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By D. B. Masson

THE martensitic transformations occurring in ,B AgCd as a result of cold working at room temp.. a-ture and cooling below room ten1erature have been reported by Masson and Barrett.1 These authors found that simple cooling to 127°K caused body-cen-tered-cubic AgCd containing 53.0 at. pct Ag to transform to an orthorhombic structure, while mechanical deformation at room temperature was found to induce a hexagonal structure. Furthermore, they observed that the temperature at which martensite first appeared on cooling, Ms, and that at which the transformation was complete, MI, as well as the corresponding temperatures for the reversion on heating, As and Af, could be ascertained to ± 1°K, although they report these temperatures only for the alloy containing 53.0 at. pct Ag (51.96 wt pct). In the work reported here the temperature range of the transformation on cooling was studied as a function of composition, and further consideration was given to the relationship between the original bcc structure and the transformation-induced hexagonal and orthorhomic structures. EXPERIMENTAL PROCEDURE All samples studied were in the form of annealed 300-mesh powder and were prepared from the constituent metals in the manner described previously.1 The compositions in weight and atomic pct are listed in Columns 1 and 2 of Table I. Method of determination of transformation temperature was the same as utilized previously1—measurement of the intensity of X-rays diffracted by the transformation product. A diagram of the low-temperature sample mount is shown in Fig. 1. This mount was made to fit a GE model XRD-3 X-ray diffraction unit. The powder sample was varnished to a flat surface on a copper sample holder at A. This sample holder was soldered to a brass rod, B, which was in turn soldered to a second copper rod, C. Rod C could be cooled to 78°K by inserting it into a reservoir of liquid nitrogen, D. A heater coil of nichrome wire was wound at E on the sample holder, and by varying the current through the heater a thermal gradient could be maintained along brass rod B. Sample temperatures from 78°K up could be attained and held as long as needed. The sample holder and nitrogen reservoir were insulated by Styrofoam, F, and the sample itself was protected by two radiation shields, G. The X-ray beam entered and left the sample chamber through windoys in the supporting container at H and I. Sample temperatures were measured by a calibrated copper-constantan thermocouple varnished to the sample holder just out of the beam at A. The low-temperature structure had been found to possess orthorhombic symmetry1 and the expected positions of the diffraction maxima were known. Therefore the sample was aligned in the X-ray beam and the Geiger counter of the spectrometer set to receive the most intense orthorhombic reflection without an adjacent interfering cubic reflection. This

Jan 1, 1961

By James A. Barr

THE writer's first experience with flotation was during World War I, in the beneficiation of Alabama graphite schist ores. One plant used a cone with a peripheral overflow; dried ore was distributed by a rotating disk onto the surface of the water. No reagents were added. The flake-graphite film floated along with some fine silica. Dewatering screens separated the flake. Naturally unit capacities were low but results were satisfactory. Minerals Separation mechanical flotation machines were used in another graphite mill. Pine oil was added directly to the head cells, with very good results. All the graphite concentrates from both types of flotation were carefully refined after drying by passing through burr mills, followed by very close sizing on sifters. In March 1927, the writer ran a very successful flotation test on minus 28-mesh, deslimed, Tennessee brown phosphate "sand,"[t] using a mechanical laboratory machine built in the shops of the Anaconda Copper Mining Co., through the courtesy of Ernest Klepetko. The machine was soon converted to a subaeration type, with improved results. This phosphate research program was greatly enlarged and led to the formation of the Phosphate Recovery Corporation. The development of nonmetallic flotation followed rapidly. The first flotation machine for beneficiation of phosphate was installed in January 1928, in a pilot mill, by International Minerals and Chemical Corporation, at Mulberry, Florida. These machines were Minerals Separation, 24-in., subaeration types, with weir control of pulp between cells. The sand bleeders were very small-about 1 ¼ -in. diameter. As the feed contained 15 to 25 per cent plus 28-mesh material, the cells choked up frequently. This was corrected by cutting a 2 by 6-in. bleeder opening above the impellers, at grid level, which short-circuited the coarse fractions directly into the feed of the next cell. The matter of size classification and grinding was a considerable problem, which does not come within the scope of this article. Because of the troubles noted above, cascade-type roughers were installed in the second flotation plant, built about 1930. Experimental cells had been tried in 1928. General plant layout is shown in Fig. I and details of the cascade construction in Fig. 2. The cascade installation in plant No. 2 gave good results and eliminated troubles due to choke-ups and coarse feed. The capacities and metallurgical results equaled those of the modified 24-in. Minerals Separation flotation machines in plant No. I The objections to the cascade installation are the floor space and mill height required. In 1939 an improved Minerals Separation machine was tested in parallel with the cascades. Extensive tests indicated

Jan 1, 1945

By E. DeGolyer

The problem of good engineering practice and. of good conservation practice in oil production is that of keeping gas in solution. This can best be done by producing a field as a water-drive field. Some degree of water drive is believed to be present in most fields but in the past most fields have been produced too rapidly to allow water drive to be effective. Ultimate recovery should be So per cent greater for fields produced by water drive over that from gas-expansion fields. Wide spacing will permit fields of low-degree water drive to be produced effectively as water-drive fields. INTRODUCTION NEGLECTING all the complicated scientific and technical argument that leads to the conclusion, the solution of the problem of good production practice, of good engineering practice, and of good conservation practice in oil production is to keep gas in solution. This can best be done by producing fields at a rate low enough to make effective the water drive, which exists, it is generally believed, to a degree great enough to be made effective in more than 80 per cent of the fields of the United States. Oil is being so produced in the Ten Section field of California; the Hobbs field of New Mexico; the Yates, Hastings, Friendswood, Anahuac fields and the main sand of the Conroe pool, in Texas; the Magnolia pool in Arkansas, and doubtless in many other pools. The measures of success in the efforts to achieve the solution of the problem are the maintenance of bottom-hole pressure and the prevention of an increase in gas-oil ratio. Some decline from the original bottom-hole pressure and the production of some gas are essential. A pressure gradient from the reservoir to the well is necessary and at least as much gas must be produced as that in solution in a barrel of oil. Once the optimum pressure and the gas-oil ratio have been established, they should be maintained. How common is water drive in oil and gas pools? General opinion is that some degree of water drive is almost universal. Theoretically, this conclusion appears to be a sound one. According to our present understanding of the processes of sedimentation, we would expect to find all sands or sandstone charged with connate water unless it had been displaced by some other fluid, such as fresh water percolating into the beds at the outcrop, or oil and gas, or gas only, in reservoirs. The fact that the original pressure of the fluid system in most reservoirs is approximately that of the hydrostatic head of water for the corresponding depth indicates that the sands are charged with water, and common experience in drilling into the extensions of the porous zones beyond the reservoir is to find them watercharged. The very fact that oil and gas pools have substantial pressures in their original undisturbed state indicates either that a reservoir is a rigid, closed system, a con-

Jan 1, 1941

By Hiroshi Oikawa, R. G. Blake, A. G. Guy

An equatzon has been derivedfor conveniently calculatzng the couvse of homogenzzation in a three-corviDonent solid solutiorz. The necessary assumptions are that the initial distributions of concentration are sinusoidal and that the four interdiffusion coefficients are constant. Employed in the deriuation were the Laplace transform and the fouter Fourier sine transfovwz. The final equation is shown to be in good agveetnent with the experimental results of Kirkaldyet al. on Cu-Sn-Zn alloys. COMMON boundary conditions for representing physical problems are those of the finite solid and of the infinite solid. Many metallurgical phenomena, however, have the geometrical form of a finite pattern reproduced throughout an infinite solid. Precipitation within a uniform matrix and segregation during solidification are examples. The purpose of this treatment is to show that a convenient and useful mathematical analysis is possible in those cases in which the finite pattern can be approximated by a sine function. The particular topic to be considered here is the elimination of segregation by diffusion, but the mathematical model has applicability in many other areas. It should be recalled that the mathematical solution of a typical problem involving a finite solid, such as diffusion into a thin layer, is quite involved.' Various types of solutions are used depending on the degree of advancement of the diffusion process and on the initial conditions. A step function is the simplest initial condition, but even in this case the general solution has the form of an infinite series. Kirkaldy et a1.2 have

Jan 1, 1968

By Herbert H. Kellogg

DEPOSITS of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine (-325-mesh) sand assaying 95.5 per cent Si02, and a very fine (-3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. EXPERIMENTAL PROCEDURE The Sample.-The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotation Feed.-By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine fraction was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper.

Jan 1, 1945

By Raymond Wilson

Grinding costs are an important item in cement manufacture, and the cost of power is one of the large items in grinding costs. Even where power is of secondary importance, cost items dependent on mill capacity are influenced by the efficiency with which the mill uses power. Tube mills typify the grinding mills most generally used in cement plants, though other types are in use. Tube mills have been widely criticized as being extremely inefficient, and this feeling regarding them has led to many attempts to design more efficient grinding equipment. Other types may be an improvement over tube mills, but the improvement has certainly not been revolutionatry, at least as applied to clinker grinding. The very fact that there is doubt regarding the relative efficacy of various types of mills is an indication that an adequate basis for an analysis of performance is lacking. The development of methods of analyzing mill performance has been hampered by the lack of a suitable method for measuring the work actually done in pulverizing. There has been no acceptable standard to which mill performance might be compared. As a consequence, it was not known whether the margin of possible improvement in pulverizing operations was large or small. Even direct comparison of different types of mills was seldom possible because most plants had only one type of mill for each type of service. Comparisons between plants were difficult because of differences in the grinding characteristics of various clinkers. Some recent developments complete a chain of circumstances that place the cement industry in a particularly fortunate position for the study of its clinker-pulverizing operations. Modern investigators of grinding phenomena have generally agreed that the best measure of mill performance is the production of new surface. For most materials this measure of performance is largely theoretical and has a rather remote connection with the practical aims of the grinding operation. In clinker grinding, on the other hand, surface area is intimately connected with the properties conferred on the cement by grinding. Morcover, the determination of surface area is rapidly coming into routine use as a tost

Jan 1, 1938

By D. R. Shoemaker

BACKGROUND AND PROJECT DESCRIPTION A 16.5 kt per day underground molybdenum mine is being developed at Questa in northern New Mexico by Molycorp, Inc., a wholly-owned subsidiary of Union Oil Company of California. Small scale underground mining first began in 1920 and large scale open-pit mining was initiated in 1965, but because of the high stripping ratio, surface mining is being phased out. Construction of the underground mine began in 1979, and first production is scheduled for 1983. Full production of 9 Gg of molybdenum per year will be reached in 1984. Access to the mine will be through a service shaft, development shaft, and a conveyor decline. 'The service shaft, 7.3 m in diameter and 396 m deep, will be used for personnel, supplies, development muck, and fresh air intake. It will be equipped with a large, double-deck man and material cage. A double-drum hoist will be used for cage hoisting and will handle up to a 27-t load in balance with a counterweight. Development muck will be loaded into skips in balance and a single-drum auxiliary hoist will provide secondary access. The 396-m-deep, 4.3-m-diameter development shaft will be used for initial mine development, mine exhaust air, and a slick line pass for concrete entering the mine. The development shaft is located 400 m north of the service shaft. Run-of-mine ore will be transported from one of two 1.4-kt storage pockets by a 1.2-m-wide conveyor belt located in the 10°, 1980-m-long, 5.0-m-wide, and 4.6-m-high decline. The decline conveyor belt has a centerline length of 2229 m and a vertical rise of 363 m. It will be powered by four 448 kw motors driving a single drive pulley. There will be two 1.2 m transfer belts on the surface with centerline lengths of 322 m and 542 m. All three conveyor belts will be reinforced with steel cables. On the surface, in addition to the headframes and hoist houses for the two vertical shafts, a large maintenance shop, warehouse, compressor house, change house, main office building and timber framing building have been constructed.

Jan 1, 1981

By W. A. Olds, H. F. Yancey, M. R. Greer

John Griffen (Consulting Engineer, Pittsburgh)— This paper presents much data offering tentative answers to many technical questions that have been raised by the employment of dense-medium processes and, particularly. by use of the drum-type vessel. It is to be hoped that the Bureau of Mines will publish complete detailed data for the 50 or more tests so that it may be studied and analyzed by everyone interested. Such study will take considerable time and effort. The writer is therefore obliged to confine his discussion to a few points he has noted as being open to question. The data in Table VI which shows the recovery efficiency for the 1 to ½-in. fraction as 100 pct raised a question concerning the accuracy of these calculated figures; washability data on the washed coal and refuse show that such a figure is impossible. When the ash content of the misplaced material in the washed coal is higher than that of the misplaced material in the refuse, recovery efficiency must be less than 100 pct. A etudy of the data In Tables 111, IV, V, and VI suggest that this error probably arose from slight errors in calculating the data in Table V and more probably from the small scale used in plotting the washability curves from the Table V data. Using the size consist data and the weight and ash percentages of individual gravity fractions of each size as given in Tables 111 and IV and the reported recovery of 80 pct of washed coal, a revised Table V was calculated, carrying weight and ash percentages to hundredths of a percent instead of rounding off at tenths. This washability data for each size of Belling-ham reconstituted feed was then plotted on large scale washability curves. From this data the new Table XIV was prepared, which agrees very closely with data shown in Table VI except for two important values, theoretical yields and recovery efficiencies. As will be noted the recovery efficiencies in Table XIV are significantly lower than those given in Table VI. A study was then made of the data showing distribution of the specific gravity fractions to washed coal and refuse. This study indicated that when this distribution was based on the fractions as tested at 0.1-sp gr intervals the number of points obtained were insufficient to establish the distribution curves with reasonable precision, which in turn would affect the accuracy of the values for the specific gravity of separation and the error areas. The data in Tables III and IV of the paper and curves prepared therefrom make it possible for the specific gravity fractions to be subdivided further. This was done using 0.05-sp gr fractions and the distribution data so obtained is given in Table XIV, with resulting values for the specific gravity of separation and the error areas. The writer believes that the derived values shown on Table XIV are more accurate than the correspond-

Jan 1, 1955

By P. K. Foh, C. G. Dunn

IN an investigation of the rolling and recrystalliz-ing textures obtained from single crystals of Si-Fe, Koh and Dunn included specimens in the (111) [lie] orientation.1,2 This note reports some additional results obtained on samples of this orientation set aside for further study. After, primary recrystallization longer annealing treatments at 980°C produced only normal grain growth; after 16 hr at 980°C the largest grains were approximately 0.010 to 0.020 in. in diam. The orientations of 13 of the larger grains were found by the transmission Laue method. The unidentified (110) poles of these grains have been added to the (110) pole figure obtained after primary recrystallization2 to give the plot shown in Fig. 1. When grain identifications are also included, such a plot reveals the following. Two orientations agree almost exactly with the peak pole ,density positions of component B of the primary recrystallization texture, one with component A, and one with component C. All the remaining orientations except one have (110) poles near weaker and unidentified pole density peaks. In fact, the correlation is good enough for an identification of eight components. Previously only components A, B, and C were certain. The results are perhaps even more informative when combined with the (110) pole figure of the cold-rolled crystal to give the plot shown in Fig. 2. Each of the 13 orientations is identified and uniquely determined by selecting two (110) poles per crystal, which are 90" apart. The plot shows that every grain of the 13, except grain 6, has a (110) pole in common with a (110) pole concentration of the cold-rolled crystal. The orientation relationship with the cold-rolled orientation may be expressed as approximately 25" rotations about <110> axes, which is in agreement with the results found for coarse-grained primary recrystallization.2 If primary recrystallization had occurred according to the <110> rule without preference, there would have been 12 components. Actually our limited data reveals nine of the possible 12. The integrated pole figure also shows preferences among them, since three are considerably stronger than the rest. In addition to the above direct information, the following conclusions may be drawn. Primary recrystallization of the (111) [110] cold-rolled crystal produces a limited number of weak and strong components with the larger than average primary grains from these components related to the cold-rolled crystal by 25° <110> relationships. In arriving at this conclusion of limited orientations the (111) [110] orientation has to be ruled out as a source of large primaries for two reasons. First, growth in this orientation was not observed, despite an expected high grain boundary mobility (due to its 25" <110> relationship with every component of the texture). Secondly, primary recrystallization usually produces grains in new orientations, particularly in a single orientation matrix; but (111) [110] is not new. Other orientations besides those of the <110> type may be considered unlikely, of course, because the <110> relationship holds quite well for primary recrystallization2 and should produce the large primaries. There are exceptions to this rule, but probably not enough to provide large

Jan 1, 1958

By F. A. Forward, A. H. Ross, J. Halpern

IN treating carnotite ores by carbonate leaching it is often necessary to subject the ore to a prior roast. Among the advantages that may result from roasting are: 1) improvement in settling and filtering characteristics of the ore; 2) destruction of carbonaceous impurities that would otherwise contaminate the leach solution and consume carbonate reagent; and 3) substantial increase in extraction of vanadium, which is usually low when the un-roasted ore is leached. Recovery of vanadium may be an important factor in processing carnotite ores, and the fact that vanadium is usually recovered in good yield by acid leaching has often been an economic advantage over results obtained with carbonate leaching. It has been found that response of an ore to carbonate leaching, following roasting, depends primarily on its composition and on the conditions of the prior roast treatment, particularly with regard to the roasting temperature, the atmosphere maintained above the charge during roasting and cooling, and the presence of certain reagents. These factors, especially the last, virtually determine the extractions of uranium and vanadium attainable by carbonate leaching, as well as the reagent consumption on leaching, whereas adjustment of the leaching conditions has only a limited influence. These variables were examined systematically to elucidate roasting and carbonate leaching of carnotite ores and to establish a treatment yielding high recoveries of both uranium and vanadium. Tests were conducted on samples of four different Colorado Plateau ores, described in Table I. Roasting Procedure: The roasting tests were conducted in an electrically heated furnace in which the temperature was controlled to +15oC. The ore (usually a 500-g charge) was ground to —100 mesh, mixed with the desired reagent, and spread in a silica tray to a bed depth of about 1/2 in. During the roast the charge was rabbled at intervals of 15 min. To provide free access of air the furnace was usually ventilated; however, in a few experiments a controlled atmosphere such as carbon dioxide or hydrogen was maintained above the charge. Leaching Procedure: The pulp, generally comprising 500 g of ore and a 2000-cc solution containing 80 g per liter of Na2CO3 and 20 g per liter of NaHCO3, was agitated in an autoclave at 120°C, under 30 psi O2. Samples were withdrawn periodically, and the solution and ore residue were analyzed. Under the conditions employed, maximum attainable extractions were generally achieved in 4

Jan 1, 1958

By H. J. Rahilly

Mine fires, in the Butte district, have been a source of trouble and expense for the past thirty years, for while the actual fire area in most of the mines has been comparatively small, the handling of the gas has generally been a much larger problem than the actual putting out of the fire. The chief causes of these fires have been: Defective electrical equipment, incendiarism or carelessness, and spontaneous combustion. While fires from the first two causes have been numerous, they have generally been discovered and extinguished before they became very serious. Spontaneous combustion has been the cause of most of the large fires. Sufficient heat has been generated by ground movement, high rock temperatures, and decomposition to set fire to inflammable materials (tarred ropes, canvas, dry timber, manure, and hay) left in the gobs some years ago. Also, the heat resulting from the oxidation of the fine broken sulfide ore is sometimes sufficient to set fire to the timbers. In 1906, a fire was discovered on the 1100-ft. level of the Minnie Healy mine about 400 ft. (122 m.) west of the Minnie Healy shaft, which presumably originated, in one of the upper levels, from spontaneous combustion. The fire was walled off by concrete bulkheads and a pillar of ground 100 ft. (30 m.) thick was left below and mining resumed underneath. During subsequent operations, this pillar became cracked and broken and the fire passed into the new workings. It travelled downward and spread rapidly until, in 1917, it had become necessary to wall off portions of the Tramway, Leonard, and West Colusa mines on the 1300, 1400, 1500, 1600, 1800, 1900, and 2000-ft. levels, and the tops of the 2200-ft. raises were sealed. The fire area now consists roughly of a block 1000 by 1200 ft. and from the 600-ft. to the 2200-ft. level. This block is composed largely of a series of parallel northwest-southeast nearly vertical veins and stringers that form a large orebody, with no well-defined walls but badly fractured and faulted, and known locally as "horsetail ore." These veins, or orebodies, are mined by the square-set or the rill method.

Jan 1, 1923

By Donald Barton

Geophysical methods of prospecting taken as a whole do not seem to offer much promise to a young man planning to enter them in the future. They have come to stay, to be sure, and they will continue in at least moderate use for a long time. Domestic use of those methods presumably is at a maximum at the present time. Geophysical pros-pecting is expensive; the effects of the law of diminishing returns in efficiency in finding prospects can already, be recognized. Small oil companies, independent oil operators, and royalty buyers have been doing much geophysical work. But in face of the increasing costs of discovery of prospects and in face of the probably decreasing favorability of the prospects discovered, I expect to see a sharp drop in their use of geo-physics within the next 10 years. The major oil companies and the larger small oil companies will continue to use geophysics indefinitely, but it seems doubtful whether the very extensive domestic campaigns of a few companies will be maintained for many years, certainly not for 10 years. The use of the geophysical methods in foreign countries has not yet reached its peak, and probaby will continue to increase for several years. That peak level of activity will probably be held for 10 years, although the site of maximum activity will shift from country to country; by the end of 25 years from now, I should expect the foreign use of the methods to have leveled off to one-fourth to one-half the peak level. Foreign service in geophysics does not seem to offer an attractive career, except for men of exceptional disposition. Young men just out of college, who are willing and who can afford to spend three or four years having a fling at adventure before settling down, will find it attractive, but when their tour of foreign service is over, they probably will have harder work finding a job than when they were seniors, and their foreign experience will count for little, except for further foreign service. A few men, mainly the higher men, will have headquarters at civilized or semi-civilized cities and towns or at permanent oil-company towns, and can

Jan 1, 1938

By D. Turnbull, J. H. Hollomon, J. C. Fisher

UNDERSTANDING of the diffusion problem has recently been furthered by the analysis of Birchenall and Mehl.1 They pursued the problem of the variation of the diffusion coefficient with composition for non-ideal solid solutions. Their analysis follows the tradition of Wagner,2 Jost,3 Dehlinger,4 and Darken,&apos; who assumed the net flux of dif¬fusing material to be proportional to the gradient of a quantity related to free energy. The basic assumption in Birchenall and Mehl&apos;s formulation of the problem is that the net flux of diffusing material is proportional to a thermodynamic activity gradient. To estimate the applicability of the derived relations, they corrected the diffusion coefficients of carbon in austenite by utilizing Smith&apos;s&apos; activity data for carbon, and the diffusion coefficients for zinc in copper with activity data computed from the vapor pressure measurements of Hargreaves.7 The corrected coefficients for carbon were virtually independent of carbon concentration. For the diffusion of zinc in copper a moderate variation of diffusion coefficient with concentration remained after the correction. The problem of diffusion can be formulated in another way than that used by Birchenall and Mehl. The method is that of chemical rate theory and has been used by Eyring and his coworkers8,9 as well as by Eyring and Zwolinski10 in a discussion to Birchenall and Mehl&apos;s paper. As yet the method has not been applied to substitutional solid solutions. On the basis of their assumptions, Eyring and Zwolinski derived a relation between the diffusion coefficient and activity that is identical with that of Dehlinger and Darken but which does not correct for the variation of the diffusion coefficient with concentration as well as does the relation derived by Birchenall and Mehl. It was thought that the theory of diffusion would be more complete if the inconsistency between the two analyses could be rationalized and if diffusion in substitutional alloys were analyzed on the basis of the concepts of the chemical rate theory. CHEMICAL RATE THEORY According to the theory of absolute reaction rates8 the reaction [A,+A2+ ... +.A;->X,+X2 +...+X; EqI] is viewed as taking place in two steps. First the reactants A1, A2, . . . A[i] combine to form an activated complex M with which they are in equilibrium; [A, + A2 + .. . + Ai = M. Eq 2] Second the activated complex decomposes to give the products X1, X2. . .. X[j]; [M->X1+X2+ ... +.X;. Eq3] It has been shown8 that the number of activated complexes decomposing per unit volume per unit time is given by the expression† [kT]

Jan 1, 1948

By H. M. Chance, T. M. Chance

The term &apos;Llow-sulfur coal," as used in this discussion, is limited to coals containing less, or very little more, than 1 per cent. sulfur. For certain purposes it might be advantageous to include coals that contain as much as 1.25 per cent. sulfur, because such coals have been considered as available for making metallurgic coke of fair grade; but the same reasoning might apply to coals containing as much as.1 1/2 per cent. sulfur, if facilities exist for utilizing such coal by admixture with coal containing less sulfur. , In Pennsylvania, it is not possible to define the exact geographic limits of the areas in which coals relatively low in sulfur exist, because variations in the percentage of sulfur occur irregularly and often abruptly. It is not uncommon to find areas, more or less circumscribed in extent, in which the coal is relatively high in sulfur although all the surrounding territory contains low-sulfur coal; and, similarly, small areas of low-sul. fur coal exist in districts where the coal is normally high in sulfur. These generalizations apply, with few exceptions, to the whole area underlaid with bituminous coal in the State of Pennsylvania, and also to each of the individual beds of coal in this state. Pennsylvania is poorly supplied with coals of the low-sulfur class. We find, however, as an offset to this condition, very large areas in which coal ranging from 1 to 2 1/2 per cent. sulfur exists in beds of good workable thickness. This coal, by washing, can be reduced in sulfur low enough to make coke of a high grade. While a considerable tonnage of workable low-sulfur coal remains in the Pennsylvania bituminous districts, it has been necessary to turn to the areas of higher sulfur coal for the supply of metallurgical coke. This utilization of the higher sulfur areas has been made possible either by washing or by selective mining. Washing.—Fine crushing of the entire feed is not generally necessary to the successful washing of Pennsylvania coals. This is due to the fact that the pyrite is rarely finely and uniformly disseminated throughout the coal but is usually confined to high-sulfur benches in the bed or to concentrations on the cleavage planes of the coal. Hence, by graded crushing and sizing it is frequently possible to at once reclaim those portions of the coal bed containing coal that is relatively free from pyrite.

Jan 1, 1920

By J. F. Myers, F. M. Lewis

THE selection of the proper type of flotation machine involves the consideration of a wide variety of factors. Under any condition, all types of machines will promote some kind of separation. Obviously, the correct way of deciding which type of machine is best suited for a particular problem would be to set up all types of machines and test them out in a competitive manner, but the vast number of small concentrating plants are unable to make such complete and exhaustive tests. The experience of the authors of this paper indicates that if all the factors involved in any particular problem can be determined, and if all the flotation-machine characteristics are understood, some basis for selection could be arrived at other than actual comparison. This paper is presented to show how the process factors and machine characteristics have been combined at the mills of the Tennessee Copper Co. from studies of several types of flotation machines. At the two mills of the Tennessee Copper Co. in the Ducktown Basin of Tennessee, four distinct flotation separations are made on the complex sulphide ore. Three different concentrates are produced from the ore-in the order of their importance: iron concentrate, copper concentrate, and zinc concentrate. Four different types of flotation machines are used. The selection of these four machines, for the plant operation, was based upon the study and operating experience of the machine units shown in the following list: Fagergren machine; square cell, size 56 inches Fagergren machine; level type, size 56 inches Denver machine; conical disk impeller, size 24 inches Denver machine; receded disk impeller, size 24 inches Minerals Separation Sub-A, size 24 inches Munro-Pearce deep cell, pneumatic Southwestern air machine, Hunt principle To date the study of flotation machines has been confined almost entirely to the rougher circuits. The various cleaning operations, being relatively a much simpler matter, have not been investigated so thoroughly. Table 1 shows a typical analysis of the two ores as well as of the mineral content. []

Jan 1, 1944

By Stanley D. Michaelson

THE selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. DEVELOPMENT OF TECHNIQUE Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.1 Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a I a by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made, to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1945

By Donald M. Liddell

From a commercial standpoint, the only beryllium mineral warranting attention is beryl, 3Be.Al2O3.6SiO2, which is of fairly widespread occurrence. The chief deposits are in Brazil, Argentina, India, Canada and Portugal. When pure this contains about 14 per cent of beryllium oxide. The mineral richest in beryllium is phenacite, (BeO)2SiO2, or beryllium orth-osilicate, which contains 45.55 per cent beryllium oxide when pure, and if anyone is ever fortunate enough to discover a large deposit of it, it will revolutionize beryllium metallurgy. Gadoljnite, zBe0.Fe0.2Yz0~.zSi0~, is also often referred to as a beryllium mineral, though the occurrences are comparatively scant and it is of more interest as a source of yttrium than it is of beryllium. There has been a good deal of talk recently regarding some occurrences of helvite in the United States. Helvite is a complex manganese-iron beryllium silicate and, apart from difficulties that would arise in its treatment because of its compo-sition, the beryllium content of the known deposits is not high. Early Metallurgy The elemcnt beryllium was first isolated by L. N. Vauquelin in 1797.~ He crushed and heated beryl, having the idea that this heating made it more readily amenable to attack by chemicals, and then mixed the mineral with three times its weight of caustic potash and fused the mixture. The melt after cooling was dissolved in hydrochloric acid: dehydrated, and again taken up with hydrochloric acid and filtered to separate the silica. The filtrate was treated with an excess of potassium carbonate and the precipitate after draining was leached with a solution of caustic potash, which dissolved the alumina. The undissolved material was, in Vauquelin&apos;s own words, "une terre nouvelle." He dissolved the precipitate with nitric acid and evaporated it to dryness, took up with hydrochloric acid and threw the iron out with potassium hydrosulphide, though he found that to make a complete separation of the iron involved a second treatment. The solution, after throwing out of the iron, had a pronounced sweet taste, and though Vauquelin always speaks of working on "terre du beril" he suggests that the new element should be known as "glucine" (glucinum) from the Greek word for sweet, because of this outstanding property. Vauquelin did considerable work on the comparison of aluminum and beryllium, working out a separation of most of the aluminum based on its precipitation as alum. His work is reviewed at some length, since ordinarily his procedure is incorrectly stated. It will also be recognized by those who have studied beryllium metallurgy that a good many technicians have followed his ideas since, including preheating the beryl before treatment. The fusion with

Jan 1, 1944

By Pierre I. Desré, Donald T. Hawkins, Ralph Hultgren

The torsion effusion method has been used to measure the vapor pressure of thallium over pure thallium in the temperature range 931" to 996&apos;k The resulting 100 cal per g-atom is in excellent agreement with previous values. Vapor pressures of T1 above seven Ag-Tl alloys ranging from 33 to 90 at. pct T1 were measured at 979K. Calculated silver activities show a positive deviation from Raoult&apos;s law, approaching ideality near the liquidus. Measured thallium activities scatter less than previous measurements using the electromotive force technique. Values for the silver component were calculated by Gibbs-Duhem integration. This paper reports the results of a study of the liquid Ag-T1 system by the torsion effusion method. The method has been discussed in detail previously1 and will be only briefly summarized here. The sample is contained in a cell which is suspended on a torsion wire. Vapors effusing from orifices on opposite sides of the cell exert a torque on the wire. The torque is related to the vapor pressure by the following relation: where D is the torsion constant of the wire, 4 the angle of torque, fl and f2 the correction factors determined by Freeman and seamy2 to account for orifice length, a1 and a2 the areas of the orifices, and ql and q2 the distances of the orifices from the axis of rotation. By using the same cell for pure thallium and the alloy, it follows that: so that most systematic errors are eliminated in determining the activities. EXPERIMENTAL 99.999+ pct T1 was obtained from the American Smelting and Refining Co. and 99.999 pct Ag was obtained from Cominco American, Inc. Seven Ag-T1 alloys, ranging in composition from x^\ = 0.33 to 0.90, were prepared by melting together weighed amounts of the pure metals in sealed, evacuated Vycor tubes and quenching the melts in water. The apparatus and graphite cells were the same as those used by Hawkins and Hultgren3 in their study of the Pb-Sn system. RESULTS Ten measurements were made on pure thallium in the temperature range 931" to 996K, and the results are presented in Table I. Values of were calculated using tabulated Gibbs (free) energy functions.4 No dependence of on temperature or orifice size was found, indicating that equilibrium was attained. Data for the Ag-T1 alloys are shown in Table 11. Values were smoothed with respect to composition by a plot of where . DISCUSSION The value = 43,520 * 100 cal per g-atom for pure thallium (95 pct uncertainty limit) is in excellent agreement with previous investigations.4 In the Ag-T1 system, this study shows positive deviations from Raoult&apos;s law at all compositions. Fig. l gives the phase diagram according to Hansen, and Fig. 2 presents the activity values measured in this study as well as those from electromotive force measurements at 950°K by Terpilowski, Kundys, and slaby.&apos;

Jan 1, 1969

By P. K. Foster, C. M. Payne, A. McNabb

Some measurements of the rate of loss of hydrogen from cylinders of iron and steel are analyzed in terrns of a trapping theory. The apeement is encozcraging and gives rise to estimates for the density of traps in the metal. Furthermore, the results appear to be consistent with the assumption that the diffusion constant D for the steady flux of hydrogen through iron lies on the extrapolation of the linear curve (log D us 1/T) obtained at higher temperatures. IT has been shown in a previous paper1 that the literature relating to hydrogen diffusion in steels contains sufficient evidence to invalidate the application of Fick&apos;s laws to hydrogen diffusion in steel at low temperatures. To account for the deviations from simple diffusion behavior a trapping theory was developed, and the fundamental laws modified accordingly. In the present paper, new data on the evolution of hydrogen from long cylinders are presented, and together with some published data are examined in terms of the requirements of the theory.&apos; Properties of solutions of the basic equations derived previously1 and appropriate to the present experimental conditions are outlined in the first section as they determine the form in which the experimental data must be presented for a comparison of theory and experiment. Two new solutions not previously reported have been derived as described in the Appendix. The experiments and treatment of data are then given in Sections 2 and 3 and the results are discussed in Section 4. 1) THEORETICAL CONSIDERATIONS The hydrogen-diffusion theory1 applied in this paper postulated that lattice-dissolved hydrogen at a concentration C diffuses randomly in accordance with Fick&apos;s first law, but that the lattice contains a uniform density N of traps which can capture and delay hydrogen atoms. kC atoms are trapped per second in a region containing unit concentration of vacant traps while p are released per second in a region containing unit concentration of occupied traps. On this basis it was shown (Eq. [30] of Ref. 1) that, when local changes in the lattice-dissolved concentration C are slow enough for equilibrium to be established with the traps, the system is represented in dimensionless form by the equation:

Jan 1, 1965

By G. L. Craig

In order to meet the unprecedented demand for copper it has proven necessary not only to develop every primary source, but also thoroughly to exploit the vast reservoir of metal in use that is not directly serving the war effort. To accomplish this, industry has been called upon to use its resourcefulness for developing methods of recovery that have not been previously utilized. It is in this connection that we have been asked to review the methods suggested for recovering the basic constituents of various types of the so-called clad or composite steels. As a background for a discussion of recovery processes, it seems appropriate to consider clad materials in general and their processes of manufacture. A composite steel has been defined as one consisting of two or more component parts of which the smallest shall comprise at least 3 per cent of the total mass. Clad materials or composites are produced commercially by several different methods of which the most important may be classified as: I. The Blacksmith or Forge Weld Method. This method is limited to carbon steels or low-alloy steels as the presence of even small amounts of common alloying metals prevent the formation of a satisfactory weld. The principle is the application of pressure to sections in contact at elevated temperatures. Because of the tendency of the iron to oxidize, borax is generally used as a flux to absorb the iron oxide. 2. Casting Process-In this method a composite ingot, billet or cake is produced by pouring one metal onto or about another. The bond may be accomplished by the action of the molten metal against a solid, in which case the bond may be referred to as mechanical; while in other instances alloying occurs at the interface producing an anchor by the formation of component parts of the two materials. 3. Cementation Process.-In this process any number of combinations of metallic and nonmetallic constituents which are mixed together may be used as a bonding agent. The combinations, whatever they may be, are placed between the materials to be joined and subjected to pressures at elevated temperatures to produce the desired bond. 4. Fusion Welding Processes.-Gas and electric joining for direct use or subsequent fabrication. 5. The, Assembly Process.-This method involves the joining together of prefabricated bars or plates followed by hot and/or cold working the assembly to the desired finished form. This method is advantageous in many cases because of its flexibility on small quantities as compared to the casting method. It is also possible to control the relative proportion of components more accurately. In this discussion we are concerned only with the scrap resulting from the manufacture of clad materials and our interest,

Jan 1, 1943