Search Documents

By W. L. Fink, H. C. Stumpf, L. A. Willey

VERY few studies of the aluminum-sodium system have been reported. Heycock and Neville1 were unable to detect any solubility of sodium in liquid aluminum. Mathewson2 prepared an equilibrium diagram consisting of two horizontal lines at the melting points of the two metals. He was unable to detect solubility at either end of the system. Scheuer3 first developed a method for the chemical determination of sodium in aluminum and then studied the solubility of sodium in liquid aluminum. His method involved heating aluminum with excess sodium under hydrogen until the two liquids were at equilibrium, chilling the melt, and analyzing the aluminum-rich layer. He reported the solubility to be 0.l0 pct sodium at 700°C, 0.115 pct at 750°C, and 0.128 pct at 800°C. This paper on the aluminum-sodium system is the twenty-first in a series from the Aluminum Research Laboratories describing equilibrium relations in aluminum alloy systems. MATERIALS The alloys for this investigation were prepared from electrolytically refined aluminum (99.99 pct) and sodium of reagent quality conforming to specifications of the American Chemical Society.4 Chemical analysis of the aluminum showed the following impurities: 0.002 pct Si, 0.002 pct Fe, 0.002 pct Cu, <0.001 pct Mn, 0.002 pct Mg and 0.001 pct Na. The sodium content of all alloys was determined by either chemical or spectrographic standard procedures5 No. IIA and No. SII respectively. DETERMINATION OF LIQUIDUS The boundary of the miscibility gap was determined by a modification of the method of Scheuer3 and of that used by Henry and Cordiano.6 As finally developed the method was as follows: A fused alumina crucible 25 mm id with 75 mm depth and 1.5 mm wall thickness was placed on the bottom of a container made from a 4-in. length of 1 ½ -in. iron pipe and pipe caps. The space around the sides of the crucible was filled with 40 mesh crystalline alundum. A charge of 60 g of aluminum and 10 to 15 g of sodium was placed in the crucible, the sodium first being trimmed and rinsed with iso-pentane to remove kerosene. The container was tightly capped, filled with argon (99.6 pct pure), and heated to the desired temperature (between 670 and 800°C). A positive pressure of argon (about 8 mm of mercury) was maintained through a small pipe to the upper cap. After soaking for from 2 to 5 hr, the container was quenched by directing a stream of water onto the base, so that solidification of the aluminum-rich layer was rapid and essentially unidirectional. Analyses were made on drillings taken from the bottom of the ingots after machining off the surface layers. Pre-

Jan 1, 1948

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

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

Jan 1, 1944

By F. M. Simonds, E. Z. Burns

The Rawley property is located in the Kerber Creek mining-district, Saguache county, Colorado, at an elevation of 10,600 ft. (See sketch-map, Fig. 1.) Ore was found in this vicinity as early as 1880, and claims covering a large area were located and in many instances patented. The ores, however, being of comparatively low grade, and the nearest railway station, Villa Grove, being some 20 miles away, it is extremely doubtful whether any considerable amount of ore was ever handled at a profit. It is safe to say that several hundred thousand dollars were spent (a large part in building mills) previous to 1902-03. About that time a very capable promoter gathered together a group of these claims and interested some New York people, who had had no previous mining experience, in the purchase of these claims. Following the organization of a company, this promoter became General Manager. As often happens under such conditions, the first step was to erect a mill to treat 100 tons per day. It may not be fair to this enterprise to say that the failure which followed might have been foretold, simply on general principles, but this much is certain—the water supply fluctuates in the course of a season through a wide range and for a considerable portion of the year was inadequate to run this mill at capacity. Further, it is currently reported that some of the mill foundations were placed on ice. However true this statement may be, it is a fact that, when first inspected by us, many of the belts in the mill were held on&apos;their pulleys by means of idlers. A shut-down followed, with a material re-adjustment of interests, at which time we became consulting engineers for the company.

Jan 1, 1914

By D. E. Thomas, C. E. Birchenall

ALTHOUGH Eoltzmann gave a mathematical solution for the diffusion equation (for planar diffusion in infinite 01. semi-infinite systems only) in 1894 allowing for variation of the diffusion coefficient with a change in concentration, it was not until 1933 that this solution was applied to an experimentally investigated metallic system. The calculation was carried out by Matano&apos; on the data obtained by Grube and Jedele3 for the Cu-Ni system. Since that time concentration dependence of the diffusion coefficient has been demonstrated for many pairs of metals. However, the nature of this dependence has never been fully elucidated. Many investigators have suspected that these variations could be related to the thermodynamic properties of the solutions, one of the earliest explicit statements being contained in a discussion of irreversible transport processes by Onsager&apos; in 1931. Development along these lines has been greatly retarded by the lack of reliable data on the variation of tliffusivity with concentration, the paucity of the thermodynamic data for the same systems at the same temperatures and compositions, and an incomplete understanding of the relation of the thermodynamic properties of the activated state for diffusion to the bulk thermodynamic properties. The last factor has been discussed by Fisher, Hollomon, and Turnbull.5 In many instances where data exist, it is difficult to know which are acceptable. This problem probably applies more strongly to diffusion data than to activity measurements. For instance, four sets of observers"-" have reported self-diffusion coefficients for copper. The average spread between extreme results is a factor of about four, though the individual sets of data are self-consistent to about 20 pct. Thus one or more factors are out of control, at least in these experiments, making estimates of internal error unreliable. The most reliable diffusion data in most systems have resulted from the use of welded couples with a plane interface from which layers for analysis are machined parallel to the interface after diffusion. The layers are analyzed, and the result is a graphical relation between distance and concentration, usually called the penetration curve. Given the same set of analytical data and distances and following the same procedure in computation, different observers will generally produce diffusion coefficients which vary appreciably, especially at the extremes of the concentration range. Experiments must be carefully designed so that the precision is good enough to answer a particular question unequivocally. In the first calculation of the dependence of the diffusion coefficient on concentration in the metallic solid solution Cu-Ni, Matano found that the coefficient was insensitive to concentration from 0 to 70 pct Cu, after which it rose more and more steeply to some undetermined value as pure copper was approached.&apos; The same behavior was reported for Au-Ni, Au-Pd, and Au-Pt.* The data used were those of Grube and Jedele which were very good at the time, but are not considered particularly good by present standards. Furthermore, the method of calculation makes the ends of the diffusion coefficient-concentration curve unreliable. For better reliability, the high copper end of the curve has been checked by incremental couples, where the concentration spread is 67.7 to 100 atomic pct Cu. The implication of the curves calculated by Matano was that diffusion is very concentration sensitive in one dilute range of this completely isomorphous system and hardly at all in the other. Matano&apos;s result is confirmed. Later Wells and Mehll0 published data on diffusion in Fe-Ni at 1300°C, which represent a thorough test of the shape of the concentration dependence curve. They ran couples with the following ranges of nickel concentration: 0-25 pct, 1.9-20.1 pct, 0-20.1 pct, 20.1-41.8 pct, 0-99.4 pct, and 79.3-99.4 pct. Although the trend of the data indicates an S-shaped concentration dependence, their curve was drawn to the pattern set by Matano. Their original data have been recalculated for the 0-99.4 and 79.3-99.4 pct couples. Wells and Mehl&apos;s points and two independent recalculations from the raw data are plotted in Fig. 1. What appears to be the best curve is drawn through them. This curve shows little sensitivity to composition in both dilute ranges with a strong dependence at intermediate composi-tions.? Similar experiments on the Cu-Pd system are reported here at temperatures where solubility is unlimited. These lead to the same type of concentration dependence for the diffusion coefficients as was found upon recalculation of the data for the Fe-Ni system. Experimental Procedure Cu-Pd: The concentration dependence of the diffusion coefficient may be determined by the use of

Jan 1, 1953

By Oscar E. Harder

THE TERM "physical metallurgy&apos; is used in the title of this lecture in preference to "metallography ?because the former has a broader meaning with most audiences, some people thinking of the latter as being limited to work with the microscope. This distinction, how- ever, has not been held by the out- standing workers in the field of metallography, such as the late Professor Sauveur. In the broadest sense metallography is essentially the same as physical metallurgy. The American Society for Testing Materials has adopted-the following definition: "Metallography is that branch of science which relates to the constitution and structure, and their relation to the properties of metals and alloys." This definition might be improved by adding after "properties" -"and their industrial and scientific uses." Metallography has also been defined as the science of metals and alloys. Physical metallurgy is to be distinguished from process metallurgy in that the latter relates particularly to the production of metals from the ores.

Jan 1, 1940

By M. A. Whiting, H. Kenyon Burch

One of the advantages presented by electric drive in many classes of work is the ease with which the electric motor can be controlled automatically. In a large number of cases certain features of the control are automatic—for example, the rate of acceleration may be limited automatically or the equipment may be stopped automatically at the limit of travel—but the equipment is started and ordinarily is stopped by hand. In other cases the motion of the machinery is utilized to start, control the speed, and stop the motor automatically, independently of any operator. A considerable proportion of the large mine hoists now in use have certain automatic features, particularly protective devices against overwinding, and, in some classes of electric hoists, devices for preventing excessive acceleration or retardation. The large automatic hoists discussed in this paper, however, are completely automatic, i.e., capable of making their trips without the presence of an operator at the control levers. According to circumstances, various advantages may be obtained . by automatic control, chief of which are decreased power consumption, increased precision and safety of operation, and decreased cost of attendance. The first step in the analysis of a prospective automatic mine hoist is to determine whether automatic operation is feasible at all. If men are to be hoisted, or levels changed, the attention of an operator is required for these purposes; but under some conditions it may be entirely practicable and advantageous to build the equipment so that, while provision is made for hoisting men or changing levels, ore can be hoisted automatically from any one level. If, however, an operator&apos;s attention is required every few minutes for changing levels, handling men or drills, or other work requiring hand control, it is obvious that automatic operation between times will not be of any practical benefit.

Jan 1, 1917

By B. M. Larsen, W. E. Shenk

This paper describes a method of reversal control of the open-hearth furnace that obtains in practice those effects considered below as essential to a completely automatic control, without appreciable interference with the natural rise and fall in temperature of the regenerative portions of the furnace system. Growing out of some studies of radiation pyrometry in open-hearth regenerators, it was developed during 1934-1938, and has been successfully applied to a number of commercial furnaces. The goal of open-hearth furnace operation is the highest rate of production of steel of proper quality consistent with good refractory life and fuel economy. This requires rapid heat transfer from furnace to charge, which in turn is dependent upon heating the furnace refractories to the highest safe operating temperature. The greater the heat input to the melting chamber, the sooner this limiting temperature will be reached, so that control methods in general that are designed to protect against overheating of refractories should help to maintain maximum fuel input and production rate. For the typical open-hearth system, with fixed regenerator capacity and with rate of fuel input usually restricted in its range of variation, our basic assumption is the idea that the only other important variable —that is, the time interval between successive reversals—should also be held constant or within a small range. The present reversal control was designed therefore so that, under normal operating conditions, no arbitrary limitation is imposed other than to keep the reversal interval within the optimum time range. However, since a near approach to completely automatic operation involves protection of refractories through the balancing of furnace end temperatures as well as emergency protection against temperature peaks above some upper limit, it was also felt essential to employ, as an integral part of the control unit, a temperature-measuring instrument to register the true temperature of the refractory surfaces at some critical region in the hot ends of the regenerative zones. As the basic elements of the control, we thus have one "maximum" and one lLminimum" time relay and a potentiometer recorder-controller, combined with switches and relays to operate as follows: I. Reversal normally occurs within an optimum range limited by the maximum and minimum settings to about 4 to 6 min. variation, but the actual interval usually is less variable and closer to the minimum setting. 2. The margin between minimum and maximum time is utilized, together with a "floating temperature limit" in the potenti-ometric control instrument, to correct any tendency toward temperature unbalance between ends of the furnace. 3. A second "maximum temperature

Jan 1, 1945

By Stephen Krickovic

THERE is today no proven continuous mining machine that can be used under all the varying conditions found in most bituminous coal mines. During the last five years, however, both the machines and methods of using them have been improved. Coal operators can now exercise a fair degree of selectivity, except those who must struggle with thick and hard partings and operate in seams less than 40 in. thick. It must be realized, then, that more than one type of continuous miner may be used advantageously in the same coal mine. Although this appears to violate the standardization generally sought after, the gain can more than offset added maintenance and, to a lesser degree, operating problems. It is equally important to realize that most coal mines will require one or more conventional loader setups in the plan of total mechanization. Seam variations, adverse physical conditions, and special designs of main headings for haulage and ventilation, to name only three reasons, may render 100 pct use of continuous miners uneconomical. Of course, this situation may change. In one mine, prior to 1954, machines that ripped coal from the face achieved only average results, as compared to the impressive performance of boring-type units in another coal mine. A boring-type unit was therefore selected for the first mine in the belief that the chance of success would be greater if it were used in a combined cutting and wedging operation, cutting a full face in its forward movement. The miner was located in a part of the mine where a special projection could be readily developed, a solid area scheduled for mining according to the overall progress plan. All openings surrounding the particular area were carefully investigated for the presence of sulfur balls. Fig. 1 (opposite) shows the representative seam section, containing the several pyritic bands 1/4 to 1/2 in. thick; a characteristic Pittsburgh seam parting averaging 5 to 6 in. of slate, coal, and slate; and the occasional py-rite lenses 2 to 3 in. thick. Under normal conventional mining methods, the top 10 or 12 in. and the bottom 4 in. are left in place to offer roof protection

Jan 1, 1958

By M. S. Burton, G. V. Smith, A. Rosen

The kinetics of re crystallization and the effect of recovery on recrystallization of pure iron were investigated within the temperature range of 517" to 632 OC. Grain growth and activation energies were also studied. Two stages can be distinguished during recrystallization; the first is gvowth-controlled and the second is nucleation-controlled. The extent to which the first stage succeeds before the second stage begins depends on both temperature of recrystallization and temperature of recovery. The softening of cold-worked metals and alloys due to transformation of the deformed crystals to strain-free equiaxed grains during annealing is well-known, and early inestiators established that recrystallization kinetics is strongly influenced by minute quantities of impurities. Generally, it is believed that impurities decrease the rate of recrystallization and the rate of grain growth. To understand the basic phenomena of recrystallization, investigations should be on pure materials. Most of the research on high-purity materials has been on metals having the fee Bcc materials, and especially high-purity iron, have been studied only in recent years. It is known that recovery may influence recrystalliation;" however, no general description of the influence, applicable to all materials, has been presented. This paper presents results of a study of the recrystallization kinetics of high-purity iron, after 60 pct cold deformation by drawing, with particular reference to the influence of recovery on the rates of recrystallization and grain growth. From observations of these rates, conclusions are drawn concerning nucleation. MATERIAL AND EXPERIMENTAL PROCEDURE Preparation of Material. High-purity iron, prepared by zone refining of vacuum-melted electroiytic iron at Battelle Memorial Institute under sponsorship of the American Iron and Steel Institute, was used in this study. Table I gives the reported chemical analysis of the material, taken at a point near the last molten zone and thereby rep- resenting the maximum impurity level for the entire bar. Prismatic specimens, approximately 10 x 10 X 50 mm, were sawed from the 50-mm-diam bar, perpendicular to the zone-refining direction, and turned to cylindrical shape. These specimens were sealed in quartz capsules under vacuum, approximately 10"e mm Hg, austenitized at 930° C for 15 min, slowly cooled, removed from the capsules, and cold-worked in air by swaging to approximately 20-pct reduction of area. The encapsulating, austenitizing, and swaging operations were repeated two additional times, after which the samples were encapsulated, austenitized, and cold-drawn to 60-pct reduction of area. A further vacuum anneal at 630°C for 120 min and cold drawing to 60-pct reduction of area produced wire 2.8 mm in diameter. The drawn wire was electropolished to remove deformed material on the surface, after which it was sawed into 4-mm-long specimens for the recrystallization study. Microscopical examination before the final cold reduction revealed equiaxed grain structure with very fine grains in a thin layer near the surfaces of the specimen and nearly uniform ASTM 3-4 grain size in the center. Recrystallization Heat Treatment. a) Direct Recrystallization of Cold-Worked The 60 pct cold drawn iron specimens were recrystallized at five temperatures, 517, 556, 590°, 612, and 632C, in a lead bath covered with graphite (the temperature was controlled to 1°C). Specimens were treated for twelve different times at each temperature, and all were water-quenched after heat &eatment. b) Recrystallization Following Preliminary Recovery Heat Treatment. After cold working, specimens were sealed into quartz capsules under vacuum, 105 mm Hg, and recovery-annealed at 362", 400", and 440°C for 20 hr prior to recrystallization. No evidence of recrystallization was found

Jan 1, 1964

By C. A. Heiland W. E. Pugh

The principle underlying the majority of magnetic intensity variometers is a comparison of the force to be measured with another force of known magnitude. The known force may be (a) of a magnetic nature, such as magnetic fields produced by magnets or coils, or (b) of a mechanical nature, such as (1) the elastic forces of wires, or (2) the force of gravity, such as employed in bifilar magnetometers and magnetic balances. In instruments of such type, usually a magnetic system is employed, free to move about either a horizontal or vertical axis, on some sort of a suspension. If the earth&apos;s magnetic field alone were present, such a system would adjust itself into the direction of the field or that of its components. Owing to the simultaneous action of the comparison force, however, the magnetic needle assumes a position that differs from the direction of the magnetic force; for all such instruments, the angle of deflection is given by tan = FIX, where F is the known and X the unknown earth magnetic force component. In other words, the magnetic system merely serves as an "indicator" for the position of balance of the two forces. Considering the matter from this point of view, it is not immediately obvious why the temperature of such a magnetic system should modify the results. In magnetic instruments, we may distinguish two different effects of temperature: (1) a magnetic effect and (2) a mechanical effect. The first is due to the fact that the magnetic moment of a magnet drops with an increase in temperature. This drop is characterized by the temperature coefficient of the magnetic moment (see equation 1). The second effect is due to the change in the torsion of suspension wires with temperature or to the expansion of lever arms of masses acting on magnetic balances.

Jan 1, 1934

By H. A. Guess

Introduction. History and Production Statements. Southeast Missouri is the oldest of the large producing districts of the United States. The first recorded production from disseminated ores was in 1869, although for many years prior to that time lead had been milled from shallow pockets, and steady growth in production has been shown ever since, the greatest increases being during the past five years. The production by years for each of the operating companies from 1869 to 1907, inclusive, is given in Table I. In Table 11. is given the production by years from 1908 to 1912, inclusive, for the district as a whole, showing the tonnage of crude ore mined, yield in concentrates, average assay of concentrates; yield in metallic lead both per ton of crude ore and in total; also approximate value of the yearly production. It is interesting to note that the production for the 5-yr. period 1908 to 1912, inclusive, is practically equal to the total previous production, viz., the 39-yr. period 1869 to 1907, the summary being as follows:

Jan 1, 1915

By Eugene N. Cameron

Graphite is the hexagonal form of crystal-line carbon. It is found in nature locally as tabular crystals but occurs mostly as disseminated flakes, foliated, platy, or fibrous masses, or microcrystalline compact to earthy material. Graphite is characterized by its gray to black color, metallic luster, black streak, perfect basal cleavage, softness (H = 1-2) and unctuous feel. It is sectile and flexible and has a specific gravity of 2.1 to 2.3. It is an excellent conductor of heat and electricity and is inert to ordinary chemical reagents. It melts at a temperature of about 3500°C and vaporizes at a temperature of about 4500°C. In the presence of oxygen it oxidizes to CO2 in the temperature range 600° to 700°C, but it is stable at ordinary temperature and markedly resistant to chemical weathering. Although some natural graphite is nearly 100 pct pure carbon, most contains mechanical impurities of various kinds; quartz, biotite, muscovite, pyrite, iron oxides, and feldspars are the most common. Commercial graphite products have a wide range of graphite content, but the better grades contain 80 pct graphite or more. Both natural and manufactured (artificial) graphite are used in industry. Natural graphite products are divided into two broad categories based on grain size. Graphite composed of visible crystals is called "crystalline" graphite, whereas graphite so fine-grained that it is not visibly crystalline is termed "amorphous." The distinction is arbitrary, for all natural graphite is crystalline, and all gradations between the two commercial categories are known. Within the two categories the terminology used in industry is confusing and not entirely consistent. Crystalline graphite comes from two principal types of sources. The first consists of rocks containing disseminated flakes of graphite. This source yields the "flake graphite" of industry. The second comprises vein deposits of graphite, of which those in Ceylon are the most important and furnish the Ceylon graphite of industry. Amorphous graphite comes from any of a variety of sources, including some of those that yield crystalline graphite. Manufactured graphite is made by electric furnace processes, mostly from petroleum coke. Occurrence and Distribution of Graphite Deposits GENERAL Graphite is widely distributed in nature, occurring in igneous, sedimentary, and metamorphic rocks, and even in nickel-iron meteorites. Graphite deposits of economic interest, however, occur mostly in metamorphic terranes. Five principal geologic types of deposits are recognized: 1. Deposits consisting of flake graphite disseminated in metamorphosed siliceous sediments. 2. Deposits consisting of flake graphite disseminated in marbles. 3. Deposits formed by thermal or dynamothermal metamorphism of coal or other highly carbonaceous sediments. 4. Vein deposits of the Ceylon type. 5. Contact metasomatic or hydrothermal deposits in marble. DEPOSITS IN METAMORPHOSED SILICEOUS SEDIMENTS Deposits consisting of flake graphite disseminated in metamorphosed siliceous sedi-

Jan 1, 1960

By Walter Nummela

INTRODUCTION By the end of 1979 pellet productive capacity from the processing of low grade iron ores in the Lake Superior-District will total 86 million tons per year. Of that tonnage, 72 million result from the processing of magnetic ores such as taconite, about 6 million tons from the flotation processing of relatively coarse grained specular hematite, and 8 million tons from the processing of fine grained non- magnetic ores at Tilden. The Tilden flowsheet represents a break- through in the processing of this type of crude ore and hence is unique in the Lake Superior District (1). Operations at Tilden started in late 1974 with Algoma Steel, J&L Steel Corp., The Cleve- land-Cliffs Iron Company, The Steel Company of Canada, Ltd., Wheeling- Pittsburgh Steel Corp., and Sharon Steel Corp. as participants. Cleveland-Cliffs operates and manages the operation. Other companies have conducted extensive pilot plant tests or are in commercial production in the processing of fine grained non-magnetic ore materials (2). Flowsheets and reagent schedules may be different for each particular application, but some of the problem areas in the processing of Tilden ores have also been experienced at operations such as the IOC Plant at Sept-Isles, at Lamco and at Bong in Liberia, and in pilot testing of Fabrica ores of Ferteco Mineracao in Brazil. With a grind of about 80-85% -500 mesh resulting in 50% of the particles in the concentrate being less than 15 microns, a discussion of the pro- cessing of Tilden type ores fits the topic of this workshop. There are very large tonnages of fine grained non-magnetic crude ores which exist throughout the Lake Superior District and other iron ore producing regions. These fine grained non-magnetic materials constitute an important source of iron minerals. Before going into specifics of the flowsheet, it is important to

Jan 1, 1979

By B. W. Gandrud

WET-PROCESS coal-washing tables as we know them today have been in use in this country for approximately 25 years. The literature records only a few table installations worthy of note prior to adoption of the present-day differential-motion table. According to Phillips,&apos; six Campbell bumping tables were being operated by the St. Bernard Coal Co., Earlington, Ky., in 1893. As late as 1923, according to Richardson, 72 such tables were in operation in the Rosedale washery of the Cambria Steel Co., at Johnstown, Pa. Bumping tables probably were not used to any considerable extent except in the plant at Earlington, and at two or three plants in Pennsylvania, but they may be considered as forerunners of the present-day coal-washing table. The transition from bumping tables to the type of tables now in use signalized a revolutionary change iii table design. The advantages of the new design over the old were so outstanding as to make its adoption inevitable. BUMPING TABLES Bumping tables, as the name implies, utilized a bumping action for conveying the refuse and discharging it over the refuse end. The tables were hung from rods attached to overhead supporting beams. A shaft and cam arrangement at one end of the table served as ,the actuating mechanism and imparted to the table an endwise, back-and-forth, swinging motion. A bumping block was fixed in the proper position at the shaft end, so that the table would strike against it at the end of each backward swing. The hanger rods or stirrups were so adjusted as to give the table a slight uphill slope toward the refuse end; that is, the end next to the driving mechanism. The feed box was near the upper end of the table and designed to spread the mixture of coal and water over the width of the table. Wash water was added a short distance below the feed box. With the table in motion and coal mixed with water being fed to it through the feed box, a certain amount of stratification took place, depositing heavy refuse at the bottom of the bed in contact with the surface of the table. Constant bumping and a certain amount of

Jan 1, 1943

By Rudolph G. Wuerker

Charles T. Holland (Head, Dept. of Mining Engineeri*, Virginia Polytechnical Inst., Blacksburg, Va.) Mr. Wuerker has presented a very interesting discussion of the use of triaxial test methods for investigating the strength properties of rocks. Such methods, no doubt, eventually will develop considerable information of interest to those concerned with the design of mine layouts, particularly in the field of pillar design. From his discussion of my recent article, "Cause and Occurrence of Coal Mine Bumps" (Holland Mining Engineering 1958, p. 933-1002), it is evident that in one place at least I did not make my meaning clear to him and perhaps others. To clear the matter up I think it best to quote from the article, somewhat more fully than did Mr. Wuerker, as follows: "4) In actual operations — because rocksare not perfectly elastic, homogeneous, nor isotropic and because local yield does occur — the maximum stress as demonstrated by Phillips (Ref. 22, pp. 64, 65) and indicated by much experience in mining, does not occur at the walls of the opening but at a short distance inside the pillar. Furthermore, the maximum stress does not reach as great a value as theoretical considerations and laboratory experimental methods indicate.* Actual distance inside the pillar, measured from the wall, at which the maximum stress exists, has not been determined. Observations in many mines, however, indicate that this distance could have a mini-value of one to six or eight times the bed thickness and that it is probably affected by width and height of the opening, depth of cover, and relative values of the elasticity and plasticity of materials comprising the roof, floor, and coal seam. The actual value of the stress produced probably lies between the theoretical maximum and the average stress concentration that would be produced if the weight of the strata above the unsupported opening were evenly distributed over the pillars for a distance equal to the opening width." The footnote reference in the above quotation referred to the following: "*For example, the Pocahontas No. 4 coal bed in southern West Virginia is mined under cover up to 1800 ft thick. Development openings are driven 18 to 20 ft wide, and the bed is about 6 ft thick. According to the work of Panek, the tangential wall stress at mid-bed height under these conditions would reach values between 4000 and 5000 psi. Actual tests of 3-in. cubes of this coal show its compressive strength would be much less than this, perhaps as low as 400 psi. Yet the pillars usually show no evidence of failure in these headings. In this same bed at a depth of 800 ft, the author has seen an opening 225 ft between supports lying between two old groves approximately 1100 ft apart. According to the theoretical considerations, the stress in the pillar walls would have been about 18,000 psi, yet the pillar showed little or no evidence of weight. In view of these observations, it is clear that the wall stress does not attain the maximum values indicated by theory." (Underlining added to original wording.) By referring to Fig. 2A of my paper it will be noted that theoretically the maximum pillar stress would occur at the pillar wall, i.e., at the passageway surface of the pillar. Obviously this cannot be correct in the cases of stress ranging from 4000 to 18000 psi since the coal at the surface of the pillar is under no constraint and cannot have a strength much greater than 400 or 500 psi. Hence, my conclusion that the maximum stress does not occur at the wall but back in the pillar some distance from the wall. Since these stresses are pushed back in the pillar from the wall, it is also obvious that the loads transferred to the pillar from the opening will be spread over a greater area and hence Pillar stresses will not rise to the values postulated by theory and photoelastic experiment. Further since to visual inspection the coal along the pillar wall did not appear to be failed the conclusion was reached that the stress shift was caused by local elastic or plastic yield and by difference in the elastic modulus of the rocks composing the mine floor, mine roof, and coal bed. Later on under the heading "Strength of Mine Pillars" (pages 1000-1002) the effects of constraint is briefly described. Also a formula taking into account constraint is developed relating pillar strength to the uniaxial strength of coal and the L/T ratio of the pillar. Since my paper was written, reports of experiments conducted in South Africa (Denkhaus, et. al., 1959), in Sweden (Hast 19581, and in Canada (McInnes, et.al., 1959) reveal that the conclusion expressed relative to the existence of a low stress area existing around the edges of pillars and solid faces as described above is generally correct. But it seems possible that where the wall stress developed is less than the unconfined strength of the rock composing the pillar and where the roof, floor, and pillar

Jan 1, 1961

By H. M. Otte, A. L. Esquivel

A new leasl-squares method is Presented for determining lattice parameters of hexagonal or tetragonul structures. The method is adapted for use on electronic computers and involves a reiterative procedure. The correction factor employed raries linearly with the lattice parameter, a (determined from the Brag, angle). In contrast, Cohen&apos;s method and recent modifications of it use a correction factor that varies incersely as the squure of the lattice paramneter, a. While the recent modifications attempt to improve the precision of the extrapolated lattice parameter, a, (or-cu). by stressing the importance of the weighting factor. the present approach emphasizes the need for choosing the correct extrapolation function. A comparison between the present method (the Linear method) and Cohen&apos;s method indicates that the Linear method may be more appropriale in certain cases. through a priori no critertion appears to he available for making a chorce between the methods. The size of hexagonal and tetragonal crystal lattices is determined by two parameters, a, and c,. Although in principle the two lattice parameters can be determined independently from reflections for which h - k = 0 and 1 = 0, respectively, in practice this may be inconvenient (because of the angular positions at which these reflections may occur) or not easily possible (because of low intensity). Furthermore, if a high precision or accuracy is required, the limited number of reflections of this type available, particularly in the high-angle region, is not sufficient for the necessary corrections (mainly due to absorption) to be determined with accuracy. Several methods1-7,11-13 have been proposed employing all reflections, to obtain the optimum values, a. and co, either by a trial and error procedure or a least-squares fit. Of the latter method, cohen&apos;s5 is the best known one since it provides explicit expressions for the optimum a. and c,. However, Cohen&apos;s method is only strictly valid if an extrapolation function is used that varies linearly with l/a2 or lie2 (see Section 3), a requirement that does not appear to be generally appreciated. On the other hand, all the better known and more widely used A recent trial and error method was proposed by Massalski and King8,7 who computed extensive auxiliary tables of axial ratios vs the functions A = [(4/3)(h2 + hk +k2) + l2+(a/c)2] and C = A(c/a)2 used in computing a and c values from the observed Bragg angles. These values of a and c were then plotted against a function which permitted linear extrapolation. As a criteria; for the "optimum" values, Massalski and King rely upon a visual fitting of the line through points representing reflections of low 1-index points to compute the extrapolated value of ao and high 1-index reflections to obtain CO. The successive computations and graphical plotting required to reach the "optimum" value are quite lengthy and tedious even on a desk calculator and no quantitative assurance is obtained of having in fact selected the optimum value.* If the method of Massalski and King is used on an electronic computer, then their published tables become redundant and a least-squares fit becomes a natural selection for the choice of optimum values. Such an approach will be called the Linear method. For work now in progress on the effect of certain physical variables on the lattice parameters of hexagonal crystals, it has become essential to determine the confidence limits of small changes in the lattice parameters. Since extrapolation functions that varied linearly with a and c actually also appeared to vary linearly with l/a2 or l/c2 when tested against published as well as our data, a comparison of Cohen&apos;s method and the Linear method was considered desirable (Sections 3 and 4). For the latter method an electronic computer was required since a reiterative procedure to obtain the optimum ao and co values had to be employed. The purpose of this paper is to describe the principles of the Linear method, illustrate its application, and compare it with Cohen&apos;s method. 1) THE LINEAR METHOD The standard practice in obtaining the lattice parameters in the Debye-Scherrer method is to

Jan 1, 1965

By W. F. Hower, W. Brown

Large-scale sand consolidation tests were conducted in an effort to determine the reasons for the successes and failures of this method of sand control. Several different consolidating materials were used in treating both clean and bentonitic sands that were packed in a chamber having a capacity of 3.3 cu ft. The results were essentially the same for all of the different consolidating materials, The data show that low-viscosity consolidating materials pumped at a relatively slow rate gave the best results. Where the formation has produced sand, the treating fluids can compress the formation, thus permitting the channeling of fluids to another horizon. Pressure-packing these zones before attempting to consolidate is recommended. Sands containing more than 4 per cent of water-swelling clays are not good candidates for consolidation. It is indicated that loose sand, particularly when it is bentonitic, can be fractured during the placement of the treating fluids. INTRODUCTION Sand production in oil and gas wells has plagued the industry for many years, and numerous cures for this problem have been suggested. Most methods have been successful to a certain degree, but the great variety of well conditions that exist in the different areas has magni- fied the problem and limited the successful use of the various systems. Four review papers1-4 present a wealth of information concerning the degrees of success that have been obtained by the different sand-control methods. The bridging of sand grains by the use of gravel packs and screens has been quite successful. However, these methods do not leave the casing clear for all types of multiple completions, and the cure does not last for the production life of the well in some instance:;. The control of loose sands by sand consolidation with resins has never been as successful as desired. It has always been hoped that such a treatment would eliminate all sand problems for the life of the well, but. initial applications, starting in the middle 1940&apos;s, were only moderately successful. Lott, et a1,3 reported a success ratio of approximately 50 per cent and made the following conclusions. The highest percentage of successes were obtained where: a. Consolidation of a zone was made at the time of initial completion or prior to the production of sand. b. The interval treated was less than 12 ft in length. c. Between 30 and 50 gill plastic/ft of producing interval was displaced through the perforations. REASONS FOR SAND CONSOLIDATION FAILURES Our own experiences in the field of sand consolidation point toward the following conditions as the major reasons for the failure of sand consolidation attempts. 1. Mud-plugged perforations and mud invasion of the formation. 2. Sand in the casing covering all or part of the perforations. This sand could be either formation sand or one of the coarser sands used as propping agents in hydraulic fracturing. 3. Holes in the casing. 4. Channels behind the casing. 5. Attempting to treat too long a perforated section. 6. Too high a percentage of water-swelling clays in the formation. 7. Formations that have produced sand. Recent attempts were made to treat perforated sections ranging from 10 to 30 ft, in wells that have produced sand, by using a straddle packer that was raised and lowered through the perforations as the consolidating material was being pumped. In most instances, the pressure required to pump fluid into the formation varied considerably as the tool was raised and lowered. This suggested the possibility that significant differences in permeability were present or that only part of the formation had produced sand. There were times when a sudden break in pressure indicated that a fracture was being formed. Research conducted several years ago concerning the problem of the control of water in air and gas drilling indicated that shale sections could be fractured quite easily. In addition, it was determined that it was easier to pump fluids into shale bodies by fracturing the shale itself, or the interface between the shale and sand, than to pump into a fluid-saturated formation. Formations that produce sand are usually adjacent to shale bodies and frequently have shale streaks of various thicknesses inter-bedded in the sand. Therefore, where shale is exposed to fluid pressure it

By H. Leslie Bullock

From aluminum to zirconium, the quantitative preponderance of quartz as a gangue material is well recognized. lf this material is to be efficiently removed, its variations must be understood. Variations with temperature are especially important. Too little attention has been given to the thermal polarization of quartz. Under closely controlled conditions, electrostatic upgrading is very reliable. For efficient separation, contact charges must be fostered and charges due to radiation, surface coatings, or thermal polarization avoided. This paper lists thermal transition points of quartz and shows their effect on actual separations. Simple separation tests with all factors except temperature held constant are recommended for determining transition points. With silica making up more than 27% of the earth&apos;s crust, its oxides comprising more than 59% of all igneous rocks, and quartz accounting for most of the main free oxides, the mining engineer is in constant contact with quartz, which may occur as a valuable mineral to be purified or, far more frequently, as a gangue material to be removed as completely and economically as possible. As a means of effecting such purification or removal, dry beneficiation is becoming more and more desirable owing to local water scarcities, wet waste disposal problems, or freezing conditions. One method that has been gaining particularly rapid acceptance is electrostatic beneficiation, or the separation of dry free-flowing materials by means of opposite surface charges, differences in potential, or differences in conductivity. Electrostatic beneficiation dates back to the 1870&apos;s, but only in recent years have newly developed methods and apparatus and a growing knowledge of solid-state physics widened the field for its economical application. Because the term "electrostatic beneficiation" has been rather loosely used in the literature, it has come to include both electrostatic and electrodynamic procedures. Attracting type separators, however, in which oper- ation is based on differences of surface charge or potential, are truly electrostatic, because the separation occurs in a substantially static field set up between oppositely charged surfaces. Separations with this type of equipment may occur at potentials as low as 1000 v and seldom require potentials as high as 30,000 v. In the new contact charge dielectric separators1 the variation in charge is produced by continuous contact and separation of the particles in the moving feed stream and these charges are fostered by the use of non-conducting support and feed surfaces and by handling the feed in the form of streams of appreciable depth. This favors uniformity of feed and allows higher production rates. The distinctive surface charge differences are set up on the separation of the particles according to Coehn&apos;s Law,2 which states that equal and opposite charges are generated on the separation of any two materials in contact and that the substance having the highest dielectric constant will be positively charged. The basic contact charge concept is reliable, but all electrical charges are transient and modified by the electrical conditions of the surroundings. Pyro-electric, photoelectric and radiant effects may modify or totally destroy the contact charges necessary for efficient separation, or contact with conducting surfaces may neutralize them. Such hostile conditions must be carefully guarded against, since they have led to many costly failures in the past. The most consistent difference in surface charges to insure good separation is produced by repeated uniform contact and separation of particles in the moving stream. The thickness of the feed stream possible with this method reduces the effect of contact with the supporting surfaces, but as some contact is inevitable, the best results may be assured by having the dielectric constants of the supporting surfaces between the dielectric constants of the substances to be separated. In general, the hard smooth surface of quartz makes it an ideal substance for electrostatic separation from most minerals. For instance, with calcite, starting with a feed containing 1.9% acid insolubles, one can produce a concentrate containing 0.30% acid insolubles with a tailings containing 21.7% acid insolubles and a yield of 92.8%. The color can be held at 92 or above and the tint at 1.7 or lower. Working with specularite iron ore, laboratory work has given a concentrate of 68.8% Fe., with an iron unit recovery of 96.7%. Excellent results are also in pros-

Jan 1, 1969

By F. C. Bond

SIZE of grinding media is one of the principal factors affecting efficiency and capacity of tumbling-type grinding mills. It is best determined for any particular installation by lengthy plant tests with carefully kept records. However, a method of calculating the proper sizes, based on correct theoretical principles and tested by experience, can be very helpful, both for new installations and for guiding existing operations. As a general principle, the proper size of the make-up grinding balls added to an operating mill is the size that will just break the largest feed particles. If the balls are too large the number of breaking contacts will be reduced and grinding capacity will suffer. Moreover, the amount of extreme fines produced by each contact will be increased, and size distribution of the ground product may be adversely affected. If the balls added are too small, grinding efficiency is decreased by wasted contacts that are too weak to break the particles nipped; these largest particles are gradually worn down in the mill by the progressive breakage of corners and edges. Ball rationing is the regular addition of make-up balls of more than one size. The largest balls added are aimed at the largest and hardest particles. However, the contacts are governed entirely by chance, and the probability of inefficient contacts of large balls with small particles, and of small balls with large particles, is as great as the desired contact of large balls with large particles. Ball rationing should be considered an adjunct or secondary modification of the principle of selecting the make-up ball size to break the largest particle present. Empirical Equation In 19521,2 the author presented the following emerical equation for the make-up ball size: B - ball, rod, or pebble diameter in inches. F = size in microns 80 pct of new feed passes. Wi - work index at the feed size F. Cs - percentage of mill critical speed. S — specific gravity of material being ground. D == mill diameter in feet inside liners. K - 200 for balls, 300 for rods, 100 for silica pebbles. Eq. 1 was derived by selecting the factors that apparently should influence make-up ball size selection and by considering plant experience with each factor. Even though Eq. 1 is completely empirical, it has been generally successful in selecting the proper size of make-up balls for specific operations. But an equation based on theoretical considerations should be used with more confidence and have wider application. The theoretical influence of each of the governing factors listed under Eq. 1 was accordingly considered in detail, as described below, and a theoretical equation for make-up ball sizes was derived. Derivation of Theoretical Equation Ball Size and Feed Size: The basis of this analysis is that the largest ball in a mill should be just sufficient to break the largest feed particle into several pieces, excluding occasional pieces of tramp oversize. In this article the size F which 80 pct passes is considered the criterion of the effective maximum particle feed size. The smallest dimension of the largest particles present controls their breaking strength. This dimension is approximately equal to F. As a starting point for the analysis it is assumed that a 1-in. steel ball will effectively grind material with 80 pct passing 1 mm, or with F- 1000µ or about 16 mesh. The breaking force exerted by a ball varies with its weight, or as the cube of its diameter R. The force in pounds per square inch required to break a particle varies as its cross-sectional area, or as its diameter squared. It follows that when a 1-in. ball breaks a 1-mm particle, a 2-in. ball will break a 4-mm particle, and a 3-in. ball a 9-mm particle. This is in accordance with practical experience, as well as being theoretically correct. Confirmation of this reasoning is supplied by the Third Theory of Comminution," which states that the work necessary to break a particle of diameter F varies as F. Since work equals force times distance, and the distance of deformation before breaka4e varies as F it follolvs that the breaking force should vary as F½ These relationships are expressed in Table I, with a 1-in. ball representing one unit of force and breaking a 1-mm particle. This establishes theoretically the general rule used in Eq. 1 that the ball size should vary as the square root of the particle size to be broken. Ball Size and Work Index: The work input W required per ton" varies as the work index Wi, and the

Jan 1, 1959

By Charles P. Berkey, Joseph H. Sinclair

This paper describes the results of a visit to the Santa Elena Peninsula, Ecuador, in January and February, 1921. On account of the complicated folding and faulting of the rocks and the peculiar occurrence of petroleum, field work was possible on only a few of the many geological problems and the results here published are only a contribution to the solution of the general problem. We know of nothing published on the geology of the district, except one account, in Spanish and German, by Theodor Wolf. The Santa Elena peninsula, the most important headland of the coast of Ecuador, lies two degrees south of the equator. The chief town, Santa Elena, is about 75 miles west of Guayaquil, the main seaport of Ecuador. The nearest developed oil field is the Zorritos of northern Peru, 120 miles south across the Gulf of Guayaquil. If newspaper reports are true, viz. that the Anglo-Ecuadorian Oilfields, Ltd., has recently brought in a 700-bbl. well on the coast of the peninsula near Ancon, the district may become important in the production of petroleum. At the time of our visit, the best way to reach the peninsula was by a steamer of the Pacific Steam Navigation Go., southward from Panama and northward from Guayaquil. About once a week, one of these steamers anchored 1/2 mile off shore near the village of Salinas, an important cable station. An alternative route, in the dry season, is by automobile from Guayaquil. In February, 1921, a railway bed had been graded part way between Guayaquil and Salinas, but this was detcriorating rapidly because of lack of funds to continue construction or to keep in repair that portion already completed.

Jan 1, 1923