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By N. J. Grant, K. M. Zwilsky

A series of copper-alumina dispersion strengthened alloys were prepared using three different copper and two different alumina powder sizes. Improvements in strength of up to ten times that of pure copper were found in room-temperature as well as elevated-temperature stress-rupture tests. Conductivity values remained from 70 to 90 pct that of pure copper. The best alloys were based on the finest metal powder (1 p) and contained from 7.5 to 10-0 vol pet of 0.018 or 0,033 p alumina. DISPERSION strengthening of metals consists of strengthening a metal matrix by the addition of oxides or other refractory constituents, followed by strain hardening, in order to reinforce the base material and extend its temperature range of application. Since the discovery of SAP by lrmannl and Von zeerleder2 in 1949, many investigators have reported data obtained on a variety of metal-inert oxide alloys.3-7 The present work deals with the mechanical mixing of copper and alumina powders. While several techniques of introducing the second phase have been investigated to date, mechanical mixing so far has been recognized as the simplest, least expensive, and most universally applicable technique that can be used to obtain an ultrafine dispersion of inert particles. In the present case, the effects of changing the metal particle size, the oxide particle size and the volume percent of oxide were investigated. This work is an extension of a previous study.6 EXPERIMENTAL PROCEDURE The starting materials were three different grades of copper powder and two different alumina powders. The three metallic powders were -74 µ(- 200 mesh) electrolytic copper powder consisting of irregular equiaxed particles having a normal screen distribution and analyzing 99.5 pct Cu, and two powders having a 5-µ and 1-µ particle size, average, respectively. The latter powders were produced by the hydrogen reduction of purified sulfate solutions at elevated temperature and pressure and were furnished by the Sherritt Gordon Mines, Ltd. The alumina used was obtained from Godfrey L. Cabot Co., under the trade name of Alon C, one batch of powder having a particle size of 0.018 µ, the other a particle size of 0.033 µ. The crystal structure of these alumina powders was found to be cubic gamma. Each batch of powder was mixed in a Waring Blendor for a total of 15 min, followed by a low-temperature reduction in hydrogen and vacuum compacting in cylinders which were hydrostatically pressed at 35,000 psi to give a compact that could be handled conveniently. Six of the compacts were hydrogen sintered at 900°C while the remaining four were sintered at 1000°C. Subsequently, all compacts were hot extruded at 760°C with a reduction of area ratio of approximately 20 to 1. Evaluation of alloys included density and resistivity measurements, room-temperature tensile tests, stress-rupture tests at 450°C plus one other temperature per alloy, hardness tests, microstruc-tural examination, X-ray analysis of the dispersed phase after dissolving the metal phase, and oxidation studies on two of the alloys. A vacuum-melted, wrought copper rod was tested together with these alloys for comparison purposes. RESULTS Table I is a summary of the compositions, sintering temperatures, densities and conductivity values of the ten alloys prepared. All alumina additions are given in volume percent, and will be referred to as

Jan 1, 1962

There were two official representatives of Foreign Governments as guests at the Chicago Meeting. One was Frederick Goransson, managing director of the Saudvikens Steel Works, who represented the Jernkonteret, the Iron and Steel In-stitute of Sweden, and who came to this country especially for our meeting. The other distinguished guest was Commander Toublet, who repre-sented the French Embassy. Commander Toublet delivered a most polished address at the Banquet, dealing with the felicitous relations existing between this country and France for the last 150 years. Com-mander Toublet was accompanied by Mrs. Toublet, who won the hearts of all the ladies present, whose guest she was.

Jan 11, 1919

By Guy Borden, Michael J. Rzasa

Laboratory data on bubble point pressures and reservoir volume factors have been correlated as functions of solution gas-oil ratio, calculated gas gravity of the pentanes-and-lighter fraction of the entire fluid, differential residual oil gravity, and reservoir temperatures. INTRODUCTION Several correlations of crude oil properties have appeared in the literature. D. L. Katz4 in 1942 presented five methods of predicting oil shrinkage. these being of decreasing accuracy for decreasing amounts of information available. M. B. Standing- in 1947 published three correlations of laboratory flash vaporization data of California crudes. From values of GOR (gas-oil ratio), gas gravity, liquid gravity, and temperature, his correlations will predict bubble point pressure, formation volumes of bubble point liquids, and two-phase formation volumes. Curtis and Brinkley2 in 1949 presented several correlations. From the gas-oil ratio, an approximation of reservoir volume factor and barrels of condensate recoverable per barrel of reservoir space may be obtained; along with liquid gravity and reservoir temperature, the GOR will allow prediction of bubble point pressure. These last cor- relations seem to be more qualitative than quantitative. Generally, laboratory bottom hole sample tests furnish information on solution gas-oil ratios, residual oil gravities. bubble point pressures, viscosities of oils, liquid shrinkages, and occasionally gas gravities. Each of these data has its own applications and use in reservoir engineering calculations. The particular uses of correlated bottom hole sample data are found in (1 Providing a basis for obtaining estimates of formation crude properties in fields where bottom hole sampling is impractical or impossible. (2) Greatly reducing the time in obtaining the desired information. (3) Determining the applicability of the results from various bottom hole samples to particular field problems. (4) Avoiding, in many cases, the uncertainties of sampling by replacing it with an element over which greater control can be exercised. (5) Permitting use of preliminary field data in application of production procedures before a bottom hole sample can be obtained and analyzed in the laboratory. (6) Serving as a check on data which may appear out of line. (7) Estimating for a particular type crude the appropriate equilibrium constants by working backward from the bubble point pressure. (8) Estimating original or other past history properties of reservoirs that were not sampled in the past. PROCEDURE Application of the published correlations4,6 to Stanolind laboratory data indicated that the general scheme presented by Standing" could give dedrable results if changes were made in parameter positions and scales. The correlation curves were drawn with all the variables having consistent gradations except the temperature increments which were drawn in to best fit the data. The variables from available Stanolind laboratory data are defined below: (1) Gas-oil Ratio: Gas is liberated at reservoir temperature by differential vaporization (or rather by a series of flashes, approaching differential vaporization) and measured at atmospheric pressure and temperature, at which the compressibility factor is assumed to be unity. The oil is the residual liquid remaining after the pressure has been reduced to atmospheric. For the gas-oil ratio both volumes are corrected to standard conditions of 14.7 psia and 60°F. (2) Gas Gravity: It was decided to arbitrarily divide the hydrocarbons of the entire bottom hole sample into pentanes-and-lighter and hex-anes-and-heavier, and use a calculated gas gravity of the pentanes-and-lighter for a correlating variable. (Sample calculation is shown in Table 111.) (3) Liquid Gravity: This is the API gravity of the residual liquid from the differential vaporization. The gravity is measured at room temperature and corrected to 60°F.

Jan 1, 1950

By Michael A. Gustafson, Fernando B. Sotelino

INTRODUCTION The degree of success of any new project will ultimately depend on two factors: (i) the underlying economic strength of the project; and (ii) how successfully the parties involved can coordinate their interests to establish a financing structure which indeed capitalizes on the project's inherent economic attractiveness. A project's underlying economic attractiveness is determined by its inherent comparative advantage in accessing the factors of production, combining these factors of production to create a product and marketing this product. Initially, the decision regarding whether to go ahead with the project - the go/no go decision - has to be based on its total ability to generate cash flow returns to project participants to compensate them for resources committed and risks undertaken. The extent to which project sponsors will actually benefit from the project's inherent economic attractiveness is fundamentally dependent on their ability to establish the long-term financing arrangement which best meets the objectives and constraints of each project participant and sources the most advantageous funding. The decision regarding the financing of the project, the "how-to-go" decision, is then an attempt to find the capital structure, consistent with market terms and conditions and sound financial management policies which will best: (i) match planned outflows with cash inflows; and (ii) maximize the return for project sponsors. This article makes use of a case study to illustrate the analytical process involved in dealing with the go/no go and the how-to-go decisions on a project. Though most commonly applied to project finance situations, the guideline described herein is equally applicable to other financing situations, such as leveraged buyouts and debt restructurings, where timely repayment of debt depends largely on the future cash flows to be generated by the entity being financed. Our case study is of a hypothetical expansion of an existing copper mining operation in Latin America. The company, Latin American Copper (LAC) currently has 147 million metric tons of proven sulphide ore reserves. The existing mine complex consists of an open pit mining operation and a conventional flotation plant with a processing capacity of 20,000 metric tons of ore per day. LAC sells copper concentrates primarily to Japanese custom smelters. The mine has been in operation for 5 years, fine copper production last year was 108,240 metric tons, and LAC currently has assets of $175 million and net worth of $135 million. Due to a recent increase in proven reserves, LAC is considering doubling the treatment capacity of its flotation plant to 40,000 tons per day, and expanding its mining capacity to meet the additional concentrator capacity (the "Project"). Project capital expenditures and incremental working capital have been estimated at approximately $210 million. The construction period is expected to cover one year, including a three-month start-up phase. The Project is to be funded by a combination of equity, senior debt, including both supplier credits and/or commercial bank term loans, and subordinated debt. LAC's shareholders are willing to commit a maximum of $50 million to the Project. A government development agency, which is also a minority shareholder, might be willing to lend approximately $40 million to LAC in the form of subordinated notes, depending on the ability of the Project to generate foreign exchange reserves. Commercial banks cannot be expected to offer loans with maturities longer than ten years. The Project is expected to increase the value of the firm because: (i) It will solidify the Company's position as a low-cost copper producer by taking advantage of economies of scale. As a result of the expansion, cash production costs are expected to decrease 7.0[c]/lb. and all-in costs 11.0[c]/lb; (ii) The potential earnings stream from the sulphide ore reserves will be realized

Jan 1, 1985

ALEXANDER, JOHN S, M,'74-6,'81-3 ASHBURNER, C A, M, '85-7 ASMUS, GEORGE, Dl, '78-80 BACON, D H , M, '00-2 BAYLES, JAMES C , M,'80-2,'91-3 P, '84-5, V,'88-9 BIRKINBINE, JOHN, M, '83-5, V, 'S7-J, P, '91-2 BLAIR, THOMAS S , M, '74-43 BLAKE, WILLIAM P , V,'71-4,'76-7,'05-6 BLANDY, JOHN F, V,'71-3,'78-9 BLOW, A A, C, '05-7 BOWIE, A J , JR, V,'93-4 BRAMWELL, J H , A1, '89-91 BROOKS, THOS B , M, '72 BRUNTON, D W , V, 97-8 BUCK, STUART M , M, '83-5 BURDEN, JAMES A, V,'80-1 BURNHAM, WILLIAM, M, '82-4. CAMPBELL, H H, M, '93-5 CHAMBERLAIN, H S , V, '96-7 CHANNING, J PARKE, V,'08 CHRISTY, S B , V. '91-2, '07-8 CHURCH, JOHN A, M, '79-8l,'94-6, V, '07-8 COGSWELL, W B, V, '74-5,'86-7

Jan 1, 1910

NEW MEMBERS The following list comprises the names of those persons who became members during the period Jan. 10 to Feb. 10, 1915: Members AGASSIZ, RUDOLPH Louis, Vice-Pres.,Calumet & Hecla Mining Co., Boston, Mass. AMIDON, RICHARD G., Min. Engr 3659 Girard Ave. N., Minneapolis, Minn. ANDERSON, HECTOR G. S., Min. Engr 323 N. 13th St., Muskogee, Okla. ANDERSON, JOHN CARTER, Petroleum Engr 2307 S. 33d St., Omaha, Neb. CLARKE, ALEXANDER FIELDER, Works Mgr., Firth-Sterling Steel Co., McKeesport, Pa. CROSS, CARLOS GARCIA, Min. Engr Calle Valparaiso 194, Vina del Mar, Chile. DAWSON, THOMAS WILLIAM, Min. Engr., Asst. Chief Engr., H. C. Frick Coke Co., Scottdale, Pa.

Jan 3, 1915

By J. Gross

G. H. Clevenger, Chairman of the Milling Methods Committee of the Institute, made the following introductory remarks: "Several years ago, a number of us felt that the time was ripe for a fundamental investigation of crushing and grinding. At the outset, it appeared that the first thing we needed was a yardstick whereby grinding efficiency could be measured. "The U. S. Bureau of Mines has been most generous in cooperating with the Milling Methods Committee in getting this work well under way. John Gross, who served the committee for a year as secretary, was placed in charge of the investigation and has been ably assisted by S. R. Zimmerley. "Messrs. Gross and Zimmerley will now present the Progress Report covering their work to date." PROGRESS REPORT J. Gross, Salt Lake City, Utah. -This has not been a simple problem, and when I tell you that we can give you a verbal report of what we have done during three years in five or ten minutes you will realize that we have necessarily done a great deal of work that does not appear in the final results. We have been in a virtually new field. Therefore, we have had to feel our way at each step so as to be fairly sure we were correct. The first thing that was realized in this problem was that unless we had some method of measuring what happened when we crushed an ore, we might as well quit and not go further. We realized that if we tried to measure the results of crushing or grinding by the ordinary sizing test we would not get anywhere. We could sieve down to say 200 mesh fairly accurately; we could even go one sieve size further with a little less accuracy, and then we were absolutely lost, unless we went into microscopic measurement. Even this latter method would not have been satisfactory. The question was: what were we going to measure? If we were going to measure size, we might measure one particular dimension of a particle and be off 200 per cent. or more. If we tried to measure surface, we would have something tangible; the surface of a particle, no matter how irregular, is something definite, far more definite than any linear dimension would be. Moreover, if we could measure surface, we would immediately have a method of checking the Rittinger law as to whether it was or was not correct. So the first thing that we sought was a method whereby we could measure surface on crushed particles or, we will say, on particles; it would not make any difference whether crushed or not. It then became a question of what kind of particles we were going to use in obtaining a method for surface measurement. We realized that it would be rather a com-

Jan 1, 1928

By R. G. Jones, W. L. Kinney, R. E. Schilson, R. S. Wilson, G. A. Clark, H. Suralo

This paper discusses the results of the Fry conventional or cocurrent in situ combustion test, which was conduct-ed in a 3.3-acre inverted five-spot. The depth of the formation was between 880 and 936 ft; the oil had a specific gravity of 28.6° APl and a viscosity of about 40 cp at the reservoir temperature of 65°F. Preceding the combustion test, air injection tests were conducted which, in conjunction with geological studies, were used to evaluate the characteristics of the reservoir. Combustion was initiated on Oct. 13, 1961, with ignition being accomplished by a 40 kw electrical heater. The test phase of the project ended on Oct. 1, 1963. During the test, the average air injection rate was 1,520.-000 scf/D. Throughout the test, production of all fluids— gas, oil, and water—was monitored. Cumulative oil production credited to the project was 100,586 bbl. The cumulative air-oil ratio was 11,500 scf/bbl oil, and the oxygen uti1ization efficiency was 87 per cent. INTRODUCTION The Marathon Oil Co. conducted a successful in situ combustion test, beginning Aug. 22, 1961, at the Fry unit, Crawford County, 111. The purpose of the project was to test the feasibility of cocurrent in situ combustion as a means of oil recovery in the Fry type reservoir. Interest in in situ combulstion as an oil recovery tool has been stimulated mostly by the existence of large reserves of heavy viscous crudes with low expected recovery, usually less than 10 per cent. These are the so-called unrecoverable reserves, and most combustion tests to date have been conducted in this type of reservoir.1-4 In contrast, the Fry combustion test was conducted in a reservoir with a relatively high gravity oil having a relatively low viscosity. This paper discusses the performance of the test. The geology of the reservoir and the field operations are discussed in separate papers.5 TEST SITE The Fry combustion test was carried out in a 3.3-acre inverted five-spot portion of a lenticular body of Robinson sandstone. Net sand was 50 ft thick, porosity averaged 19.7 per cent, oil saturation was 68 per cent of pore volume, and water saturation was 20 per cent of pore volume. The oil in place was estimated at 1,040 bbl/acre-foot, or 171,600 bbl within the 3.3-acre pattern. The water in place was 326 bbl/acre-foot, or 53,800 bbl. The oil has a specific gravity of 28.7° API and a viscosity of 40 cp at the reservoir temperature of 65°F. AIR INJECTION PERFORMANCE Air injection took place in two phases, the phases separated by ignition of the reservoir. In the pre-ignition phase, air injection tests were conducted in the summers of 1960 and 1961. These indicated that the Fry reservoir was confined, and a high return rate of injected air could be expected. This proved to be an outstanding characteristic of the project, as cumulative gas production was 95.3 per cent of the air injected. The difference between the cumulative air injected and the total gas produced can be largely accounted for by the quantity of air stored in the reservoir. An estimated 16 X 10 cf of air remained in the burned-out portion of the reservoir and an indeterminate amount of gas was stored in the unburned but pressurized reservoir. Hence, a complete material balance of gases would account for nearly 100 per cent of the air injected. At times during the test, the daily gas production rate was 98 per cent of the daily air injection rate. Fig. 1 presents the air injection history of the Fry com-

Jan 1, 1966

By Gary Mack, Donald E. Scheck

To help management keep key machines or processes at peak efficiency, draglines, shovels and blast hole drills have been equipped with micro- processor based monitors. The dragline and shovel monitors log operating data such as energy consumption, bucket position, cab rotation and the clock time of key elements within each cycle as well as maintenance information. The output consists of maintenance and management data such as the number of cycles, average cycle times, swing angles, depths and downtimes. Extensive experimentation has shown that a blast hole monitor can provide a log of each blast hole in which the hard to fragment zones are indicated. The logs are printed on board the drill and serve as the design basis for decked loading of the holes. As a result, the fragmentation of the overburden has improved markedly with the same or less powder consumption. This material is based upon work supported in part by the Bureau of Mines, Department of the Interior under Grant Numbers G1115006, G1115391 and G5 105006. Any opinions, findings, and conclusions or recommendations expressed in this publication are those of the authors and do not necessarily reflect the views of the Bureau of Mines, Department of the Interior.

Jan 1, 1983

ALEXANDER, JOHN S, M,'74-6,'81-3 *ASHBURNER, C A, M, '85-7 ASMUS, GEORGE, Dl, '78-80 BACON, D H , M, '00-2 BAYLES, JAMES C , M,'80-2,'91-3 P, '84-5, V,'88-9 BIRKINBINE, JOHN, M, '83-5, V, 'S7-J, P, '91-2 BLAIR, THOMAS S , M, '74-43 BLAKE, WILLIAM P , V,'71-4,'76-7,'05-6 *BLANDY, JOHN F, V,'71-3,'78-9 BLOW, A A, C, '05-7 BOWIE, A J , JR, V,'93-4 *BRAMWELL, J H , A1, '89-91 BROOKS, THOS B , M, '72 BRUNTON, D W , V, 97-8 BUCK, STUART M , M, '83-5 *BURDEN, JAMES A, V,'80-1 BURNHAM, WILLIAM, M, '82-4. CAMPBELL, H H, M, '93-5 CHAMBERLAIN, H S , V, '96-7 CHANNING, J PARKE, V,'08 CHRISTY, S B , V. '91-2, '07-8 CHURCH, JOHN A, M, '79-8l,'94-6, V, '07-8 COGSWELL, W B, V, '74-5,'86-7

Jan 1, 1910

By W. C. Maurer

A drilling-rate formula for roller-cone bits is derived from rock crater-ing mechanisms. This formula holds for "perfect cleaning", which is defined as the condition where all of the rock debris is removed between tooth impacts. Under these conditions, the drilling rate is directly proportional to the rotary speed and to the bit weight squared, and inversely proportional to the bit diameter squared and to the rock strength squared. Under imperfect cleaning conditions, such m those usually present in field drilling, re-grinding of the cuttings occurs under the bit, and the drilling rates fall below those for perfect cleaning. INTRODUCTION Drilling with a roller-cone bit consists of two fundamental operations: the formation of craters under the bit teeth, and the removal of the broken rock from the craters. To delineate the mechanisms of drilling, it is first necessary to understand the mechanisms involved in the formation of individual craters and then to relate these mechanisms to the over-all drilling operation. CRATER FORMATION When a bit tooth impacts a rock, the rock is elastically deformed until the crushing strength of the rock is exceeded, at which time a wedge of crushed rock is formed below the tooth1, 2 as shown in Fig. 1. As additional force is applied to the tooth, the crushed material is compressed and exerts high lateral forces on the solid material surrounding the crushed wedge.- en these forces become sufficiently high, fractures are initiated below the tooth and propagate to the free surface of the rock. The trajectories of these fractures intersect the principal stresses at a constant angle,'s4, as predicted by both Mohr's and Griffith's theories of failure. It has been experimentally 3,1 .6 that the volume of an impact crater, V,, varies as where Ec is the total energy imparted to the rock during the formation of the crater, and E is the threshold energy required before cratering is initiated (Fig. 2). Indexing experiments' have shown that, when there is a second free face to which the crater can fracture, a larger volume of material is removed and that above a transition zone there is a linear relationship between V, and Eo., which has a slope equal to the relationship for the single crater. The beneficial effect of indexing tends to offset the threshold energy requirement to a large extent, as shown in Fig. 2, and the relationship is closely represented by Rate-of-loading effects have been used to explain the nonproportional relationships usually obtained between drilling rate and rotary speed. The data4, in Fig. 3 show that the Ec/Vc parameter remained constant for craters produced with impact velocities ranging from 10 to 8,000 ft/sec. Since no rate-of-loading effects were detected over this wide range of loading rates, it seems improbable that there would be detectable rate-of-loading effects present in the small changes in impact velocities produced by varying the rotary speed. The writer has shown (Fig. 4) that, for craters produced in rock by impacting spheres, where X is the depth of penetration. Simon2 has shown that this relationship also holds for craters produced in rock by penetrating wedge-shaped chisels. The energy used in crater formation is

Arizona ROBERT W. THOMAS, Chairman 'ROBERT W. HUGHES, Vice-Chairman HUBERT O. WOODS, Secretary-Treasurer American Smelting & Refining Co., Hayden, Ariz. FRANK A, WARDLAW, JR., National Director THOMAS C. CHAPMAN P. D. I. HONEYMAN CARL H. COLE JOHN E. LANNING. JOHN H. DAVIS JOHN C. LOVING EDW. G. DENTZER ALBERT MENDELSOHN MORRIS G. FOWLER BRENT N. RICKARD H. C. HENRIECHARLES F. WILLIS Black Hills BANCROFT GORE, Chairman N. P. GOODRICH, First Vice-Chairman C. N. KRAVIG, Second Vice-Chairman C E. SCHMIDT,

Jan 1, 1941

By N. J. Themelis, A. W. Last

Autogenous grinding, broadly defined as the self-comminution of ore in a cylindrical tumbling mill without using auxiliary grinding media, was first applied in the early 1930s in the so-called Hadsel mills (Hardinge, 1955) as seen in [Fig. 1], mainly used in gold mines on the US West Coast. A number of these mills were built in sizes of 6.1 to 7.3 m (20 to 24 ft) in diameter and 1.5 to 1.8 m (5 to 6 ft) in length but were later abandoned because of operating and maintenance problems. Further development of autogenous grinding led to the Hardinge wet grinding mill, the Aerofall dry grinding mill, and others. A tabulation of all known autogenous and semiautogenous installations around the world was sent recently to the authors by G. J. Kushner of Koppers Metal Processing Division. In the size range of 7.3 to 9.8 m (24 to 32 ft) in diameter, it lists a total of thirty mills used in the processing of copper ores. An additional 120 units are listed in the same size range for the processing of iron ores and other minerals. The largest mills in existence are 11.1-m-diam X 4.6-m-long (36-ft-diam X 15-ft-long) mills in Minnesota, powered by 8950 kW (12,000 hp) motors and used in taconite beneficiation. It is indicative of the phenomenal growth of this technology that none of the above 150 mills was in operation prior to 1964. Grinding rates in autogenous mills are very sensitive to variations in ore breakage characteristics and even with fairly constant ore type can exhibit large fluctuations in throughput. Autogenous grinding also depends on availability of a fraction of the ore to be ground in the form of "competent" rock which can withstand to a certain degree the tumbling and impact action within the mill. If the ore is uniformly "soft", it will soon be crushed into small size particles and further grinding, which depends on the existence of large size rocks in the mill, becomes impossible. This apparently is the case with most copper ores, as discussed by Bachman, Last and Nabbs (1970) who described pilot testing of a large number of ore samples from Kennecott 's Chino Mines. Under circumstances where autogenous grinding is impractical, it is possible to realize most of its benefits by utilizing large grinding balls in place of the "competent" rock and practice what has become known in the mining industry as semiautogenous grinding. In semiautogenous grinding (SAG), a grate discharge tumbling mill is fed the product of a primary crusher or run-of-mine rock. Forged steel or alloy grinding balls, in proportions ranging from 3.10% of the total volume in the mill, are used to complement the self-grinding ability of the ore. Semiautogenous circuits can be either designed for single-stage closed-circuit grinding which produces a final product for recovery processes, or for two-stage operation in which the SAG mill product is fed to a second stage of grinding in a conventional ball mill.

Jan 1, 1981

By G. F. Comstock

The subject of transverse fissures in steel rails has been discussed very thoroughly in recent years from various points of view and the final opinions expressed may be roughly classified into two groups: That these fissures are the result merely of fatigue of the steel and are independent of the quality of the metal. That the quality of the metal and the mill practice must have something to do with them. The first point of view was most ably and logically presented by J. E. Howard,' and in the discussion of this paper, as well as in several other preceding publications, Dr. P. H. Dudley, Consulting Engineer of the New York Central Lines, has strongly defended the other opinion.2 The steel makers of the country have naturally rallied under Mr. Howard's banner, hoping to place the entire blame for these failures upon the railroads with their high wheel-loads, hard steel, heavy traffic, and sometimes inferior road-bed; while the railroad men, on the other hand, have been busy accumulating statistics and records in the effort to find an actual reason in the rails themselves that would account for the seemingly haphazard occurrence of transverse fissures under varying conditions and at widely separated points. The view that these fissures were due merely to normal fatigue under alternating stresses seemed most reasonable to the writer until quite recently, because from work done in this laboratory, as well as the work of other investigators, of the problem, no structural differences, in the vast majority of cases, were found between metal at the nuclei of transverse fissures and metal at similar positions in the same rails or in other rails that had not developed fissures. Within the last year or two, however, since the method of etching polished sections for the microscope with a oupric chloride solution has been tried systematically on lengthwise sections, passing through the nuclei of transverse fissures, evidence

Jan 1, 1920

By Eugene Guccione

Fear of inflation is one important reason for the rebirth of interest in the stock market, and natural resources are the best investment in an era of rapid inflation.

Jan 6, 1976

By Edward Keller

Someone some time ago remarked that some chemists still insist on telling us how to determine copper by the electrolytic method. The writer must confess that he believes that everything is not known definitely as yet as to how the exact amount of copper is determined in such material as purest commercial electrolytic copper. Some of us are not yet convinced that the pure copper atom is at all times deposited from an acid solution and that no oxygen or hydrogen will, under certain conditions, accompany it. Difficulties with impurer material are frequent, but erroneous results are not always apparent, since these do not speak so readily for themselves as those in the first case. By this is meant that if a chemist finds 100 per cent. of copper in electrolytic copper it is pretty certain that he perceives the result to be wrong, while if he finds 0.1 per cent. of copper too much in a very impure material he is far more likely to be unconscious of his error. These, in a wide experience, are frequent happenings and the writer has always looked with interest to publications on this subject. In turn, he feels justified in giving to others a few of his own observations. Up to a few years ago there were two methods of electrolytic copper assay in technical use, which had in common, that the amount of copper deposited on the cathode did not exceed 2 g., and they differed in that, for the one a large sample1 (20 to 80 g.) was taken from the general sample by a splitting device, dissolved, and from the solution a small portion (to contain 1 or 2 g. of copper) measured out for the electrolytic deposition; in the other method the copper was determined, generally in 1-g. portions of the separated coarse and fine portions of the sample, all of which passed a 16 or 20-mesh screen, being accomplished by a 40-mesh screen,

Jan 1, 1914

By Edward Keller

THE ASSAY FOR COPPER. SOMEONE some time ago remarked that some chemists still insist on telling us how to determine copper by the electrolytic method. The. writer must confess that he believes that everything is not known definitely as yet as to how the exact amount of copper is determined in such material as purest commercial electrolytic copper. Some of us are not yet convinced that the pure copper atom is at all times deposited from an acid solution and that no oxygen or hydrogen will, under certain conditions, accompany it. Difficulties with impurer material are frequent, but erroneous results are not always apparent, since these do not speak so readily for themselves as those in the first case. By this is meant that if a chemist finds 100 per cent. of copper in electrolytic copper it is pretty certain that he perceives the result to be wrong, while if he finds 0.1 per cent. of copper too much in a very impure material he is far more likely to be unconscious of his error. These, in a wide experience, are frequent happenings and the writer has always looked with interest to publications on this subject. In turn, he :feels justified in giving to others a few of his own observations. Up to a few years ago there were two methods of electrolytic copper assay in technical use, which had in common, that the amount of copper deposited on the cathode did not exceed 2 g., and they differed in that, for the one a large sample 1 (20 to 80 g.) was taken from the general sample by a splitting device, dissolved, and from the solution a small portion (to contain 1 or 2, g. of copper) measured out for the electrolytic deposition; in the other Method the copper was determined, generally in 1-g. portions of the separated coarse and fine portions of the sample, all of which passed -a 16 or 20-mesh screen, being accomplished by a 40-mesh screen,

Jan 8, 1913

By C. Y Garber

Interest in the sink-float plant operated by the Bunker Hill and Sullivan Mining and Concentrating Co. at Kellogg, Idaho, has recently been largely confined to the preparation and reconditioning of the galena medium. The ease and simplicity of these operations in a concentrating mill producing a flotation lead concentrate has occasioned enough comment to possibly justify a brief description. The subject is treated from an operator's viewpoint only; no effort is made to discuss it in a theoretical or technical manner. Plant at Bunker Hill The sink-float plant was first operated in December 1941. It was built for the purpose of quickly increasing mill capacity with a minimum of equipment. It is the Huntington-Heberlein sink and float process using galena medium and is operated under license from the Sink and Float Corporation, New York city. Lead-zinc ore from the Bunker Hill mine is crushed to pass a 1¼ in, screen and is then washed and separated on vibrating screens having 4½-mm. round-hole punched plates. The screen under-size, amounting to 33 per cent of the mill feed, is dewatered and sent to the ball-mill feed bins. The screen oversize goes to the sink-float plant, where 50 per cent of the oversize, or 33 per cent of the mill feed, is float, which is washed free of medium and disposed of as a tailing. The remaining 33 per cent is sink, which, after washing, is crushed to pass 36 in. and joins the under-size in the ball-mill feed bins. The normal plant capacity is 1700 tons per 24 hr. About 1.7 per cent of the total lead and 3.3 per cent of the total zinc in the feed is lost with the sink-float tailings (Table I). More complete descriptions of the H. H. process and Bunker Hill sink-float plant are given in several recent articles, listed at the end of the paper. Preparation OF Medium at Bunker Hill New medium is made intermittently as needed, usually 12 hr. in each 6 days. A small portion of the current mill flotation concentrate is diverted from the lead cleaner cells to a I-in. vertical sand pump and pumped to a two-cell 36-in. Agitair flotation machine, where NaCn and ZnSO4 are added and the grade of concentrate is improved as shown in Table 2. Tailing from this operation is returned by gravity

Jan 1, 1947

By C. Y. Garber

Interest in the sink-float plant operated by the Bunker Hill and Sullivan Mining and Concentrating Co. at Kellogg, Idaho, has recently been largely confined to the preparation and reconditioning of the galena medium. The ease and simplicity of these operations in a concentrating mill producing a flotation lead concentrate has occasioned enough comment to possibly justify a brief description. The subject is treated from an operator's viewpoint only; no effort is made to discuss it in a theoretical or technical manner. Plant at Bunker Hill The sink-float plant was first operated in December 1941. It was built for the purpose of quickly increasing mill capacity with a minimum of equipment. It is the Huntington-Heberlein sink and float process using galena medium and is operated under license from the Sink and Float Corporation, New York city. Lead-zinc ore from the Bunker Hill mine is crushed to pass a 1¼ in, screen and is then washed and separated on vibrating screens having 4½-mm. round-hole punched plates. The screen under-size, amounting to 33 per cent of the mill feed, is dewatered and sent to the ball-mill feed bins. The screen oversize goes to the sink-float plant, where 50 per cent of the oversize, or 33 per cent of the mill feed, is float, which is washed free of medium and disposed of as a tailing. The remaining 33 per cent is sink, which, after washing, is crushed to pass 36 in. and joins the under-size in the ball-mill feed bins. The normal plant capacity is 1700 tons per 24 hr. About 1.7 per cent of the total lead and 3.3 per cent of the total zinc in the feed is lost with the sink-float tailings (Table I). More complete descriptions of the H. H. process and Bunker Hill sink-float plant are given in several recent articles, listed at the end of the paper. Preparation OF Medium at Bunker Hill New medium is made intermittently as needed, usually 12 hr. in each 6 days. A small portion of the current mill flotation concentrate is diverted from the lead cleaner cells to a I-in. vertical sand pump and pumped to a two-cell 36-in. Agitair flotation machine, where NaCn and ZnSO4 are added and the grade of concentrate is improved as shown in Table 2. Tailing from this operation is returned by gravity

Jan 1, 1947

By Fun-Den, Wang

An open pit rock structure usually contains geological planes of weakness. They are formed by joints, faults, bedding planes, fractures, and cleavages. Rock slope failures often occur in the form of sliding of rock blocks along a plane or several planes. Such a rock block has three- dimensional geometry, and consequently a vector method of analysis has been found useful in evaluating its stability. Rock blocks of a tetrahedral shape with one free surface have been analyzed by Londe. Rock blocks formed by sets of parallel joints have been analyzed by Wittke. The present chapter presents the analysis of rock blocks formed by planes of weakness of arbitrary orientation, and of three-dimensional geometry.

Jan 1, 1972