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By Barry D. Lichter, Pierre Sommelet

High-temperature heat contents of InSb, GaSb, and AlSb were measured over the temperature range 400" to 1450°K using a diphenyl ether drop calorimeter. Smoothed ualues of the thermal properties, H$ - H:9s, have been derived and are tabulated at even temperature intervals. The heats of fusion of the three compounds were determined as, respectively, 5707 *100, 7780 * 100, and 9800 5 3011 cal per g-atorn at the determined melting points of 797" l°, 985" * , and 1330" * 5°K. The calculated entropies of fusion are, respectively, 7.16 0.12, 7.90 * 0.11, and 7.37 * 0.23 cal per deg per g-atotn. The heat capacities increase substantially on tnelting in contrast to the behauior of structurally related germanium and silicon. Deriations from the Kopp-Neumann rule are negative for solid compounds and positive for the liquid phases. Previously obsevrrd "post melting" in InSb is confirmed. The high-temperature thermal properties of 111-V compounds are presently not well-established, despite the technical importance of these semiconducting cbmpounds. Uncertainties in available heats of fusion and heat contents have seriously hampered thermo-chemical evaluations' and thermodynamic analyses of phase equilibria2"" in these systems. This paper reports results of high-temperature heat content investigations of InSb, GaSb, and AlSb measured in the range 400" to 1450°K employing a diphenyl ether drop calorimeter. Similar measurements for InAs and GaAs will be reported in a following publication.~ EXPERIMENTAL PROCEDURES Samples. High-purity samples of the compounds InSb and GaSb were supplied in the form of crushed crystal fragments by Dr. Carl Thurmond of the Bell Telephone Laboratories and in the form of single crystals by Dr. A. Strauss of the M.I.T. Lincoln Laboratory. Single-crystal samples of semiconductor-grade AlSb were supplied by Dr. W. P. Allred of the Bell and Howell Research Center. Chemical analyses indicated that all compounds were stoichiometric to *0.1 at. pct, which is within the experimental uncertainties of the analyses. Samples were crushed, weighed, and encapsulated in evacuated, thin-walled, fused silica capsules. The capsules were nearly identical in external shape, 2 cm by 2 cm diam, but varied in weight due to differences in wall thickness. One sample of AlSb was contained in a thin-walled, high-purity alumina cup encapsulated in silica and used for heat content determinations of liquid AlSb. The capsule materials showed no visual evidence of reaction with any of the compounds. The sample and capsule weights are given in Table I. Calorimeter. A Bunsen-type calorimeter, similar in design to a previously described instrument6 but employing diphenyl ether (C6H5)'0 as the calorimetric substance, was used for measurements of heat contents above 300.0°K, the melting point of diphenyl ether. Heat input to the calorimeter caused isothermal melting of diphenyl ether, and the resulting increase in volume was measured by displacement of mercury from the calorimeter into a 200 cm horizontal, calibrated capillary, 1.25 * 0.01 mm diam, or into a weighed beaker. The advantages of diphenyl ether over water have been previously pointed out7 and include: i) an increase by a factor of 3.5 in sensitivity as measured by the ratio of the volume change to the enthalpy change on fusion, ii) the smaller required extrapolation from the melting point to the standard temperature of 298.17"K, and iii) the positive volume change on fusion of diphenyl ether in contrast to the contraction which occurs on fusion of ice. Diphenyl ether was purified by fractional crystallization to 99.95 mol pct as determined from the melting point depression with fraction crystallized. During assembly of the calorimeter, the ether was repeatedly outgassed under high vacuum to remove dissolved air. The calorimeter receiving vessel consisted of an 8-in.-long by 1:-in.-diam copper tube with twelve horizontal 3-in.-diam radiator "fins" for dissipating heat to the surrounding mantle of diphenyl ether. Before forming the mantle the chamber surrounding the receiving vessel contained 3300 cu cm of liquid diphenyl ether above 250 cu cm of mercury which was

Jan 1, 1970

By Robert M. Dreyer

Diamond is the hardest known material. The diamond industry is separated into two major segments: (1) industrial and (2) gem. The major industrial use of diamonds is as a high-grade abrasive in a wide variety of applications for which no effective low-cost substitute is available. Table 15.4C.1 gives data on diamond production and on United States diamond imports. In both gem and industrial diamond production, but especially in the case of gems, the quantity of stones produced is not directly correlated to value, for the determination of value is markedly affected by quality. South Africa long has been the world leader in gem diamond production with an estimated production in 1968 (including South-West Africa) of about 4 million carats, equal to 43% of world production. During the past decade there has been a marked rise in gem diamond production in the USSR, so that nation is now the second-ranking world producer, with Angola (Portuguese West Africa) third. In natural industrial diamonds, the major production comes from Congo-Kinshasa (former Belgian Congo) which had a production of 15 million carats in 1968, equal to 47% of world production. In the past six years, there has been a remarkable increase of industrial diamond production in the USSR, which now ranks second with a production of 6 million carats, equal to about one-fifth of world production. South Africa now ranks third followed by Congo-Brazzaville and Ghana. No natural diamonds are produced in the United States. Since the discovery of an economical method for synthesizing small diamonds, there has been a rapid increase in synthetic production from 4 million carats in 1962 (1 carat equals 200 mg) to 11.5 million carats in 1968. Table 15.4C.1 gives the gradually increasing imports of industrial diamonds and indicates a relatively stable price. In 1968, the imports of industrial diamonds were 13.5 million carats with a value of $59.2 million. Table 15.4C.2 shows the rapidly increasing imports of cut and uncut gem diamonds into the United States. In 1968 the United States imported 2.5 million carats of uncut stones with a value of $166 million and 1.5 million carats of cut stones with a value of $175 million. Diamond marketing is dominated by DeBeers Consolidated Mines, Ltd., which, through subsidiaries, controls many of the major African diamond deposits. More important economically, however, through its Central Selling Organization, DeBeers markets 86% of the non-Communist world's natural diamonds, including most of those derived from deposits outside of corporate control. In 1967 the mines owned by DeBeers produced nearly 8 million carats, and the Central Selling Organization sold over $600 million in gem and industrial diamonds. The dominance of diamond sales by

Jan 1, 1976

By C. S. Herzig

In the Engineering and Mining Journal of Jan. 27, 1900, there appeared an article by Fred T. Greene, describing a method of measuring stopes by the use of strings, a clinometer and a tape. In the early part of 1899 I had occasion to use a somewhat similar method, for the purpose of accurately ascertaining the volume of a very irregular drift connecting two mines. This drift was an important factor in a lawsuit for $150,000, in which it was necessary to determine accurately the volume originally occupied by the material extracted from the drift. While it is unnecessary to give all details of the controversy, some explanation of the conditions is required for a clear understanding of the following description. This drift was started by the a E." Co. from its workings, in about 7 ft. of ore, and headed toward the " G " property, the plan being to extract all the ore between foot- and hanging-walls, but to take out nothing but ore. As the " G." workings were approached; the ore-body became very irregular and much thinner. It was, however, very closely followed, and where the drift broke through into the " G." workings, the hole was only 11 inches high by 14 inches wide, or just about the size of a man-hole in a boiler.. Thus, this connection-drift tapered sharply; and the very irregular foot-wall sloped toward the " G." workings, and entered them at a place where the stope was only about 3 ft. high, and the ground had already commenced to cave. It may be added that the ore-body was a nearly flat deposit of lead carbonate in limestone, the ore and gangue being quite soft. I was called upon by the " G." people to make the survey

Jan 1, 1901

By K. A. Svanstrom, W. R. Opie

Problems associated with deposition of titanium infused chloride baths using TiCl4 as a feed material are reviewed. A potentially workable cell design using Alumdum diaphragms is discussed. Problems inherent to diaphragm-type cells are emphasized and experiments described which led to the development of a cell without a diaphragm. INVESTIGATORS assigned the problem of developing a commercial electrolytic process for making titanium metal have found that they are faced with the difficulty of overtaking rapidly improving magnesium and -sodium reduction processes. The ultimate goal must be a simple process that can be conducted with low-cost raw materials in equipment that is inexpensive to build and operate. This paper deals with some work of a preliminary nature conducted by personnel of the Research Laboratories, National Lead Co., Titanium Division. Many others have reported on the general subject of depositing titanium from fused salt electrolytes. A partial bibliography is attached. PRIMARY CONCEPTS Anyone attempting to deposit titanium metal from fused chloride electrolytes using TiC14 as a source material is faced with at least the following problems: 1) Getting the titanium into solution in the bath. 2) Controlling oxidation and reduction reactions associated with multiple valency of titanium, which result in low electrical efficiency. a) Prevention of dichloride oxidizing to trichloride and trichloride to tetrachloride at the anode. b) Prevention of trichloride from being reduced only to dichloride at the cathode. 3) Prevention of metal formation away from the cathode thru disproportionation of titanium dichloride to satisfy the equilibrium. 3 TiCl2 == 2 TiCl3 Ti [I] 4) Protection of reduced chlorides in the bath from reaction with oxygen from the bath container or air. 5) Removing from the bath the deposited metal formed without allowing it to become contaminated with oxygen or nitrogen. 6) Selecting cell materials of construction which will not contaminate the metal product or cause cell inefficiencies. EARLY EXPERIMENTS The initial attempts made to electrodeposit titanium from fused chloride baths using TiCl4 as a source material indicated that TiCl4 was essentially insoluble in molten chlorides of alkali and alkaline earth metals. If, however, the TiCl4 was introduced close to a nickel cathode some metal could be formed and appreciable quantities of reduced chlorides would be present. These reduced chlorides reacted with any oxygen present in the atmosphere above the bath and would also diffuse quickly to the anode (graphite) where they would react with chlorine and be evolved as TiCl4. The net electrolytic efficiency of such a cell is essentially nil. The TiCl4 evolves from the anode at almost the same rate that it is introduced in the vicinity of the cathode although a small amount of titanium metal can be formed on the walls of the cathodic pipe. It becomes apparent from the operation of such a cell that: metal can be deposited on the cathode, although of small particle size and a nonadherent nature, and when reduced chlorides are present in the electrolyte, TiCl4 can be absorbed, probably by the following reactions: TiCl4 + TiC12 - 2 TiCl3 [21 2 TiCl3 + 2 faradays - 2 TiC12 + C12 [31 DIAPHRAGM CELLS These facts encouraged the development of cells containing diaphragms to prevent migration of reduced chlorides from the cathode area to the anode. Small-scale cells for process development are described by Alpert, Schultz, and Sullivan.20 A larger brick-lined cell incorporating internal heating by passing ac through the bath is shown in

Jan 1, 1960

By Scott, Turner

IN 1933, the monetary metals were produced in a ratio of 6.7 oz. of silver to 1 oz. of gold, the lowest relatively for silver since the period from 1851 to 1865. At the beginning of that period, the virgin placers of California -and Australia began producing almost simultaneously; in consequence, there was an abrupt fall in the high silver side of the ratio that had prevailed through the centuries after America was discovered. The output from the Comstock and other silver-bearing lodes, located a decade after this influx of new gold began, soon restored the production ratio to its former proportion, that is, a relatively large weight of silver compared to gold.

Jan 1, 1934

By Roy A. Anderson, Roger H. McConnel

The Metaline district from 1906 through 1965 has produced nearly 16 million tons of ore yielding 400,808 tons of zinc and 178,062 tons of lead. The sediments, ranging from Precambrian into the Devonian, are schist, phyllite, conglomerate, greenstone, quartzite, limestone, dolomite, and carbonaceous slates. Quartz monzonite and granodiorite of the Cretaceous(?) Kaniksu batholith are exposed over a large area beginning a few miles south of the district. There is minor Tertiary conglomerate and basalt. Abundant glacial deposits obscure much of the bedrock. Northeasterly trending folds, having steep to overturned northwesterly limbs, and thrust faults, striking northeasterly and dipping southeasterly, characterize the deformation of the Precambrian and Paleozoic rocks. Recent drilling indicates that the Slate Creek fault, too, is also a southeasterly dipping thrust fault. This new evidence suggests that the Valley Block, like the other major structural blocks, may in part be a thrust plate rather than a graben, and that thus the district may have greater structural unity with the surrounding region than has heretofore been apparent. The Cambrian Metaline Formation of limestone and dolomite contains essentially all the lead-zinc mineralization. Most of the ore mined to date has come from a carbonaceous and irregularly siliceous breccia of sedimentary and diagenetic origin, locally called the "Josephine Horizon," that makes up the top few to 200 feet of the formation. The Josephine Unit is a breccia of gray limestone, coarse light gray dolomite, very thin-bedded, dark gray siliceous dolomite, and "zebra rock" believed to be of algal origin, embedded in a matrix of black dolomite or black jasperoid or some mixture of these two. The jaspcroid fragments and irregular blobs embedded in the black dolomite matrix and the jasperoid laminae in the very thin-bedded dark gray dolomite suggest that jasperoid was formed at the same time as the breccia matrix and as the laminated dolomite. So too, the marked tendency of sphalerite and galena to be disseminated mainly in the dark breccia matrix and their tendency to be most abundant in the jasperoid laminae of the very thin-bedded dark gray dolomite, even where the mineralized and laminated dolomite is fragmented and engulfed in black dolomite breccia matrix, suggests that in these ha bits the sulfides also were deposited contemporaneously with the sediments. Yet very commonly, veins and vague veinlets and patches of white quartz, usually carrying rather coarse sphalerite and sometimes galena crosscut all rock types. These contradictory relationships throughout the Josephine Unit suggest that both the black jasperoid and the sulfides were original constituents of the Josephine Unit but that, locally, they later were remobilized and deposited as white quartz and coarser grained sulfides. Possibly, at least in part, this "metamorphism" is one of the effects produced during or shortly after the intrusion of the Kaniksu batholith.

Jan 1, 1968

By F. P. Schwarz, A. J. Erickson, S. M. Weber

The Mount Hope stockwork molybdenum deposit, Eureka County, Nevada, occurs In a small igneous cauldron complex 32 to 38 mybp in age. The deposit is characterized by alteration zones and mineralized shells symmetrically distributed around multiple intrusive. For samples from the 25000 meters of 1981-82 drilling, an appraisal of error in in-house sample preparation and in commercial assaying has been made. After splitting, drill core is crushed and reduced in volume (8: 1 at 10-mesh, 8: 1 at 70-mesh, 6: 1 at 200-mesh) to supply 25 grams of 200-mesh pulp for replicate analyses. A 0.5 g portion of the pulp 1s used for dissolution and atomic absorption analyses. Factors which initially generated unacceptable data variances have been minimized by planning, repeat analyses, introduction of standards, and routine monitoring. Assuming that error between replicate assays defines a Gaussian distribution with its average ideally equal to zero, precision (from preparation including core splitting and analysis of 8400 pairs of molybdenum samples) is on the order of ± 25%. The error distribution is not based. Accuracy is approximately 4% lower than absolute as determined by replicate assays of commercially prepared samples.

Jan 1, 1984

By L. H. Morris, M. Mordecai

The emission of firedamp from the strata surrounding active workings of longwall coal faces normally follows a cyclic pattern associated with the mining operations being carried out on the face. Mining a coal seam disturbs the uniform distribution of overburden weight such that the working is enveloped by a region of high stress, both above and below the seam, advancing with the coal face. There is a decrease in stress as the strata is relaxed following the passage of the face (1). The paper describes experimental results obtained during the initial investigations of a study intended to examine the permeability of samples of coal measure strata under various triaxial stress conditions. Several investigators have studied the effect of stress on the permeability of rock, usually in the field of oil reservoir engineering. Two general lines of approach are evident. One used by among others Fatt and Davies (2), McLatchie et a1 (31, Brace et al (4) and Gunther (5) was to subject a rock sample to hydrostatic pressure. Gas or water would be passed through a sealed core mounted inside a high pressure hydraulic bomb. The second approach, favoured by Wilhelmi and Somerton (6) and also by Gray et a1 (7) was to employ a so-called 'triaxial' state of stress. Here radial confining pressure and axial stress are applied independently. The permeability is measured in the direction of the axial load. With the exception of Gunther who tested coals and Brace et a1 who tested a granite, the rocks used have been sandstones with permeabilities generally appreciably higher than British coal measure rocks.

Jan 1, 1971

By John G. Reid

Subsequent to extensive laboratory testing, a pilot plant was constructed to further evaluate the process for high temperature copper cementation on iron in continuous operation. The plant was designed to produce 1 to 10 tons of copper a day depending on the copper concentration of the pregnant solution. The energy required to operate the plant was very low as virtually all the solution heating requirements were supplied by the exothermic cementation reaction. The results demonstrated that this, process has potential for commercial application with respect to operating considerations, and yields a copper product in excess of 95% copper at high reaction rates, low iron consumption, and at least 98% copper recovery from solution.

Jan 1, 1976

By G. W. P. Rengstorff, H. B. Goodwin

A method is described for purifying iron in batches of 150 Ib or more. Oxygen, carbon, nitrogen, and sulphur are removed from flakes of electrolytic iron by treatment in wet and then dry hydrogen. A special consumable-electrode arc furnace is used to remove hydrogen and to melt the flakes into ingots. A LTHOUGH iron and steel have been made and used for hundreds of years, there are still wide gaps in the knowledge of these metals. One reason for this is that the true properties of the base metal, iron, are not well known, simply because really pure iron has never been readily available for research purposes. Commercial iron, and even the "pure" irons which are readily available, are really complex alloys of iron with oxygen, nitrogen, hydrogen, carbon, sulphur, silicon, and other elements. With such materials, it has not been possible to prepare steels and other iron-base alloys in which all elements are present in exactly controlled amounts. It is an almost impossible task to determine the separate effects of each element or the intereffects of several elements unless alloys can be prepared containing only the desired elements. The ability to make controlled alloys of low impurity content is of far more than academic interest. It will assist in the study of such practical problems as the brittle failure of ship hulls or high pressure gas pipelines, determination of the factors which affect the deep-drawing properties of steels, the improvement of magnetic and electrical properties, the phenomenon of aging, and many others. In 1949, the American Iron and Steel Institute asked Battelle Memorial Institute to prepare high purity iron on a larger scale than had been attempted up to that time. The work was directed toward the preparation of iron as low as possible in carbon, oxygen, nitrogen, hydrogen, and sulphur because these are the contaminating elements of major interest in many research investigations on iron-base alloys. The initial goal was to prepare iron with not more than a few thousandths of one percent of each of these impurities. The purpose was to set up a "bank" of high purity iron on which researchers could draw when a supply was needed and which could be replenished from time to time as the stock became depleted. Of several possible purification methods, heating in hydrogen was selected as the most promising for the removal of both carbon and oxygen from solid iron. Metallic impurities are not removed by hydrogen. Electrolytic iron was chosen as a starting material because it is the most readily available form of iron low in metallic impurities. Decarburization and desulphurization were obtained by annealing in wet hydrogen at relatively low temperatures. The optimum temperature, as found by previous investigators,' is about 1400°F. This treatment also removes nitrogen. Deoxidation is favored by dry hydrogen and higher temperature. There is evidence from previous work that an anneal at 2200°F in dry hydrogen will deoxidize the iron, reduce the nitrogen content, and reduce the sulphur content. These facts suggested the use of a two-stage process: treatment in wet hydrogen followed by treatment in dry hydrogen. To carry out the purification process in a reasonable time, it was necessary that the iron be in thin pieces. Flakes of electrolytic iron were used. Because most research requires that the iron be in massive form, it was nccessary to melt the purified iron to consolidate it into large pieces. For this work, it was decided to prepare ingots weighing at least 15 lb by using a cold hearth arc-melting process in which the iron was melted in an inert atmosphere of argon or helium. This method of melting was selected because the water-cooled copper crucible is not wet by the iron and does not contaminate it as a refractory crucible would. Because the crucible is also the ingot mold, contamination from mold materials was also eliminated. Contamination from electrodes was avoided by making them from purified electrolytic iron. The melting procedure finally developed was an adaptation of the Kellogg electric ingot process.' Two batches of purified iron weighing about 100 Ib each in ingot form were prepared with analyses as shown in Table I. The remainder of this paper describes the procedures in more detail. In general, the process con-

Jan 1, 1956

'T. C. MERRIMAN, New Haven, Conn. (written discussion *).-This most interesting paper gives much carefully obtained and valuable data. However, there are two points in connection with the commercial application of such data that might possibly be a source of trouble and mis-understanding. In the first place, the examination of thin sections of annealed brass, subjected to standard Brinell test.(500 kg. on a 10-mm. diameter ball) at thicknesses from 0.075 to 0.150 in. (1.9 to 3.8 mm.) indicates that cold work has been performed on the specimen during the application of the load sufficient to extend way through the section, and have the hardness of the metal affected by the backing. If the usual steel support is used, this effect is very evident. If a piece of soft brass is used as a support, there is an indentation in the surface of the support accompanied by a bulge on the under side of the tested specimen opposite the impression made by the 10-mm. ball. The result is that Brinell specifications (governmental or otherwise) on stock for a certain purpose may not be fair in all cases, since manufacturers, owing to differences in equipment, etc., may not all use precisely the thickness and conditions of stock on which the Brinell specifications were based. This would mean that the government Brinell specification might not give them the temper of stock best suited for their manufacturing methods. The concluding paragraph of this paper, stating that "the hardness of cartridge brass may be determined with greater accuracy by the Brinell hardness measurement than by attempting to judge it from the grain size" appears to be in line with the movement of the last year or so among brass manufacturers toward substitution of the Brinell test for microscopic examination as an acceptance test on cartridge brass. The statement may he strictly true that the Brinell test is the best test for hardness, but it is not, as the statement might readily be construed, a sufficient test of suitable condition. For instance, some cartridge brass might accidentally have been overheated to a point where it would be distinctly unsafe to use for the manufacture of small arms cases and then be so rolled (reduction 3 to 5 per cent.) as to pass perfectly proper Brinell specifications. Under such conditions the microscope would reveal its unsuitability where the Brinell test had failed so to do.

Jan 4, 1919

By Charles Will Wright

FRANCE will be given a large portion of the Marshall Plan funds for relief, reconstruction, and industrial development in France and in her colonial possessions. At present that country is not in position to export any large tonnages of the mineral products needed in the United States. However, if, through Marshall Plan aid and private capital investment, production of these products can be largely increased, agreement should be made by which the United States is to get a substantial portion of such increased output in part payment for the material and technical aid they receive. France issued a decree, published in the "Official Gazette" on Feb. 3, 1948, which permits conversion into dollars of dividends derived from dollar investment in Government-approved projects and will agree to a scale of capital amortization to be determined in each individual case. This decree is to encourage private capital to cooperate with local capital in industrial undertakings. The French Colonial Ministry has prepared an extensive program for aid to the mining industries in overseas possessions to be undertaken by

Jan 1, 1948

By Edgar Bain

SUMMARY THE results of an enquiry into the structural nature of some 70 iron alloys containing both nickel and chromium over a considerable range of concentration 'are briefly described in this paper. This study is a part of a comprehensive program of investigation into alloys involving chromium in progress at the Union Carbide and Carbon Research Laboratories. All the compositions studied contained carbon in amount comparable to commercial alloys.

Jan 3, 1927

By P. S. Roller

P. S. ROLLER,*College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author's quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should he defined as the distance between an F line and a y line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and y lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. x between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling. It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz." This would require many data to prove, but no proof is given; any consideration would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.," Cambridge, Mass. Contrary to Mr. Bond's stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids., Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystalline solids. Andreasen and Berg' have ground

Jan 1, 1941

By P. S. Roller

P. S. ROLLER,* College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author's quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should be defined as the distance between an F line and a [?] line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and [?] lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. I between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling. It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz," This would require many data to prove, but no proof is given; any considera¬tion would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.,* Cambridge, Mass.¬Contrary to Mr. Bond's stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids.5 Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystalline solids. Andreasen and Berg8 have ground

Jan 1, 1941

By P. S. Roller

P. S. ROLLER,* College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author's quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should be defined as the distance between an F line and a y line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and y lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. I between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling, It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz." This would require many data to prove, but no proof is given; any consideration would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.,* Cambridge, Mass.-Contrary to Mr. Bond's stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids.5 Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystal-line solids. Andreasen and Berg8 have ground

Jan 1, 1941

By Edward L. Beutner, Robert M. Crump

Benson Mines low-grade iron ore reserve is a replacement deposit within the Grenville gneisses of the Adirondacks. The average grade of the crude ore is about 23 per cent iron. The iron minerals are principally magnetite and hematite disseminated throughout the gneisses as discrete grains about 1 mm in size. Both the ores that are beneficiated by magnetic means and the non-magnetic ores that are treated by a gravity method of concentration are found in stratigraphically different positions but are mined from one open pit, which is about 2.5 miles long and 800 feet wide and are treated separately. The gneisses are described and named according to the distinctive mineralogy of each. Recognition of the gneissic types permits the interpretation of the structure as a northward-plunging syncline overturned to the west. A large subsidiary anticline-syncline complex occurs on the east limb of the major syncline. Iron minerals in sufficient quantity to be considered ore are confined to areas where the footwall rock has been overturned to become the hanging wall.

Jan 1, 1968

By J. Marin, H. A. B. Wiseman

This paper presents results of an experimental study dealing with the plastic stress-strain relations of aluminum alloy 14s-T6 subjected to combined biaxial tension and compression stresses. Plastic stress-strain relations for both constant and various types of variable stress ratios were determined and a comparison made with the simple flow theory of plasticity. THIS investigation was undertaken for two main purposes: 1—to obtain plastic stress-strain relations for Alcoa 14s-T6 when subjected to various combinations of biaxial stresses, and 2—to determine the validity of the flow theory in predicting plastic stress-strain relations for combined stresses. The biaxial stresses used were tension combined with compression. These stresses were produced by subjecting a thin-walled tube to axial tension and torsion. Various types of combined stress conditions were investigated. To provide control test data and information on the influence of biaxial stresses on the strength, the usual constant stress ratio type tests were made. These tests showed that the biaxial yield strength agrees approximately with the distortion energy theory. However, the Prager semi-empirical theory agrees best with the test results. For these constant stress ratio tests the plastic stress-strain relations were found to agree approximately with the flow theory. (For constant stress ratio the flow and deformation type theories are identical.) In view of the fact that the constant stress ratio tests cannot distinguish between the flow and deformation theories, a number of variable and special biaxial stress tests were made. These tests showed that the flow theory is inadequate, since large differences exist between the experimental and theoretical results. Special tests were also conducted to check the validity of the isotropic yielding assumption. This assumption is made in the linear-type flow theory. In these tests specimens were loaded in tension to predetermined values beyond the proportional limit stress. The tension load was removed from the specimen and the specimen was subsequently loaded in torsion to fracture. Other tests were applied with the order of torque and tension loading reversed. The nominal stress-strain curves for these two tests approximately coincide indicating that the assumption of isotropic yielding is valid. In addition, these tests verified the requirement by the slip theory of plasticity that prestraining below 140 pct of the proportioned limit stress in either torsion or tension did not influence the subsequent plastic stress-strain relations in tension or torsion. Another type of test was conducted to check the assumption that the axes of principal stress and strain coincide during plastic flow and variable stress conditions. These tests were conducted by applying combinations of axial tension, internal pressure, and torsional loads such that the axes of maximum principal stress could be rotated through 90" during the loading path. The results of those tests showed that the angle between the axes of principal stress and strain varied greatly and that the assumption of coincidence between axes of principal stress and strain is not valid. Introduction In various metal processes including forming of sheets, rolling of bars, and extrusion of rods, metals are subjected to stresses beyond the yield strength of the material. Often these stresses are not simple stresses acting in one direction, but are combined stresses acting in more than one direction. Structural and machine members are often subjected to combined stresses. To adequately determine the factors of safety for these members, it is necessary to know the plastic stress-strain relations of the material used for various combined stress conditions. Various theories have been proposed for predicting the plastic stress-strain relations for combined stress in terms of the simple tension plastic stress-strain relation for the material. To determine which theory might be adequate, test results must be obtained for various combined stress conditions, in order to compare the actual with the assumed theoretical behavior. Although there is considerable test data available for combined stress conditions in which the principal stress ratios are constant, relatively little test information is available for vari-

Jan 1, 1954

By A. Singh, W. M. Fleming, J. J. Santana, A. E. Torma

The present study investigated the extraction of uranium from a low-grade ore by sulfuric acid and the removal of radium from the leach residues by brine solutions. The optimum leaching conditions were found to be 1.5 mole [£]-1 H2SO4, 6.Og[9]-1NaClO3, 75°C,50% pulp density and 750 rpm with respect to the rate of uranium extraction. The maximum theoretical attainable rate of uranium extraction was determined to be V = 1.09 x 10 -2 mole [¢]-1h-1 with respect to pulp density. The order of reaction was found to be zero regarding the effect of sodium chlorate concentrations. The apparent activation energy of uranium extraction was calculated from the Arrhenius plot to be [A]E a - 2.36 kcal mole-1 (9.87 kJ mole-1). Based on a statistical design, an empirical equation has been developed describing the rate of uranium extraction with the leaching parameters. This equation can be used to generate data which may be useful for scaling up the process to pilot plant level. The most efficient brine solution for the extraction of radium - 226 from the leach residue was found to contain one mole[f] -1 CaCl2 and 0.01 mole [2] -1 HC1. The two step leaching with this solution resulted 87% removal of radium from the leach residue. On the basis of this study, a flow-sheet has been proposed for the extraction of uranium and radionuclides from the ore and leach residue to yield almost innocuous tailings.

Jan 1, 1982

By Reynold Q. Shotts

The small domestic underfeed stoker as now designed is unusually sensitive to the coking and plastic properties of coals, and when the attempt is made to burn the high rank coking and caking coals of the Appalachian fields, serious trouble may occur. Part of the trouble may be caused by the ash content and clinkering characteristics of the particular coal in use, but more often it is the result of an excessive ac-in of coke in the form of ''trees,, or masses in the fuel bed. Often the zone of burning becomes so restricted that the fire goes out, and in extreme cases the fire may be lifted completely out of the fuel bed and away from the air ports. Even if the fire is preserved, excessive coking may cause response to subsequent heat demand to be slow and irregular. Coke trees seem to grow most readily during periods of hold-fire operation, so that in order to preserve the fire in mild weather it is often necessary to set the timer for such a long operating period each hour that much unneeded and unwanted heat is developed, causing discomfort. During demand periods the fire may pick up so slowly that the intense burning which develops just before stoker cut-off causes a serious overshooting of the thermostat. Stoker coal performance tests which have been made in a number of labora- tories1,2,3,4,5 are one method of approach to the problem Of selecting stoker coals- As yet these tests have not been standardized and they are expensive and time-consuming. The value of the results probably increases with the length of the test. Sherman and co-workers at Battelle Memorial Institute believe that even a one-week test is not too reliable a trial2 and it is quite likely that Others will agree. A great advantage of such tests is that characteristics of coals, other than the tendency to form coke masses, may be evaluated from them. The selection of stoker coals would be greatly facilitated if there existed some simp1e measure Of the tendency Of a a to form coke in a fuel bed. Although possessing all Of the disadvantages Of any small-scale laboratory test, the test for the free-swelling index of coal has been found useful for evaluating the probable effects Of swelling properties upon combustion.7 Ostborg, Limbacherand Sherman, have shown that the test can Predict, roughly, the tendency Of coke to accumulate in the fuel bed Of a domestic stoker. They also gave instances of its value in predicting combustion behavior in some large installations. The use of the test for free-swelling index has become well established and has been adopted as a standard by the American Society for Testing Materials.8 When domestic stoker coal investigations began at the State Mine Experiment Station in the summer of 1944, it was decided

Jan 1, 1949