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By Raymond E. Zimmerman

The Batac jig is a newly-designed Baum-type jig for coal washing. Multiple air changes directly under the bed permit uniform distribution of air pulsations, giving greater capacity and more accurate separation than standard Baum jigs. It is equipped with a new concept of air valves, automatically and electrically controlled, permitting infinite variation in jigging pulsations. It can be used for either coarse or fine coal cleaning. Sharp separations and high organic efficiencies are obtained at separation densities above 1.45 specific gravities.

Jan 1, 1976

By R. N. Kust, J. C. Agarwal, H. E. Barner, N. Beecher, D. S. Davies

Kennecott Copper Corp. has developed a new hydro-metallurgical process for extracting nickel, copper, cobalt and molybdenum from manganese nodules. This process (KCC Cuprion Process) utilizes a reduction leach operating at atmospheric pressure in the temperature range of 45 OC to 50°C and has been successfully demonstrated in a continuous pilot plant processing 350 kg/d of nodule ore. The chemistry, basic flowsheet and operating characteristics are presented and discussed.

Jan 1, 1980

Discussion of the paper of Ernest Stütz, presented at the New York Meeting, October, 1913, and printed in Bulletin No. 79, July, 1913, pp. 1257 to 1265. H. 0. HOFMAN, Boston, Mass.:-I have hunted in vain for a figure which gives a dehydration temperature of cement. I was wondering if Mr. Stütz had a figure of that kind. E. STÜTZ :-No, I am afraid I have not. H. 0. HOFMAN :-I believe some figures have been published that cement disintegrates at 400° or 500° C. It seems tome a very low temperature. Mr. Stütz just made the statement that the disintegration of this briquette begins at the temperature at which the briquette begins to sinter. E. STÜTZ :-Above the sintering point. H. 0. HOFMAN :-The sintering point would be about 1,200° C. E. STÜTZ :-Yes, roughly speaking. I am afraid I have no information about the exact degree of dehydration. H. O. HOFMAN :-I have looked a great deal for such a figure and I have not been able to obtain it. Professor Howe says I might have tried myself. But my interests were directed in another line; although-that is, perhaps, a good suggestion. CHAIRMAN RICHARDS:-Is not Portland cement sometimes used for the test in a cupellation furnace and a drying-out furnace ? H. O. HOFMAN :-Yes, it is used, but that is a higher temperature, of course. But I have been trying to get a figure as to where it loses all its water. I understand it begins to lose water at 500° C. But there are different silicates or hydrosilicates in the cement, so that one would have four or five or sit figures to show the different stages at which the cement loses its water. CHAIRMAN RICHARDS :-I would ask whether the use of this material in blast furnaces has shown increased capacity of furnace or decreased consumption of fuel.

Jan 12, 1913

By Joseph E. Pogue, F. G. Coqueron

Prior to 1934, published information covering expenditures by oil companies for properties, plant, and equipment and the residual investments in the various segments of the petroleum industry was not readily available. This paper presents an analysis of the funds invested by a selected group of oil companies during the years 1934 through 1942, based on an extensive study of the individual records of 30 oil companies. The sources, disposition, and characteristics of the capital employed by this group of companies are traced through a series of combined statements and appropriate charts, covering financial and operating data. The analysis of these statements not only constitutes a record useful for comparative purposes but also serves as a basis for a number of broad conclusions, as follows: 1. The petroleum industry generates the capital needed for its expansion almost entirely from its own operations and is therefore a singularly self-contained economic unit. 2. The long-term debt contained in the financial structure of the industry is small relative to the total capital employed, is less than the net working capital, and has increased but slightly during the period under review. 3. The capital expenditures of the group of 30 oil companies for 9 years amounted to 6.1 billion dollars, 58.3 per cent of which went into crude-oil producing facilities. The trend of this segment of expenditures has been downward since 1937. In 1942, expenditures for refining facilities increased sharply and for marketing facilities declined drastically. 4. The net working capital of the group increased 380 million dollars over the period, and the industry is in a favorable position of liquidity. The ratio of combined current assets to current liabilities was 3.21 at the close of the period. 5. The net income of the group displayed an upward trend for the period but with rather sharp cyclical fluctuations. 6. In 1942, Government funds entered the industry for the first time in substantial amounts and were utilized for the provision of war facilities needed for emergency use. 7. The group has paid out a relatively low percentage of net income in the form of dividends. For the period the average was 55 per cent of net income. 8. The rate of return on invested capital for the group averaged 6.4 per cent for the

Jan 1, 1944

By Joseph E. Pogue, F. G. Coqueron

Prior to 1934, published information covering expenditures by oil companies for properties, plant, and equipment and the residual investments in the various segments of the petroleum industry was not readily available. This paper presents an analysis of the funds invested by a selected group of oil companies during the years 1934 through 1942, based on an extensive study of the individual records of 30 oil companies. The sources, disposition, and characteristics of the capital employed by this group of companies are traced through a series of combined statements and appropriate charts, covering financial and operating data. The analysis of these statements not only constitutes a record useful for comparative purposes but also serves as a basis for a number of broad conclusions, as follows: 1. The petroleum industry generates the capital needed for its expansion almost entirely from its own operations and is therefore a singularly self-contained economic unit. 2. The long-term debt contained in the financial structure of the industry is small relative to the total capital employed, is less than the net working capital, and has increased but slightly during the period under review. 3. The capital expenditures of the group of 30 oil companies for 9 years amounted to 6.1 billion dollars, 58.3 per cent of which went into crude-oil producing facilities. The trend of this segment of expenditures has been downward since 1937. In 1942, expenditures for refining facilities increased sharply and for marketing facilities declined drastically. 4. The net working capital of the group increased 380 million dollars over the period, and the industry is in a favorable position of liquidity. The ratio of combined current assets to current liabilities was 3.21 at the close of the period. 5. The net income of the group displayed an upward trend for the period but with rather sharp cyclical fluctuations. 6. In 1942, Government funds entered the industry for the first time in substantial amounts and were utilized for the provision of war facilities needed for emergency use. 7. The group has paid out a relatively low percentage of net income in the form of dividends. For the period the average was 55 per cent of net income. 8. The rate of return on invested capital for the group averaged 6.4 per cent for the

Jan 1, 1944

By S. P. Brown, J. E. Worthington

Increasingly, exploration geology seems to be growing into a practice overly separated from production geology, yet the eventual purpose of exploration is a producing mine. To help in the search for new deposits, exploration "lessons" will be drawn from the papers given at the Fall AIME Applied Geology Symposium. In addition, examples will be re-emphasized to show how early observation of matters now routine to mine work (for example, hydrology, rock mechanics, oxidation state) can help increase the effectiveness of exploration.

Jan 1, 1984

By J. E. Campbell, G. D. Calkins, N. M. Ewbank, M. Pobereskin, A. Wesner

GRINDING for size reduction affects the economics of many processes and products. It is essential as the first step in many industrial processes and is also a finishing step for materials with properties depending on particle size, such as talc, cement, and silica sand. Intermediate and fine grinding are vital operations in the U. S. cement industry, which is producing more than 250 million bbl of cement per year.' Wear of the grinding media is a large part of the grinding operation cost. Problems encountered in grinding cement are so complex that evaluation of efficiency and economy of grinding media is difficult.2 It has been especially difficult to evaluate the relative effectiveness of different types of balls because there are no good testing techniques. Many other industrial operations can be evaluated on a laboratory scale with reasonable accuracy. This does not hold true for evaluation of grinding balls. The consistent results obtained in a laboratory test under a given set of conditions are not always borne out in field application. Rough evaluations of the effectiveness of various compositions and types of grinding balls have been made in the field by using a full charge of one type in a mill and comparing the production record with another run using another type of ball. This method is time-consuming and not very precise, as the second run may not have been carried out under identical conditions. Laboratory-scale tests, on the other hand, have yielded inconclusive results, and many investigators have turned their attention to the development of a field testing technique. Field testing small sample lots of grinding balls has been impractical because it is difficult to identify and recover the test specimens from the grinding mill, and individual groups of balls that have undergone different heat treatments can not be separated.".4 To overcome these difficulties, previous investigators have identified the balls by distinctive marks, notches, and drilled holes, but this procedure has three serious drawbacks: 1) Grinding characteristics and quality of the steel balls may be affected. 2) Physical markings may be worn away in the grinding process, especially during a prolonged run. 3) Recovery from the bulk of the charge will be extremely difficult because the markings are hard to see and may be masked by a coating of the product. To circumvent these difficulties, a radioactive-tracer technique was proposed for recovery and separation of steel grinding balls and subsequent evaluation of the various compositions of the balls. The proposed technique involved five basic operations: 1) Thermal-neutron irradiation activation5 of each group of test grinding balls to a different level of specific radioactivity. 2) Addition of groups of radioactive steel-ball specimens into a ball tube mill. 3) Recovery of radioactive steel-ball specimens from the bulk of the mill charge. 4) Separation of the various groups by their specific radioactivity. 5) Evaluation of actual grinding ball wear. Before any physical tests were performed, required neutron irradiation intensity and time were calculated. Probable composition of the steels to be used was ascertained. An examination was made of the radioactive nuclides8 to be formed which would contribute measurably to the radiation level immediately after irradiation and during the test operation. The radioisotopes formed, their types of radiation, and their half lives are listed in Table I. Of these radioisotopes only iron-59 and chromium-51 were significant for the actual wear test. The intensity of radiation that could be detected by a Geiger counter when the test was completed was the basis for the minimum activation level established. The intensity of radiaton that could be safely handled at the beginning of the test was the basis for the maximum activation level, although this was not considered a major problem. Ten groups of grinding balls of various composition and/or surface or heat treatment were to be tested. One group was designated for the minimum irradiation time. The remaining groups were designated for irradiation periods that increased by increments of 33 pct from that of each preceding group. This difference was considered enough for separation and identification of the groups by comparison of specific activity. Potential Hazards: Possible radiation hazards that might be encountered during this experiment were evaluated for the three important phases: 1) the radiation hazard of placing balls and removing them from the mill, 2) contamination of the product cement by radioactive material worn from the balls, and 3) contamination of the steel by the radioactive balls left in the mill. The radiation intensity expected from the whole group of radioactive balls was calculated to be 250 milliroentgen per hr at 1 ft. This meant the balls would require special shielded packaging and warning labels on the shipping containers. In a radiation field of 250 mr per hr a man can work for 1 hr without exceeding maximum permissible weekly exposure. Since the balls could be dumped into the mill in a matter of seconds, relatively little radiation exposure was anticipated at this stage of the operation. If the weight loss in the balls was 7.7 pct per month and the cement feed through the mill was

Jan 1, 1958

By C. A. Queener, W. L. Mitchell

In recent years bismuth telluride has received considerable attention, primarily due to its value as a thermoelectric material. Since there seems to be little dependence of thermoelectric effect upon crystal orientation1 many of those who have previously worked with this material have been unconcerned with knowing its crystallographic angles to any degree of accuracy. However, when one studies deformation modes, lattice stresses, or the aniso-tropy of mechanical properties of a material, knowledge of the interplanar angles and their relationships as depicted in a stereographic projection becomes essential. Lange2 originally published crystallographic data which described the structure of bismuth telluride as rhombohedral with the space group A more recent work by Francombe3 presented data which corroborated and refined the original findings of Lange. In addition, Francombe established that there is no coexistence of rhombohedral and hexagonal lattices in bismuth telluride as was suggested by Vasenin and Konovalov,4 and that the structure can be treated as a pseu do hexagonal structure with an axial ratio, c/a, of 6.95 for convenience of presentation and study. The interplanar angles and standard projections for several unalloyed hexagonal metals have been previously published,5-9 but the present work appears to be unique in that it presents data for a rhombohedral compound in pseudohexagonal form. As a pseudohexagonal structure, this material has an extremely large axial ratio as compared to the usual range of approximately 1.5 to 1.9 for the hexagonal metals. Since the interplanar angles are a function only of c/a, the angular relations between various planes for Bi,Te3 will be grossly different

Jan 1, 1965

By J. Walter Snavely

PRESENT-DAY processing in many industries, calcining, sintering, briquetting, beneficiation and nodulizing, increasingly calls for the handling of large volumes of hot bulk materials. Various types of conveyors have been employed. This discussion will cover the factors governing their selection. For temperature ranges up to 400°F, or approximately 200 °C, a wide range of conveyors is available. Special constructions of rubber conveyor belts, steel conveyor belts, vibrating and shaker conveyors, apron conveyors, and drag chain conveyors, all are used successfully. As temperatures go well above 400 2F, however, choice of conveyors is narrowly limited. This paper will consider the problem of handling bulk materials only where the temperatures exceed 400°F. The arbitrary selection of 400 °F as a dividing point undoubtedly can be challenged, as special conveyor belting constructions are available which are suitable for temperatures in excess of 400°F. However, when the relatively short life of such belts and the cost of their replacement, with the attendant down time, are balanced against the reliability and long service life of the properly designed steel constructed units to be discussed, there is little question in any operator's mind that the special belts are more expensive to use. Because the conveyors under study are for the handling of bulk materials, inevitably including a high proportion of fines, obviously wire mesh belts cannot be included for consideration. Even though this type of conveyor is widely used at high temperatures, i.e., for carrying glassware through a lehr, it is unsuited for the conveying of bulk materials, and therefore will be excluded from further discussion in this paper. Preliminary to the study of the conveyor itself is the determination as to whether the material is to be cooled while it is being handled, or whether the processing requires retention of all heat and the maintenance of a given temperature range. In the majority of cases cooling is incidental to or part of the handling process, when the handling, for example, follows completion of sintering, roasting, calcining, refining, or some other process. To meet such operating conditions successfully, the conveying medium used must have: 1—a construction capable of withstanding maximum initial temperatures of the material being handled. 2—a construction providing efficient heat transfer for cooling. 3—a construction providing dependable operation and long life with minimum service requirements, and 4—a construction providing controlled and efficient conveying. Under the usual conditions of cooling during the handling, the construction selected to withstand the initial maximum temperatures does not necessarily involve using alloys, as excellent results can be achieved with normal carbon steels and cast irons, when they are properly applied and proportioned. The earliest and simplest type of conveyor for handling very hot materials is the cast steel drag chain conveyor, still widely used for handling hot cement clinker, as illustrated by Figs. I and 2. Because of the rugged and generous proportions of the chain link design, low carbon steels are entirely suitable for the links. The pins, however, must be alloy steel. The simple, rugged construction of this type of conveyor makes it readily capable of withstanding high initial temperatures, even though the chain is operating buried in the material. The drag-chain type of conveyor has advantages and limitations. Although the efficiency of the heat transfer is relatively poor, the life of the conveyor is reasonably long, and because of its crude simplicity it does not require much servicing. However, as a conveyor, it is limited in capacity, and largely limited to horizontal runs. Furthermore, because of the crude design, heavy weight, and the chain operating at the temperature of the material, greatly reducing permissible operating chain pulls, this type of conveyor is limited to relatively short centers. Another type of conveyor that has been used for very hot materials is the cast pan conveyor. Because of its very generous proportions the cast pan, which is made of either cast iron or malleable iron, can withstand initial maximum temperatures. It also provides efficient heat transfer for cooling. Further, it is on efficient conveyor construction, which can be used for inclines. Because the chain employs rolling friction instead of sliding friction, and is not in the maximum temperature zone, much longer centers are possible. It is this type of conveyor that is frequently used in the casting of various metal pigs, pig iron, and aluminum; it is obvious, therefore, that very high initial temperatures are being handled. With this kind of conveyor the return run is frequently sprayed with water to accelerate heat transfer. The build-up of residual heat in the very heavy cast pans is thus overcome. The outboard roller steel pan conveyor is an improved pan conveyor' which provides high rates of heat transfer and substitutes formed steel pans for the heavy cast pans. It is a very efficient conveying medium. The details of this particular construction are clearly shown in Fig. 3. An early application of this type of conveyor is shown in Fig. 4. In this case the conveyor units are handling roasted phosphate rock at average temperatures of 1000" to 1500°F, and frequent maximum temperatures as high as 1900°F. Several widths are used. The capacity of the unit at a speed of 50 fpm is approximately 30 tph per inch of width at peak loadings, average capacity being about 1/3 of peak loading. The assembled conveyor is shown in Fig. 5, with views of both the top and the underside to show all the construction details. In particular, the following general design principles were carried out in this construction:

Jan 1, 1954

By R. S. Handy

Does hand-sorting of mill feed pay? The fact that the practice is so general would seem to indicate that there must be good reasons for following it; yet, to my mind, the advantage in many cases is doubtful enough to invite a thorough discussion of the subject, in an effort to determine under what conditions hand-sorting does pay. A typical sorting plant in the Cœur d'Alene district handles about 800 tons of mine-run ore per day. The material is run over a grizzly with about 4-in. opening; the oversize is crushed to about 4-in. size and joins the undersize, and the whole product is washed and screened on trommels with about 11/2-in. openings. The oversize from these trommels is spread on a belt and the waste and first-class ore are sorted out by hand, the residue being further crushed, and passes, together with the trommel undersize, into the mill feed bins. About 50 tons of shipping ore and 150 tons of waste are sorted out per day, leaving about 600 tons of mill feed. To sort out this material requires 20 ore sorters and five bosses and repairmen. The normal cost of hand-sorting is about 16 c. per ton of run-of-mine ore, or 65 c. per ton of material sorted out. I believe that straight crushing and milling of the run-of-mine ore, in this case, is not only cheaper, but is metallurgi-cally more economical and efficient than hand-sorting followed by milling of the residue. At one of the Bunker Hill rock-houses, 1250 tons of run-of-mine ore are reduced to pass 30-mm., or about 11/4-in., opening in six hours, requiring five men. The equipment and the power consumed are not greatly different from that required in a sorting plant of the same capacity, and the normal cost of thus preparing the feed for milling is under 3 c. per ton of mine product. The relative cheapness of this method as against hand-sorting is thus clearly established, the saving being about 13 c. per ton. If there is any virtue in, hand-sorting, then, it must lie in the metallurgical results achieved. Possibly the favor that hand-sorting has found is due to a misconception as to what happens to the clean shipping ore if it is sent to the mill. If a mill makes, say, an 80 per cent. recovery on its feed, the hand-sorting advocate may assume that if the hand-sorted product be sent to the mill, 20 per cent. of the metals contained in it will be lost. This, I believe,

Jan 1, 1920

By C. E. Berry

NATURAL stone or manufactured porcelain pebbles are used as the grinding elements in pebble mills and the mills are lined with stone or porcelain blocks. Steel balls usually form the grinding medium in ball mills, and the mills may he lined with steel. Pebble mills are used to avoid contamination by metal of the material being ground, to avoid rapid wear due to corrosion, and for other specialized reasons. Belgian silex (silica) blocks formerly were used widely for lining pebble mills, because of their good resistance to wear, light color, and low cost, but the importation of these blocks was suspended early in the war. The stocks of Belgian silex diminished so rapidly that it was found desirable to compare the properties of a number of domestic substitutes for use in critical operations. Accordingly, a study was made to determine the best substitutes for fine grinding of white pigment and for certain heavy-chemical process work such as the disintegration of phosphate rock in weak phosphoric acid solution. A comprehensive survey of substitute lining materials was not contemplated. It was stipulated at the beginning of the investigation that only materials on the market, available in large quantities, would be tested and that these should be reasonable substitutes from a wear-resistance viewpoint as determined from known physical characteristics. Marble and limestone, for example, could be eliminated immediately. Quartzite, granite and porcelain were selected as likely substitutes for Belgian silex and were subjected to quality tests. MATERIALS TESTED I. Belgian Silex.-Sample blocks of a standard grade were taken from crates in which they had been imported. The stone was light gray in color and contained inclusions of small clusters of quartz crystals as well as light colored portions that might be calcite. This was a siliceous rock which appeared to be of igneous origin. 2. Quartzite.-Samples of quartzite were obtained from a quarry in Minnesota. The blocks exhibited a crystalline fracture. The stone was pink in color and composed of fine quartz crystals bound together with a siliceous cement. 3. Granite A was a fine-grained granite quarried in South Carolina. 4. Granite B was coarse-grained granite quarried in South Carolina. 5. Granite C was a coarse-grained granite quarried in North Carolina. 6. Porcelain A.-Sample blocks of the pebble-mill lining material were obtained direct from the manufacturer. It was white in color and of lower specific gravity than the natural materials. It did not have a crystalline structure. 7. Porcelain B.-Sample pebble-mill lining blocks of the porcelain were received direct from the manufacturer. It was white

Jan 1, 1946

By R. W. Jenkins, O. J. Neslage

The need for mechanized drilling to decrease mining costs has resulted in the development of the jumbo from column-and-bar drill carriages to hydraulically controlled jib jumbos. Resultant savings from use of improved jumbos have been high enough usually to retire the machinery investment within a year. The jib jumbo unit has been applied successfully to development and preparation programs and flat-bedded stopes, and an even wider field is forecast for its use.

Jan 1, 1950

By A. L. Bailey, B. A. Landry

W. W. ANDERSON and G. E. KELLER*—We want to compliment the authors on this very thorough paper. It gives information which the coal industry has needed for some time. We hope that the additional information which the authors are collecting will he available shortly. The mixing and riffling procedure that was followed for experimental purposes is obviously not practical in routine float-and-sink testing because of the particle size degradation which would result in handling the sample so many times. It is important to obtain our tloat-and-sink fractions with a minimum amount of handling of material. A statement is made in the paper (p. 80) that "the variable most likely to affect the size of sample required to meet a given preassigned accuracy would be the state or degree of mixing of the coal." We agree that this is a large factor, but do not believe it is the most important factor. Our own opinion is that the most important single factor governing the total gross weight of sample that must be collected is the percentage of the weight of material in the smallest fraction that results from the screening and float-and-\ink operations. In other words, size of sample is governed by the total number of fractionations that must he made, and the distribution of material within the fractions. We can imagine a coal with perfect mixing, but with such a small amount of material in some float-and-sink fraction in one of the coarse sizes that a much larger sample would have to be taken than would be the case with very poorly mixed material, but with a large percentage of coarse material more evenly distributed in all float-and-sink fractions. Our own observation of many float-and-sink tests that we have run in our own organization on many types of coal is that the size of sample that must be used on fine size float and sink is governed more by the requirements for weight of material to be used for analysis in the laboratory than by weight of material necessary to obtain accurate float and sink percentage of weight values. In other words, it is our opinion that very small samples can be used for float-and-sink fractionation in the fine sizes, but that accurate analysis of the fractions will depend on a larger weight of sample being pulverized for the laboratory than is necessary to establish the float-and-sink distribution with respect to weight. A. L. BAILEY and B. A. LANDRY (authors' reply)—The authors thank Messrs. Anderson and Keller for their comments based on long experience. It is agreed that the involved mixing and riming technique used may be disadvantageous from the standpoint of degradation. Fortunately, the paper does point out that the extended riming was unrewarding in causing further mixing. Two large unknowns remain, however: (1) how much of the mixing from the presumed highly unmixed state in the bed was achieved toward the random state during blasting, loading, transportation, screening, and further transportation to the point where the gross sample was taken, and (2) how much of the mixing took place during the preparation described preceding riming. As has been pointed out by one of the authors.6 the degree of mixing has a very large effect on the size of sample required and there are still too few experimental data to show at what stage of coal handling most of the mixing occurs. The discussion states that the weight of material in a screened fraction, or in a float-and-sink fraction, is more important than the mixing factor. We do not believe that these factors are comparable in this instance inasmuch as our purpose was to give minimum sampling requirements to achieve a preassigned accuracy in the percentages of float, middlings, or sink, and nothing more. The gross sample had already been screened and no further division by screening was made or contemplated; also, it was not intended that the middlings and sink fractions would necessarily be adequate for percentage ash or other determination. In other words, the sample obtained by the method outlined is not intended for washability studies but only for preparation plant control. Further experimental work has been done, since the paper was prepared, to investigate the effect of increasingly larger top and bottom sizes on the variability of float, etc., of a double-screened coal from Western Pennsylvania. Results will be published and eventually attention is to be given to the preparation of sampling specifications. E. H. M. BADGER*—I should like the authors to explain more fully the fundamental assumptions on which their Eq 4 is based. The equation is of the form s2 = p(l - p) which is the usual expression for the (standard deviation)2 when the chance of finding a particular kind of particle in the sample is proportional to the number fraetion, p. But instead of the number fraction, the authors have used the weight fraction, WF/W. The chance of finding a particular kind of particle in the sample can only be proportional to the weight fraction, if the average ?eig?ts of all kinds of particles, that is, float, midlings, or sink, are the same. Surely a much more justifiable assumption would be that the average volumes of the particles are the same, and, if this is so, Eq 4 would not be true. This may be demonstrated as follows: Let be the weight fraction of float, middlings, or sink, dl the density of this fraction, and d2 the density of the rest of the coal. Then assuming that the average volumes of the pieces in the three classes are the same, the number fraction, p, is given by ? P = d1/l-?/d2 + ?/d1 = ?d2/d1 + ?(d2-d1) The weight fraction, w, in terms of p is given by ? = pd1/(l-p)d2 + pd1 = pd1/d2 + p(d1-d2) _____ [61

Jan 1, 1950

Plans for a multi-million dollar uranium mill in southeastern Utah were confirmed by Vanadium Corp. of America officials. Atomic Energy Commission has been dickering with two private firma for mill construction at Hite, Utah and Shiprock, N. Mex. Colorado Fuel & Iron has contracted to purchase all manufacturing business, plants and inventories of John A. Roebling's Sons Co., through a wholly owned subsidiary. CF&I currently the nation's ninth largest steel producer, will operate the new subsidiary under the Roebling name.

Jan 1, 1953

One of the best known of the early mining districts of Nevada was that at Eureka, where gold was first produced in 1864. While the most productive period for the district was during the early 1900s, production was maintained until World War II. During this time it is said that the Eureka district produced as much as $200 million from over 3,200 km (2,000 miles) of tunnels and workings. The Windfall Mine was discovered in 1902; and a 100 ton cyanide plant was operated there during the periods 1907 to 1912 and 1932 to 1940. Since 1974, however, Windfall Venture has been operating a unique heap leaching facility on the site of the old Windfall Mine. Gold at the Windfall occurs primarily in the Cambrian Hamburg dolomite which has been "sanded" and in which the original cementing material has been destroyed. The gold is micron sized and is accompanied by anomalous amounts of arsenic and mercury; however, the ore is low in base metals which distinguishes it from the typical Eureka district ores. Mining the ore causes it to break down into very fine sizes, approximately 50% minus 150 microns (100 mesh). Conventional heap leaching by spraying proved to be impossible; and extensive test work on alternate methods, aided by the United States Bureau of Mines, was conducted using 9 m (30 ft) high columns and then large tonnage heaps. The leaching procedure adopted was to build 10.6 m (35 ft) high heaps on a 50 mm (2 in.) asphalt pad covered with a layer of gravel for drainage assisted by perforated pipe drains and gravel-packed sumps. A berm is constructed around the top of each pad and cyanide solution is allowed to “pond” on top of the heap. The resulting percolation is at the rate of 24 liters per sq m per hr (0.6 gal per sq ft per hr), a figure about 2.4 times the average used in other heap leaching operations. The dolomite furnishes its own protective alkalinity and no lime or caustic is used in the solution which contains 0.4 kg per mt (0.8 lb per st) of NaCN. Pregnant solution from the heaps averages about 0.7 g Au per mt (0.02 oz per st) with a cyanide content of about 0. 2 kg per mt (0.4 lb per st) and a pH of plus 9. Addition of NaCN brings the pH of the leach solution up to 11 to 12. Carbon adsorption of the gold was originally done in six enclosed columns mounted in a trailer but five conventional columns have been substituted as the size of the operations has increased. Carbon stripping is with hot caustic, cyanide and 10% ethanol solution. The

Jan 1, 1981

By C. F. Knutson, Martin Felsenthal, J. D. Lung, J. D. Sudbury

This paper has been written to summarize the laboratory and field studies leading to injection of Pacific Ocean water into the Third Grubb formation. Laboratory tests for studying the permeabilities of cores to Pacific Ocean water are described. Resu1ts of these tests showed that the water could be successfully injected if it is chertzically stable, sterile, and properly filtered. Clay swelling was shown to be a comparatively minor factor in permeability reduction. Studies leading to development of an economical system for properly conditioning the ocean water are discussed. It was necessary to control pH, calcium carbonate deposition, bacterial growth, iron solubility. oxygen, and suspended .solids. A relatively simple system combining the use of sulifur dioxide, a quaternary rlmmonium chloride, and filtration was shown to adeqrrately control all these factors. This combination has not previozrsly been used as a water treating technique. Evaluation tests of the water conditioning system prior to injection are described, and results of such tests are presented. Control measures needed to insure a continusously adequate water treatment are reviewed. INTRODUCTION The utilization of sea water for waterflooding has only recently been attempted. Knowledge of treating requirements has been limited, and in most cases no water treating has been done. Prior to the initiation of the Third Grubb zone pilot water flood at Ventura, Calif.,1,2 extensive laboratory research was done in an effort to define the treatment necessary to provide a suitable injection fluid. Close coordination and cooperation between field and research personnel were found to be the foremost requisites in developing the most economical and practical method for treating ocean water. LABORATORY STUDIES Plugging Tendencies of Untreated Ocean Water A program of laboratory studies was initiated for the purpose of evaluating whether ocean water could be injected into the formation for prolonged periods of time without causing plugging of the pore spaces of the formation. Because of the fairly low permeability of the formation (averaging about 40 md), it was realized from the outset that a water of exceptionally high quality was needed. Ocean water used in the laboratory studies was carefully filtered through ultrafine fritted glass filters, having nominal maximum pore sizes of 1.4 microns. Permeability tests were then performed on representative core material from the formation. It soon became apparent in these tests that filtration alone was not a sufficient treatment for the ocean water. Prolonged injection of the filtered water caused rapid plugging of the core samples. The permeability decline of a typical sample is illustrated by Curve A in Fig. 1. The several factors which can contribute to this plugging tendency are briefly reviewed below: Interaction with Formation Clays An evaluation of clay swelling effects on liquid permeability was made using core samples from the Grubb formation. An oil was chosen for determining the reference liquid permeability because it would not tend to react with any clays that might be present in the formation. The samples were then extracted, dried, and resat-urated with either treated ocean water, formation water, or fresh water. Permeabilities to the various waters were then determined. Results of this test series are summarized in Fig. 7. The difference between the oil permeability and water

Jan 1, 1957

By Margaret Dienes, Karl L. Fetters

IT is recognized that ingot structure is important in determining the quality of finished steel. Such elements of ingot structure as the size, shape and distribution of primary crystals; the size, shape and distribution of shrinkage cavities; the degree and kind of segregation (both macro and micro); the distribution of nonmetallic impurities and gas cavities are each the result of solidification phenomena. These structural features vary with variations in ingot practice: Their control requires a knowledge of their correlation with the variables of ingot practice. The structures of full-sized commercial steel ingots have been studied extensively by the Steel Ingots Committee of the Iron and Steel Institute.1 This work has contributed much to our knowledge of ingot structure and the factors that control it. These investigations of necessity have been limited in scope because of their costliness and the great amount of time required. Studies of the structure of small steel ingots have not been very satisfactory because of the rapid rate of freezing of the ingots. Their structures are thus difficult to resolve and to interpret. Medium sizes of steel ingots in the ranges of 150 to 500 lb. do offer opportunity for study, but their cost also becomes greater and their use less expeditious. Experimental models for the study of steel-ingot structure have been made of nonferrous metals. Gathmann2 used lead-tin alloys and reported that they were satisfactory for determination of the location of the major shrinkage cavities and the "planes of weakness" resulting from the pattern of crystallization. Certain features of mold design have been illustrated by these results. Bezdenezhnihk3 has used copper and aluminum ingots similarly and has shown the effect of mold-wall thickness on the depth of columnar grains. These studies of nonferrous models have been useful, but their results have been limited because the analogy with steel ingots obviously is not a close one. Model wax ingots have also been useful in pointing out important features of mold design. Gathmann4 watched the solidification of wax ingots, through a glass-walled mold, and has reported the effects of mold type on the location of the shrinkage cavities. Segregation was simulated by using a mixture of paraffin and stearine and adding a red dye that on freezing segregated preferentially to the stearine. Brearley5 also used wax models to determine .the effects of some of the principal features of mold design. Carlsson and Hultgren6 used stearic acid to show that the columnar crystals of ingots tend to slope upward progressively from the mold wall toward the center of the ingot. As with the nonferrous ingots, the wax ingots have

Jan 1, 1943

By Thomas M. Barczak

This paper describes a computerized simulation system developed by KETRON, Inc., in cooperation with the US Bureau of Mines. The system is designed for use by mine managers and engineers rather than computer specialists and allows the novice user to arbitrarily create room-and-pillar mine plans, with the computer illustrating the simulation of face activities on a video screen. Extensive visibility of the simulation is provided by visual display of the machine elements and their movement in the section to accommodate the extraction, roof control, place changing, and haulage operations. Performance statistics are also generated to analyze the operations. Both conventional and continuous mining methods can be simulated. In addition to simulating arbitrary scenarios, the system can also be used to display actual time-studied operations and maintain a file of these activities. Practical applications and case studies are provided in the paper to demonstrate the usefulness of the system. The system is designed to operate on a desk-top mini computer and is currently installed on a Hewlett-Packard 9854B system with 450 K bytes of memory.

Jan 1, 1983

By H. Brandhorst, C. G. Dunn

THE main purpose of the present note is to provide further information on the effect of orientation on strain-induced grain boundary migration in sheet material. A secondary purpose is to draw attention to the presence of a possible (hkl) dependent surface energy driving force, when the two grains of a bicrystal specimen have different (hk2) planes at the gas-metal interface, and to suggest means of eliminating this effect from the strain effect. If two grains A and B of a bicrystal have specific surface energies ? A and ? B respectively, there exists a driving force for grain boundary migration equal to 2(? B - ? A)/t, where t is the thickness of the specimen.' When t is quite small and the residual strain energy densities €A and €B of grains A and B, respectively, adjacent to the boundary are nearly equal, it is possible that the surface effect may be larger than the strain effect. he difference € B - € A gives a driving force per unit area of grain boundary due to an applied strain.)2 However, with increasing t the surface effect decreases while the strain effect presumably remains constant. Therefore, at some sufficiently large value of t, it is possible to ignore surface energy driving forces.* On the other hand, the surface effect is eliminated for all value of t when both grains have the same (hkl) orientation or have identical values of surface energy. Although the previous study provided information on the effect of orientation on strain-induced erain boundary migration,' no orientations with the (100) plane in the plane of the sheet were included. The present paper provides information of this kind. A seed crystal reorientation technique4 was used to prepare a large bicrystal of silicon iron (3 1/4 pct Si) consisting of two (100) grains in the (100) [001] and (100) [Oil] orientations. After a cold rolling strain (7 pct reduction to 3/4 mm final thickness), small specimens were prepared as before2 and then annealed in hydrogen. The boundary migration was such that the grain in the (100) [011] orientation always grew. After overnight anneals at 1100° or 1000°C, none of the initial (100) [001] grain remained. An anneal of 30 min at 950°c, however, provided information on the start of boundary migration. Fig. 1 shows both the initial position (lower line) and the final position (upper curved line) of part of the boundary, as revealed by thermal grooves. In a different area, part of the final boundary was not revealed by thermal grooves; the rate of boundary migration presumably was too high for the growth front to be defined by the time dependent thermal grooving process. In another area there was a vacated intermediate position connected to the initial position of the grain boundary. This indicates that only part of the boundary

Jan 1, 1962

By Ralph D. Parker

The Frood mine (Fig. 1) is 2½ miles north of Sudbury, at a general elevation of 1000 ft. above sea level. It includes the original Frood location, which was the No. 3 mine of the Canadian Copper Co., and the adjoining Frood Extension, which until 1929 belonged to the Mond Nickel Co. Mine Openings The mine is operated from two vertical shafts and one underground incline shaft. No. 3, for men, supplies and ore, is a vertical six-compartment shaft in the footwall of the south mining area and extends froin surface to a depth of 3045 ft. No. 4 is a vertical three-compartment shaft, 2500 ft. north of No. 3 and adjacent to the north mining area. It extends from surface to a depth of 3345 ft., and is used exclusively for hoisting ore. No. 5 shaft, inclined at 65' to the horizontal and of three compartments, is near No. 4 and extends from the 2600-ft. level to 77 ft. below the 3100 level. It is used for development below the 2800 level. The ventilation shaft, No. 1, with which this paper is directly concerned, will be described fully in the following pages. Mining Methods The orcbody is laid out in 80-ft. sections at right angles to the gcneral strike and numerically indicated from the southern extremity toward the north. Each section consists of a 45-ft. stope and a 35-ft. pillar, extending the full width of the orebody. Pillars and stopes are superimposed from level to level. The present mining territory is divided into two vertical areas, a north and a south area. Both areas have the same number of stopes on each level, the two areas being separated by 10 sections to be mined later. In a few instances, at the ends of the south area, where the orebody has narrowed, longitudinal stopes are used, of more than the usual 45-ft. width. It will thus be seen that mining operations hare been concentrated in two relatively small areas, offering

Jan 1, 1934