was blown with a small Weimer blowing engine, delivering 300 ft. of air per minute, at from 1 to 1.5 lb. pressure. The slag was tapped every 2 hr. and was black and glassy in appearance. Broken or egg size anthracite was used. The fine sizes of anthracite could not be made to work. Mr. Taylor expressed the belief that bituminous coal could be used if not too fine, but no experiments were made with this coal. Runs were made of four weeks' duration, with no stops or changes. Mr. Taylor reported the following advantages: 1. Excellent gas, very uniform in quality. The high fuel bed permitted no air to pass unconsumed, and the gas was almost entirely free from carbonic acid.

Jan 1, 1914

By Paul Gordon, A. S. Iyer

Apreliminary investigation has been carried out on the annealing behavior of explosively deformed copper as compared to that of conventionally deformed copper, using hardness, stored energy, and the sharpening of X-ray spots on Debye rings as the measures of annealing. Some metallographic observations were also made. Two purities of copper were studied. The first was tough pitch copper obtained as a small piece from the guard ring of a sample explosively deformed to approximately 500 kbars.* The second consisted of mentedand deformed these samples quite badly so that only a few small fragments were available for the annealing experiments. As a result, little in the way of quantitatively significant data was gleaned from the measurements. However, several important qualitative differences between the annealing of explosively deformed and conventionally deformed copper have been observed. It is these effects which are described in this report. Fig. 1 presents the results of hardness and stored energy measurements during the annealing of the high-purity copper after explosive and after conventional deformation. The data for the conventionally deformed copper were taken from the work of Gordon;' in this copper it is evident that most of the stored energy is released concommittantly with the hardness drop. As shown in orddon's' investigation, these effect:: are definitely associated with recrystal-lization. In the case of the explosively deformed copper, on the other hand, a large fraction of the stored energy is released before any appreciable hardness change takes place. At 160°C about 40 pct, and at 170°C about. 80 pct, of the stored energy appears to have been released prior to the onset of softening, i.e., undoubtedly prior to recrystallization. In view of the facts that the hardness and energy measurements were made on separate samples and that there is reason to suspect that the intensity of the explosive deformation from sample to sample may not have been uniform, a check run was carried out in which both energq release and hardness changes were measured cnn the same sample. The results of this run, shown in Fig. 2, confirm the conclusion that much of the stored energy is released before the hardness begins to change appreciably. In an effort to obtain some idea of the activation energy associated with the early release of energy, a single callorimetric sample was annealed in two steps, first at 149.5"C and then at 159.5"C, Fig. 3.

Jan 1, 1962

By R. J. Fruehan

The oxygen activity and concentration were measured in Fe-Cr-0 melts in equilibrium with an oxide phase at 1600°C (2912°F). The activity was determined by ,use of the following solid oxide -electrolyte galvanic cell CY-Cr8,(s) I ZrOz(CaO) I Fe-CY-G(saturated)(l) The oxygen concentration decreases with increasing Cr concentration to about 270 ppm 0 at about 7pct CY and then increases gradually. The activity coefficient of oxygen (fo) decreases with increasing Cr. In melts containing up to about 20 pct Cr, log f is approximately a linear function of wt pct Cr with a slope (e q 2) of —0.037. The activity of chromium was calculated and found to exhibit a small negative deviation from Raoult's law. From the activity and solubility data for low chromium melts, the free energy of formation of chromite, FeCr204, was found to be -79.8kcal per mole where pure liquid chromium and oxygen at I wt pct in Fe are the standard states. ThE effect of chromium on the chemical behavior of dissolved oxygen in liquid iron is of great importance in controlling the deoxidation of steels containing a significant amount of chromium. Chen and chipman' equilibrated Fe-Cr melts in the presence and absence of slag with hydrogen-water vapor mixtures. They concluded that at 1595°C chromite was the oxide phase in equilibrium with Fe-Cr alloys containing less than 5.5 pct Cr while at higher chromium concentration Cr,O, was the stable phase. In the composition range 0 to 10 pct Cr they found that the interaction coefficient, was equal to -0.041. Turk-dogan,' Schenck and Steinmetz,, and pargeter4 measured egr) in a similar manner and found the value to be -0.064,-0.04, and -0.052, respectively. McLean and Be11 evaluated egr) from their data on the equilibrium of Fe-Cr-Al-0 alloys with H2/H20 mixtures and found it to be -0.058. However, McLean and Bell's value should only be considered an estimate because the effect of aluminum on the activity coefficient of oxygen is about a hundred times greater than that of chromium. Consequently, an error in the value of egl) used, which at the present time is not well-known, or an error in aluminum analysis, which is present in very small quantities, will result in a significant error in egr). Fischer et a1.6 determined the interaction coefficient (eEr) in Fe-Cr-0 melts not in equilibrium with an oxide phase and containing less than 18 wt pct Cr at 1600°C electrochemically. They determined a value of -0.031 for egr). Hilty et aL7 measured the oxygen content of Fe-Cr melts in equilibrium with an oxide phase containing up to 50 pct Cr. They found that the solubility of oxygen decreased as the chromium content increased to about 6 pct Cr and then increased gradually. They concluded that the equilibrium oxide phase was chromite below 3 pct Cr, distorted spinel from 3 to 9 pct Cr, and Cr,04 above 9 pct Cr. Adachi and lwamotoa also investigated this system, but did not find Cr30,. They X-rayed the equilibrium oxide phases and did not find the presence of Cr,O,. They also X-rayed the oxide phase extracted from a 65 pct Cr melt which was heat treated and did not find metallic chromium as would be expected if Cr3O4 were the equilibrium oxide phase as indicated by the reaction : 3Cr3O4 — 4Cr2O:, + Cr [lj It was the purpose of the present investigation to determine the effect of chromium on the activity coefficient of oxygen in Fe-Cr melts by measuring the activity and solubility of oxygen equilibrated with an oxide phase in the composition range 0.18 to 50.5 wt pct Cr at 1600°C (2912°F). The activity of oxygen in the melts was determined by use of the following galvanic cell: The relationship between the partial pressure of oxygen in equilibrium with the melt and the reversible electromotive force of the cell (E) is where 11 = 4, F is the Faraday constant, pb, is the oxygen pressure in equilibrium with the meit and is the oxygen pressure in equilibrium with Cr203 as determined from the free energy data compiled by Elliott et al? The oxide phase in equilibrium with pure chromium was assumed to be Cr If Cr30, were the equilibrium phase the activities derived would be approximately the same, since the best estimated free energy of formation of Cr,O,, if it does exist, is approximately % the free energy of formation of The activity of chromium in Fe-Cr alloys at 1600° C was also determined from the measured electromotive force. The activity of chromium (aCr) is related to the electromotive force as follows: , The oxide phase in equilibrium with pure chromium and Fe-Cr melts from 10 to 52 pct is assumed to be Cr203 so that n equals three. If future work proves the existence of Crs04 in equilibrium with Fe-Cr melts and pure chromium, the experimental results can be reevaluated using a value of $ for n in Eq. 141. A value of ^ for n will make the activities about 10 pct higher. In order for Eqs. 131 and [4J to be valid the electrolyte, ZrOa(CaO), must exhibit predominantly ionic conduction at the temperature and oxygen partial pressure of its use. Previous work1' has demonstrated that ZrOz(Ca0) is predonlinantly an ionic conductor

Jan 1, 1970

By E. P. Klier

The work of Klier and Lyman1 on the bainite reaction has led to the full description of this reaction for medium-carbon low-alloy steels. Certain experimental data reported by Klier and Lyman appear, however, to be improbable of obtaining in very high-carbon steels. The study herein reported consists of an analysis of the transformation characteristics of a 3 per cent chromium, I per cent carbon steel by means of X-ray and metallographic methods. From the results obtained it is concluded that the bainite reaction in this steel is more complicated than that reported by Klier and Lyman for the low-carbon low-alloy steels. Henever, the mechanism of bainite formation in this steel appears to be essentially the same as proposed for the low-carbon low-alloy steels with the modification that the first volumes to transform are carbon-rich rather than carbon-poor. Experimental Procedure The analysis of the steel studied was: Cr, 2.89; C, 1.02; Mn, 0.33; Si, o.35; S, 0.012; P, 0.020. The stock was received from the Carpenter Steel Co. in rolled bars. Specimen Preparation Specimens were austenitized at 1200°C. (2192°F.) for 25 min. in purified nitrogen. The resultant A.S.T.M. austenite grain size was 0-1. Isothermal transformation was allowed by quenching the specimens into liquid metal baths controlled to 5°C. at 540°C. (1004°F.) and above, and to 2°C. at 490°C. (914°f.) and below. Following the desired degree of transformation all specimens were water quenched. Metallographic Examination Specimens 1/4 in. round by 1/4 in. were used for nietallographic examination. Etch-ants used were Vilella's reagent (I gram picric acid, 10 ml. hydrochloric acid to make 100 ml. ethyl alcohol solution) and alkalinc sodium picratc (electrolytic). Vilella's reagent was effective in revealing all microconstitucnts, while the alkaline sodium picrate developed the carbon-rich structures or carbides only. X-ray Examination Two X-ray techniques were used—the Debye-Scherrer method and a back-reflection method. 'The small specimens necessary for the Debye-Scherrer method were austenitized and then allowed to transform, or were cut from the larger metallographic specimens. The metallographic specimens were used for the back-reflection X-ray analysis. When it was desirable to determine the structure of the carbide present, sliver specimens were electrolytically etched to develop a carbide coating, then were exposed in the cameras. This technique was frequently necessary to allow positive identification of the carbide.

Jan 1, 1945

By A. A. Collins

OF widespread interest to all lead metal¬lurgists is a dross smelting process that will consistently give mattes and speiss of low lead and high copper contents. It is -a problem that has intrigued operators through the years. In the past few years, with increased amounts of copper and overcrowded metallurgical furnaces, it has been almost imperative to evolve some method. Experience at El Paso has shown that high copper-to-lead ratios can be consistently attained by use of the new soda process.* OLDER METHODS Three older methods; i.e., I. Dross plus siliceous ore and limestone, 2. Dross plus one fourth blast-furnace speiss and siliceous ore, 3. Dross plus sulphides and siliceous ore and scrap iron, do not give consistent results and the products are higher in lead and lower in copper than those produced with the soda process. Any radical changes in the blast-furnace drosses will throw any of these older methods into a tailspin. For almost ten years, El Paso used the third method, with varying results. It was noted that as the arsenic and antimony contents of the dross increased, the speiss and matte steadily increased in lead and decreased in copper. A dross containing four parts of copper to one part of arsenic was considered to be ideal. At times this balance of copper to arsenic was almost impossible to obtain, especially when handling copper Cottrell dusts, which vary so greatly in copper and arsenic. Custom smelters like El Paso, handling a great variety of dusts, concentrates and ores, are apt to have anywhere from 1 to 15 pct copper on charge, with a variable in arsenic of from 0.5 to 2 pct. With such conditions, and drosses varying from 12 to 40 pct copper, an operator using the older dross reverberatory practices would find it just about im¬possible to control the reverberatory operations and to produce a good product low in lead and high in copper. These conditions prevailed at El Paso from the year 1940 to the present. In May 1942 it, was decided to experiment with the soda process. THE SODA PROCESS After considerable work the soda method was finally evolved to the present state. It has never failed to produce a good product even under the most adverse conditions. It might be well to remember that many men of thought in the smelting industry have at one time or other suggested the removal of all copper, arsenic and antimony from the lead blast-furnace charge as an aid to better metallurgy.

Jan 1, 1947

SINCE I have told you of the distillation of waters and the extraction of oils from things-all ingenious and useful processes-I wish to continue to speak of the arts. I am reminded that I first wish to write down for you the discourse that I made to you on another occasion a few days ago concerning the operation of a mint, so that if you have lost it from your memory (and this would not be surprising since you have had no practice in it) you may repossess it if you should need to practice it or even to talk about it, and if you find yourself in this activity it may not be new to you. I told you, as I believe you will remember and so again I repeat, that a very great and constant diligence is required by one who wishes to operate a mint well, or to have it operated, because it has many parts which it is necessary for anyone who enters this work to understand very well; indeed, if he wishes that all the work of his helpers be done well, he will have to do it himself. And if it were possible, he should have a hundred hands and a hundred eyes together, and at the same time should be in different places, because some loss often occurs in what is handled because of the carelessness of the workers or because of their little real trustworthiness. It is understood that such things require great hard labors and many kinds of knowledge. The material handled in this work is most often either gold or silver all ready to be spent, and these things are very attractive to those who handle them. In addition to being made into pieces of money likely to disappear, it is very easy to lose, and whatever part is lacking for any reason whatever [132v] brings detriment and loss to the patron since it is a thing of value. Therefore, in short, anyone who wishes to exercise this art well, without being fraudulent therein, must watch the weights with all care because the essence of this art is naught but the division of a quantity of weight into many exactly defined pieces and the alloying of gold by carats and of silver by leghe, exactly established in accordance with what the prince grants you as remedy.* If through negligence you exceed in perfection

Jan 1, 1942

By W. M. Brown

SEQUACHEE Valley includes portions of the counties of Marion, Sequachee, Bledsoe and Cumberland. It extends in a general direction parallel with the Great Valley of East Tennessee, some 75 miles northward from the Alabama state line. It is separated from the Valley of East Tennessee by a mountain arm about a dozen miles wide, known as Wallen's or Walden's Ridge. The genesis of the valley, geologically speaking, is simple. It is an erosion-valley along the axis of an anticlinal. The point of the wedge remains in a row of low "saw-tooth" hills, running up the center of the valley. These hills are of Lower Silurian age, and consist of Trenton limestone, covered and protected in some measure by chert. Walls of Carboniferous age form the inclosing mountains on either side; whilst a magnificent arch of Hudson River, Niagara, Hamilton and Subcarboniferous strata closes the valley at the

Jan 1, 1886

By Frederick A. Mumpton

For more than 200 years zeolites have been familiar minerals to geologists and mining engineers as minor, but ubiquitous constituents in vugs and fractures of most basalt and traprock formations. More than 35 different species have been recognized including several, such as chabazite, erionite, mordenite, and faujasite, whose adsorption properties rival those of many synthetic molecular sieves. Igneous occurrences of this type have never been mined on a commercial scale; they are generally small and low grade, and they commonly contain as many as four zeolite minerals, along with small quantities of calcite, quartz, and other silicate minerals. Their economic beneficiation to monomineralic zeolite products has also not been achieved and, for the most part, industrial applications of zeolites have relied primarily on synthetic molecular sieves that were crystallized from hydrous aluminosilicate gels. During the middle 1950s, when Union Carbide Corp. and other groups were developing processes and markets for their fledgling synthetic zeolite businesses, natural zeolites

Jan 1, 1975