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NEW MEMBERS The following list comprises the names of those persons who became members during the month of December, 1913: Members ADAMS, CHARLES W., JR., Smelting, American Smelting & Refining Co., Murray, Utah. ARMSTRONG, FRANK H., Mechanical Engineer Vulcan, Mich. ARNOLD, JOHN B., Attorney-at-Law Room 840,200 5th Ave., New York, N. Y. BOYS, HARRY R., Mining Paso Robles, Cal. CAVANAGH, JOHN L., Min. Engr., Nova Scotia Steel & Coal Co.; Wabana, Newfoundland. CHOW, KAI CHI, Student 1120 Amsterdam Ave., New York, N. Y. CLARK, ARTHUR B., Mining Engineer 72 Harbour Bldg., Butte, Mont. CURLEY, JOHN R., State Mine Inspector P. 0. Box 633, Leadville, Colo. DAVIS, ANGUS W., Mining Engineer, Consolidated Mining & Smelting Company of Canada, Trail, B. C., Canada.

Jan 1, 1914

NEW MEMBERS The following list comprises the names of those persons who became members during the period Mar. 10 to Apr. 10, 1914. Members ADAMS, DANIEL T., Mining 1416 Westminster Bldg., Chicago, Ill. BLICKENSDERFER, FRANK CLEVELAND, Sampler Morenci, Ariz. CAVERS, THOMAS WILLARD, Met Copperhill, Tenn. CLAPP, CHARLES H., Geol Univ. of Arizona, Tucson, Ariz. CLARK, PATRICK T., Pres Bullwhacker Mine, Spokane, Wash. COMSTOCK, HARRY, Mining Mineville, Essex Co., New York. COOKE, JOSEPH H., Gen. Mgr Mono Dev. Co., Gisborn, Tooele Co., Utah. COOPER, DEXTER P., Civ. and Hyd. Engr 101 Park Ave., New York, N. Y. DEGOLYER, E. L., Geol Apartado 269, Tampico, Mexico. DESHAYES, ERNEST F., Mine Foreman Baldy, N. M. DOUGLASS, WILLIAM CRAWFORD, Min. Engr 107 Leonard Hotel, Butte, Mont.

Jan 5, 1914

(There are 67 Student Chapters in the United States, 2 in Canada ) University of Alabama-University, Alabama, Mining & Metallurgical Society, James W Stewart University of Alaska-College, Alaska, Mining Society of the University of Alaska, Guinn Metzger University of Arizona-Tucson, Arizona, Miners' Society, A E Gorum Birmingham-Southern College-Birmingham, Alabama, Pick & Hammer Club, Daniel D Arden University of British Columbia-Vancouver, British Columbia, G M Dawson Club, A Gunning University of California-Berkeley, California, Mineral Technology Association, Bernard York Carnegie Institute of Technology-Pittsburgh, Pennsylvania, The Metals Club, Gerhard Derge Case Institute of Technology-Cleveland, Ohio, Pick and Shovel Club, Gerald M Cover City College of New York-New York, New York, CCNY Student Chapter AIME, Daniel T O'Connell

Jan 1, 1954

By Robert H. Richards

The object of this paper is to enlarge the field of settling velocities treated by me in my former papers, Close Sizing Before Jigging, and Sorting Before Sizing.' There seemed need of work both on coarser and on finer sizes. Messrs. A. Sidney Warren and M. L. Nagel undertook the investigation for the coarser sizes, from 12.85 mm. down to 2.05 mm. diameter. Their work, because of the closer spacing and because of the increased number of observations and consequent stronger averages, called for a revision of the former work. Messrs. G. A. Barnaby and Ralph Hayden undertook this part of the work, from 2.49 mm. down to 0.28 mm., which is near the limit of sifting. Their work covers the ground much more minutely and comprehensively than the former papers. There remained to be covered the portion of the field too fine for investigation with the sieves. This part, covering the range of grains from 0.48 to 0.03 mm., was undertaken by Mr. E. S. Bardwell. For it he used the elutriation method. 11. The First, or Coarsest, Section of the Field. The first desideratum in this investigation was a series of sieves for sizing the sands preparatory to making the tests. It became necessary, therefore, to decide on a sieve-scale for this purpose. A sieve-scale is a series of sieves in which the dimensions of the holes of the successive members form a series increasing by a definite ratio. There are three natural sieve-scales: 1, progressing by doubling the width of the hole for each suc-

Jan 1, 1908

By S. C. Miller, R. D. Carter, H. G. Riley

A merhod based on theoretical considerations is presented for estimating srabilized performance of gas wells from data obtained from short-term flow tests. Calculated results are presented for II gas wells evaluated with four-to five-hour flow tests. For seven wells on which comparison war available, satisfactory agreement between measured and calculated performance was obfained. Although the final results are favorable, diflerences in flow capacity k,h obtained from flow tests conducted at different rates were observed. This inconsistency did not seriously affect the calculated results in the test wells. INTRODUCTION In order to negotiate the sale of gas from newly drilled gas wells and to plan otherwise for the orderly exploitation of gas reserves, it is necessary to predict the long-term withdrawal rates physically possible for each well. Since low permeability gas wells do not stabilize except after a lengthy period of flow, the operator is unable, without wasting large amounts of gas, to measure directly the stabilized performance prior to contracting for sale of the gas and connecting the well to a pipeline. This problem is not new, but is one which has recently received considerable attention because of the rising importance of natural gas in the economy and new emphasis on the development of gas reserves. Thus, a method is needed for estimating stabilized performance from information obtained from short-flow tests in which stabilized flow is not attained. Such short tests would be suitable for wells not connected to a pipeline since the tests require only small amounts of time, labor and production of gas. Other requirements of such a method are short calculation time and use of a model of the gas well and its reservoir drainage area which is as consistent as possible with present theoretical, experimental and field knowledge. The method presented herein was developed in an effort to meet these requirements. It has been named the "two-Row method". In addition to the section on the evelopment and description of the method, a description of field application of the method and a discussion of well test results are in- cluded. Finally, there is a section on an additional consideration prompted by some of the test results. The Appendix contains certain details of the derivation of the two-flow method and a sample calculation for Test Well 1. MATHEMATICAL DEVELOPMENT AND DESCRIPTION OF THE METHOD The method is based upon the following assumptions: (1) The system can be described as a finite cylindrical gas reservoir produced by a single cylindrical well located at the center of the reservoir. Flow is strictly radial. (2) Flow obeys Darcy's law except very near the well where a quadratic flow law is assumed. (3) The system is also characterized by a laminar flow skin factor. The use of a quadratic or turbulent flow term is more fully justified in Refs. 1 and 2. When such a system is opened at the wellbore to constant rate flow from an initial condition of uniform pressure distribution, the radius or drainage quickly passes through the non-Darcy flow and "skin effect" region imrnediately surrounding the well. After the drainage radius or "transient" has passed beyond this region and before it reaches the external boundary, drawdown can be expressed by'." After the drainage radius reaches the external boundary, stabilized flow exists and drawdown is given by'," Eqs. 1 and 2 form the basis for the method. The validity of these equations will now be examined in the light of some well test data previously available. It has been found by actual measurements on gas wells that the relationship between drawdown and flow rate, that is the performance curve, is closely approximated by' Eq. 3 is a straight line on log-log graph paper. Eq. 2 and, for any lixed value of time, Eq. 1 plot as a curve which is slightly concave upward on log-log graph

By L. D. Clark

Pay-out, the ratio of capital expenditure to annual cash return can be very revealing when employed to gage the merits of a proposed mining venture. It is a quantity, numerically but not physically equal to a factor, by which cash return may be discounted to equal required investment. The fewer the pay-out years, the greater the average annual earning power of the project. Considerable material has been written in the past about the most important and difficult problem mining management faces in evaluation of a property: the optimum production rate for the venture. What daily capacity should be adopted for the proposed operation that it may be the most rewarding to those underwriting it? H.C. Hoover has said, "The most important objective is the least cost per ton mined; and minimum working costs can only be gained by the most intensive production." Theoretically and practically, the quicker the ore is removed, the more profitable should be the venture because of the saving in fixed charges, the consequent lower unit costs and the interest unearned by the profit from the unmined ore. Yet, such performance will be tempered by policy and a restraint imposed by the ratio of capital expenditure to annual cash return. * The 'Annual net profit after taxes but including depreciation and depletion allowances. former will be based upon an analysis of the economic climate anticipated, with its many ramifications including the demand for the product, prices, tax laws, the conduct of labor and related wage scales. The variables and intangibles involved in estimating the optimum production rate for a mine do not readily lend themselves to mathematical relationships or equations and the problem can become complex indeed. The more readily applicable and relatively faster appraisal method will be used which, in addition to being reasonably adequate in most cases, will serve as a guide, should a more expanded and detailed estimate be desired. The first question is what is company policy with respect to the rate of investment return? This will depend on that significant period of time referred to as the pay-out, which can be defined as the estimated interval during which capital investment is returned through application of annual cash return to its recovery. Expressed another way, it is the ratio of capital expenditure to the annual cash return. For example, if the capital investment is $C and the annual cash return is $A (after Federal tax), then $C/$A = the pay-out period in years. That is to say, if $C = $1,500,000 and $A = $214,000, then which is the mathematical expression for the period of time it takes to recover the funds originally invested in a mining enterprise. The relationship yp = $C/$A can be useful — an expression which requires only the significant tern yp to be fixed; in other words, the long range investment policy. For what is the pay-out period but an expression of that policy in terms of the speed, within practical limits, with which an individual or company may want to recover capital expenditure? A policy by which pay-out will be fixed is imperative. This ratio must lie within relatively narrow limits governed by what is economically feasible. It cannot be too low, rendering the ratio impractical or absurd, nor too great, delaying the return of capital for reinvestment beyond a commensurate interval. What pay-out will govern this policy? Federal corporate taxes would seem to present the best measure. Consider mining in Canada where new mines are exempt from Federal income tax for 3½ yr from date of first production. This exemption expressly allows The Act states: "A corporation . . . i s not required to include the profits . . . for the period of 36 months commencing with the day on which the mine came into production in reasonable commercial quantites." a company to recover capital expenditure during the early stages of production. With such an incentive, it is logical to strive for as much capital redemption as possible during this period. Thus, 3½ yr would be a major contribution to the pay-out in Canada. Furthermore, it generally takes from 1½ to 3 yr to develop

Jan 1, 1963

By W. D. Ludemann, J. E. Dor, L. A. Shepard

Recovery of the creep resistance of 99.99 pct pure Al was studied at temperatures 540°, 573°, 600°, and 611°K. Poly-crystalline specimens crept under a stress of 950 psi to a strain of 5.5 pct were allowed to recover for periods of from 1 min to 16 days under a residual stress of 4.4 psi. Increased creep rates upon reapplication of the 950 psi stress evidenced softening of the material. The activation energy for the recovery process was found to be 64,000 cal-per mol. THREE major observations have clearly revealed that the creep of polycrystalline aluminum above about 510°K is controlled by the rate of climb of jogged edge components of dislocations past the barriers impeding their motion: 1) Extensive polygonization, characteristic of climb, takes place during creep.'9' 2) The activation energy for creep equals the estimated activation for self-diffusion. 3) The stress law for creep coincides with theoretical predictions based on the climb mechanism.4,5 The decreasing rate during the primary stage of creep must be ascribed to the introduction of additional barriers to the glide of dislocations. During secondary creep the density of barriers must remain constant, the rate at which new barriers are formed being equal to the rate at which they recover. For this reason the nature of the barriers, and their rates of formation and recovery, are significant to a complete understanding of creep. It is proposed here to study the kinetics of the recovery of barriers to creep of high-purity polycrystalline aluminum in the climb range. The technique will consist of creeping a series of tensile specimens under a prescribed stress to given state following which the specimens will be recovered for various times and temperatures under approximately zero stress. The amount of recovery will be determined by comparing creep curves following the recovery. EXPERIMENTAL PROCEDURE The material used for this investigation was a 99.99 pct pure Al supplied by the Aluminum Co. of America in sheets of 0.001-in. thickness having an H-18 temper. The spectroscopic analysis of impurities gave Cu-0.004 wt pct, Fe-0.002 pct, Si-0.001 pct, others-0.000 pct. Creep specimens were machined with their tensile axes in the rolling direction of the sheets. Annealing in a molten potassium nitrate-potassium nitrite bath for 1 hr at 686K, followed by air-cooling, resulted in a grain diameter of from 0.22 to 0.27 mm. Creep machines were equipped with constant stress lever arms of the type suggested by Fullman, Carreker, and Fishher,' which maintained the applied stress to within 0.04 pct of the reported value of 950 psi. Creep strains were calculated from extensions over a 6-in. gage section to the nearest 3 X 10"5. The temperatures of testing, 540°, 573", 600°, and 611°K, were obtained by immersing the specimen and extensometer assembly in a molten KNO2-KNO3 bath. Temperatures during the periods of creeping could be maintained constant to within better than ± 1°K. Correlation of creep data at different temperatures of testing demanded the use of an appropriate temperature-compensated time 8. Previous investigations have shown that the creep of pure metals at temperatures above about 0.55 of their melting temperatures can be correlated by the functional relationship3" e=f(0), s= const. [1] where e = total plastic strain during creep f = a function that depends on stress a = stress 6 = te-Q/RT = a temperature-compensated time t = time under test Q = 35,500 cal per mol = activation energy for creep R = gas constant, and T = absolute temperature In the present series of tests all of the creep was conducted under the same stress of 950 psi. To provide identical initial conditions for the recovery all specimens were first crept to a temperature-compensated time of = 47 X 10-14 min which resulted in a creep strain of 0.055 + 0.005 regardless of the creep temperature. During recovery, a small stress (4.4 psi) was permitted to remain upon the specimens to maintain tautness in the pulling assembly and to prevent mechanical damage to the soft specimen. After each recovery period, the full stress was reapplied and

Jan 1, 1961

Molybdenum was first discovered on the Continental Divide in Colorado in 19 11; and in 1918 two companies, one of which was the Climax Molybdenum 250 stpd operation, were producing. In 1919 both mines were shut down, but in 1924 the Climax mine reopened and has been producing ever since. Over the years mine production has been increased until today the mill capacity is 48,000 stpd (43,500 mtpd). From 1935 to 1973 all ore was mined underground by block caving with scraper slushing. Since 1973 ore production has been supplemented by open pit mining. As demand for molybdenum increased, upper levels of the mine became worked-out and new lower levels were opened. In late 1972 production began on the 600 level with the mined ore being primary and secondary crushed in the new No. 4 crushing station. Ore is hauled to the primary in 10 st (9. 1 mt) capacity Granby side-dump cars. The trains, carrying 200 st (181 mt), pass throug.h a dump block and discharge directly into a 54 in. by 84 in. Traylor TC gyratory crusher set at an 8 in. (203 mm) open side setting. Since the ore has already been broken to a size which will pass the finger raises in the mining areas, there are only rare oversize boulders to block the crusher opening. Selection of the crusher was based more on its capacity than its receiving opening, as is often the case with gyratories used in open pit operations. Nominal and maximum capacities for this unit are 1250 and 1800 stph (1 135 and 1635 mtph). The ore is often wet and sticky, so a mechanical car cleaning device is utilized. Under the crusher is a 300 st (270 mt) ore pocket equipped with a 60 in. (1.52 m) wide inclined conveyor which delivers the ore to a 2,000 st (1800 mt) surge bin. Near the head pulley of this conveyor is a wood picking station whe re tramp wood and steel are put into cars for removal. Two more apron feeders remove the primary crushed ore from the surge bin, and impact belts and conveyor belts transport the ore to the crushing station for screening ahead of two 7 ft (2. 13 m) standard crushers. Secondary crushed ore is then conveyed to the surface for tertiary crushing. The entire underground crushing room containing the primary, two secondaries, screens and related equipment is serviced by a 100 st (90 mt) crane with a smaller capacity fast hook. A control room is situated so the operator can observe all the crushing operations. Rotoclone collectors remove dust from the entire operation (including the fumes from the zinc pots) and discharge it to the final product belt. The Rotoclones are vented through drifts and raises to the surface. Shop facilities, electrical gear, and lubrication equipment are located

Jan 1, 1978

By Arthur Hockman, Howard W. Jaffe, Howard F. Carl

THE problem of establishing practical techniques for evaluating natural corundum ores arose from the desire to improve the existing purchase specifications for crystal corundum procured by the Federal Government for stockpiling. Such purchase specifications had been developed for a large number of products, including minerals and ores considered to be critical and strategic because of the country's dependence on foreign sources of supply. It will be apparent to anyone considering the subject of purchase specifications that the determination of whether or not a submittal meets the specifications is a prime consideration. Consequently satisfactory specifications for minerals should not only incorporate physical, chemical, and mineralogical requirements that are practical but should also establish the means by which both buyer and seller can be assured that materials meet stated requirements. The Munition Board's Material Purchase Specification for corundum, dated February 1954, covered two grades of crystal corundum intended primarily for abrasive use. It established limits of size and of composition: Grade A, at least 95 pct +3 mesh and 92 pct alumina (Al2O3); Grade B, at least 95 pct +7 mesh and 90 pct alumina. It also specified that "of each lot of material, not less than 95 percent shall have a hardness equal to or exceeding nine (9) on Moh's Scale." These specifications were unsatisfactory for several reasons. They did not allow the purchase of lower grades of corundum being used by industry. Neither did they specify exactly what a corundum ore actually was, except that it should contain so much alumina and have a certain scratch hardness. Nothing was said about the abrasive qualities of the material that was to be purchased primarily for abrasive use. During the period 1946 to 1948 supplies of the two grades defined in the then current specifications continued to decrease. Early in 1949 industry recommended that the finer sizes of corundum known as "crystal corundum concentrates," ¼-in. to 16-mesh material, be added to the stockpile specifications and that lower grades of crystal corundum also be included. In the spring of 1949 the National Bureau of Standards and the Bureau of Mines were asked by the Munitions Board to cooperate in developing more practical specifications for grading natural corundum ores. The National Bureau of Standards indicated that it would investigate physical test methods for determining abrasive characteristics of corundum, and the Eastern Experiment Station of the Bureau of Mines at College Park, Md., stated that it would consider the mineralogical aspects of the problem. Members of these two agencies, in consultation with representatives of the Office of Materials Resources of the Munitions Board and of the Federal Supply Service of the General Services Administration, and with the cooperation of industrial suppliers and consumers of abrasive corundum, developed, in a fairly short time, more practical specifications for six grades of corundum ores. The new specifications, issued January 1951, were for corundum, natural crystals, as by this time high-purity synthetic corundum was available and widely used. Nevertheless, for certain specific applications the natural product was preferred over the synthetic material. It has been extensively used in forming snagging wheels. It breaks into more or less cubical particles, whereas artificial corundum fractures into splintery shapes. These produce irregular cuts in glass, while the more symmetrical shape produces a more uniform ground surface. For this reason natural corundum has been preferred for certain types of lens grinding. Pure corundum is clear, colorless alpha alumina (A12O3), but it is rarely found in nature in this form. The most important gem varieties include the blue and yellow sapphires and the ruby. Minute impurities of titanic oxide impart a blue color,

Jan 1, 1955

By M. L. Lissberger

entirely suitable for certain uses and even possess advantages over present solders at high temperatures, but they are certainly not suitable for all uses. However, it is conceivable that their use in many places will effect a material conservation of tin. It will probably be wiser to use cadmium in solders where good results have been obtained than in brass where this element will be a source of possible trouble. F. F. Colcord,* New York, N. Y.-—The cadmium-lead solder has a higher conductivity than the lead-tin alloy, which ought to make it advantageous in the manufacture of electrical machinery. In the case of solder for tin-can work, one of the high officials of the American Can Co. said that he fully intended to .give it a good trial, and I do not believe they are going to reach hasty conclusions and run the risk of poisoning the public. Solder, Its Use and Abuse Milton L. LissbeRgeR,† New York, N. Y.—Solder is a mechanical mixture of tin and lead, a fact which is susceptible of very simple demonstration. A bar of solder of a grade even as low as 30 per cent. tin and 70 per cent. lead, passed through a buffing machine, will show a surface practically identical with that of a bar of second-quality or reclaimed tin. The buffings, on chemical analysis, will prove to be almost pure lead. According to the best practice, solder is made in the following manner. Virgin pig lead is first melted, and when it is thoroughly liquefied, virgin pig tin is added, together with a small amount of flux; the latter is for the purpose of bringing to the surface the so-called "liver," consisting of impurities that may have remained in either the lead or the tin as a result of incomplete refining. The combined material, when completely liquid, is thoroughly stirred for some hours, and is then cast into small pigs. Just before casting, and continuously during this operation, the molten metal yields dross, consisting largely of the oxides of lead and tin; this should be carefully skimmed off. After the pigs have cooled, they are taken to a smaller kettle, re-melted, and cast into the desired shape for use; or if wires, ribbons, etc., are to be made, the solder is cast into slugs suitable for extrusion and rolling. During this second operation, the skimming of dross should be even more carefully done than at first. Hand mixing has proved to be the only reliable method for the production of the best quality of solder, irrespective of its percentages of lead and tin. The best quality of solder is not necessarily that which contains the highest percentage of tin, but rather is that composition which performs best on the required piece of work. In order to produce

Jan 1, 1919

By Louis Ware

At the beginning of the present war, the United States faced the need to multiply its production of magnesium metal almost roo times within the shortest possible period. Urgently needed for construction of military aircraft, incendiary bombs, and other war uses, magnesium was at the top of the list of critical war materials. Up to that time only one Producer, employing Only One process and using One raw material in a single plant, represented this country's entire Production capacity. It was immediately desirable to plan a numher of plants located strategically over the Count'Y and to expand an industry more rapidly and to a greater extent than ever before had been attempted in so short a time. In '939 this American producer, the Dow Chemical .9 was making about 7 million Pounds of magnesium metal. It utilized as a raw material brine containing magnesium chloride, which was pumped from wells in Michigan. The logical ap-preach to certain increased production was to employ the same process on the same raw material in other sections of the United States. In Europe and other countries, different processes had been used and different raw materials were being utilized, but all of the skill and knowledge for production in this country was that of using the Dow type of cell and magnesium chloride as a raw material. The Government was testing other processes, but none then had been proved commercially feasible. Development OF New Process The International Minerals and Chemical Corporation had available a large quantity of magnesium ,chloride at Carls-bad, New Mexico, contained in the end liquor produced from its potash plant sillce the buildng of the plant in 1941 It was the only company mining commercially a magnesium ore and having anything like this quantity of magnesium chloride liquor. The magnesium ore, langbeinite, was being mined from the 800-ft. level for its content of potassium sulphate. This ore, a potassium magnesium sulphate, was reacted in a base exchange plant with potassium chloride to produce potassium sulphate, the discharge liquor having a high content of magnesium chloride. Three potash-producing companies are in the Carlsbad area, but only International has developed and is commercially producing the potash-magnesium ores. The lower, or No. 4, veil, in that area is potassium chloride ore, and is continuous through all the properties. The langbeinite, or No. 2, vein is minable at the International property, but occurs sparsely, if - at in other areas being mined in that vicinity. It is understood that most of the mag. nesium produced] by Germany comes from magnesium and magnesium-potassium ores obtained from potash mines. It was logical for International to pro-pose to make available this magnesium material and to offer its technical staff to design, construct and manage a magnesium plant for the account of the Government. The plan that finally was adopted by the

Jan 1, 1944

By D. C. Seidel, E. F. Fitzhugh

The Colorado School of Mines Research Foundation has developed a hydrometallurgical process for recovering nickel from oxidized ores, including both the iron-rich laterites and magnesium-rich, soft silicates. Known as the HSO-HTCP (Hydrothermal Sulfidization Oxidation-High Temperature Cementation in Pulp) Process, the system consists of feed preparation, sulfidization, oxidation, precipitation (cementation), and calcination and melting. This paper deals primarily with the sulfidization and oxidation phases of the process. Nickel sulfide ores lend themselves readily to concentration before smelting or pressure leaching, but neither of the major oxidized ore types — the iron-rich laterites and magnesium-rich, soft silicates-has been directly concentrated. Upgrading of crude ore in current practice is limited to cobbing lumps of the harder, lower grade rock (protore) from the softer, enriched silicate ore, because only a small percentage of the nickel in these ores occurs in discrete, contrasting mineral particles. There is no apparent mineralogical contrast in the superficial, iron-rich lateritic ores and, in fact, there is no firm assurance that the nickel atoms are within the lattices of the limonite minerals, or else irregularly adsorbed on the limonite. Within underlying, "rotten rock" silicates, occasional veinlets of gamierite and related nickeliferous silicates are found.' The gamierite veinlets, however, are only incidental, and the majority of the nickel atoms occur as erratic replacements of magnesium atoms in the micaceous chlorites of the ore mass. Pyrometallurgical processes account for nearly all the nickel currently being recovered from oxidized ores. Plants in New Caledonia, Oregon, Japan, Brazil and Greece smelt the ores to make ferronickels or a sulfide matte,2 and the ammoniacal leaching3 at Nicaro, Cuba, is preceded by a reducing roasL4 The only fully hydrometallurgical installation has been the politically ill-fated Freeport Nickel Co. enterprise, and this process was limited to lateritic ores.5p6 A hydrometallurgical technique which might handle both laterite and the typically richer silicate ores has been an enticing goal. The technique described here was developed at the Colorado School of Mines Research Foundation, Inc. on behalf of the Republic Steel Corp. It uses sulfur, heat, air and metallic iron to recover nickel from these ores. Inasmuch as the ores are treated as aqueous slurries, the cost of drying a plant feed that normally carried 30% or more moisture is eliminated, and there are no dust problems. Patent applications have been filed by Republic Steel Corp. on the procedures used for getting the metal into solution and on the subsequent recovery without liquids-so l ids separation of the slurry. It is referred to as the HSO-HTCP Process (Hydrothermal Sulfidization Oxidation — High Temperature Cementation in Pulp). THE PROCESS FLOWSHEET The results of various experimental studies have been combined to form the continuous flowsheet illustrated in Fig. 1. This flowsheet serves as a general pattern showing the sequence of the operations which include: (1) feed preparation; (2) sulfidization; (3) oxidation; (4) precipitation (cementation); and (5) calcination and melting. This paper deals primarily with the sulfidization and oxidation phases of the process. Feed preparation: In general, the silicate ores can be considered as earthy rather than hard, and preparation might be carried out before the addition of water in a trommel, or with water in a scrubber or log washer. This breaks up the earthy particles, allowing the hard lumps to be separated by screening. The oversize material from most ores is relatively barren and may be discarded. If there is occasion to control the iron-magnesium ratio in the plant feed, laterite, which normally breaks up readily into fine sized particles, and silicates are mixed in suitable proportions. The fine product slurry from the initial steps is mixed with elemental sulfur and fed to a conventional ball mill, although the amount of actual grinding that takes place is relatively small. The object of milling is to complete the disintegration to natural particle size and to produce an intimate mixture of the ore and elemental sulfur. This pulp, which is almost entirely -200 mesh, has a pH of about 7.0 and normally can

Jan 1, 1969

By J. W. Snavely

THE high tension belt conveyor is introducing a new and tremendously expanded era of low cost bulk material handling. High tension belt conveyors are generally those installations involving very long centers, high lifts, or drops, in which the belts are stressed up to their maximum tension values, and further, where the belt construction provides tension capacity far beyond what is possible with conventional belt constructions. With these high tension installations, the magnitude of the forces involved demands careful refinement of accepted design practice in order to achieve optimum balance of all factors.

Jan 3, 1951

By E. C. Franz

IN an effort to understand high cycle fatigue, as well as to study the mechanism of fracture in general, a number of researches have been undertaken whereby the fracture properties of a metal have been determined under conditions of repeated cyclic prestsain employing high strains with relatively few cycles. Liu, Lynch, Ripling, and Sachs' showed the effect of a definite strain history consisting of small increments of tensile or compressive prestrain upon the flow and fracture characteristics of metals after a limited number of repeated loadings. Liu and SachsW emonstrated by cyclic straining of the aluminum alloy 24s-T4 that the least reduction in ductility was obtained when the compressive stress was nearly equal to the tensile stress. Attempts have been made to correlate the reduction in ductility under such prestraining with conventional fatigue data, as was done by Liu and co-workers.' The following research was conducted to determine if such reduction as has been noted could be correlated with conventional fatigue data or whether the effect resulted from causes extraneous to the basic mechanism of fatigue failure such as strain aging. It might be pointed out in this connection that the reduction in ductility has occurred in alloys that might be susceptible to strain aging or similar behavior, such as alloy 24s-T4. In the present investigation, the test specimens were prepared from the aluminum alloy 75s-T6 and annealed electrolytic copper. Alloy 75s-T6 is a complex precipitation hardenable alloy, whereas electrolytic copper would not be expected to exhibit strain aging effects, particularly at room temperature. The experimental procedure consisted of prestraining test specimens repeatedly by increasing tension in small increments to various total prestrain values, followed by testing to failure in a tension test. The prestraining as well as the final testing was performed on a tensile machine using special fixtures9 0 maintain concentricity of the specimens. After prestraining, the specimens were unloaded as rapidly as possible and immediately reloaded so that the total time lapse between strains was held to a few seconds. Logarithmic or true natural strain was used in all calculations. The results are presented in Figs. 1 and 2. In Fig. 1, the retained ductility, area after prestrain area at fracture is plotted against the tensile prestrain, initial area area after prestrain as abscissa with the number of stress reversals executed to attain the required prestrain plotted as parameter. Several specimens that fractured during

Jan 1, 1955

Trans No. Place Date Vol. 1. Wilkes-Barre, Pa. May, '71 1 2. Bethlehem, Pa. Aug., '71 1 3. Troy, N. Nov., '71 1 4. Philadelphia, Pa. Feb., '72.. 1 5. New York, N. Y. May, '72 1 6. Pittsburgh, Pa. Oct., '72 1 7. Boston, Mass. Feb., '73 1 8. Philadelphia, Pa. May, '73 2 9. Easton, Pa. Oct., '73 2 10. New York N. Y. Feb., '74 2 11. St. Louis, Mo. May, '74 3 12. Hazleton, Pa. Oct., '74.. 3 13. New Haven, Conn. Feb., '75 3 14. Dover, N. J. May, '75 4 15. Cleveland, Ohio Oct., '75 4 16. Washington, D. C. Feb., '76 4 17. Philadelphia, Pa. June, '76 5 18. Philadelphia, Pa. Oct., '76 5

Jan 1, 1944

Trans No. Place Date Vol. 1. Wilkes-Barre, Pa. May, '71 1 2. Bethlehem, Pa. Aug., '71 1 3. Troy, N. Nov., '71 1 4. Philadelphia, Pa. Feb., '72.. 1 5. New York, N. Y. May, '72 1 6. Pittsburgh, Pa. Oct., '72 1 7. Boston, Mass. Feb., '73 1 8. Philadelphia, Pa. May, '73 2 9. Easton, Pa. Oct., '73 2 10. New York N. Y. Feb., '74 2 11. St. Louis, Mo. May, '74 3 12. Hazleton, Pa. Oct., '74.. 3 13. New Haven, Conn. Feb., '75 3 14. Dover, N. J. May, '75 4 15. Cleveland, Ohio Oct., '75 4 16. Washington, D. C. Feb., '76 4 17. Philadelphia, Pa. June, '76 5 18. Philadelphia, Pa. Oct., '76 5

Jan 1, 1936

Trans No. Place Date Vol. 1. Wilkes-Barre, Pa. May, '71 1 2. Bethlehem, Pa. Aug., '71 1 3. Troy, N. Nov., '71 1 4. Philadelphia, Pa. Feb., '72.. 1 5. New York, N. Y. May, '72 1 6. Pittsburgh, Pa. Oct., '72 1 7. Boston, Mass. Feb., '73 1 8. Philadelphia, Pa. May, '73 2 9. Easton, Pa. Oct., '73 2 10. New York N. Y. Feb., '74 2 11. St. Louis, Mo. May, '74 3 12. Hazleton, Pa. Oct., '74.. 3 13. New Haven, Conn. Feb., '75 3 14. Dover, N. J. May, '75 4 15. Cleveland, Ohio Oct., '75 4 16. Washington, D. C. Feb., '76 4 17. Philadelphia, Pa. June, '76 5 18. Philadelphia, Pa. Oct., '76 5

Jan 1, 1942

Trans No. Place Date Vol. 1. Wilkes-Barre, Pa. May, '71 1 2. Bethlehem, Pa. Aug., '71 1 3. Troy, N. Nov., '71 1 4. Philadelphia, Pa. Feb., '72.. 1 5. New York, N. Y. May, '72 1 6. Pittsburgh, Pa. Oct., '72 1 7. Boston, Mass. Feb., '73 1 8. Philadelphia, Pa. May, '73 2 9. Easton, Pa. Oct., '73 2 10. New York N. Y. Feb., '74 2 11. St. Louis, Mo. May, '74 3 12. Hazleton, Pa. Oct., '74.. 3 13. New Haven, Conn. Feb., '75 3 14. Dover, N. J. May, '75 4 15. Cleveland, Ohio Oct., '75 4 16. Washington, D. C. Feb., '76 4 17. Philadelphia, Pa. June, '76 5 18. Philadelphia, Pa. Oct., '76 5

Jan 1, 1941

Trans No. Place Date Vol. 1. Wilkes-Barre, Pa. May, '71 1 2. Bethlehem, Pa. Aug., '71 1 3. Troy, N. Nov., '71 1 4. Philadelphia, Pa. Feb., '72.. 1 5. New York, N. Y. May, '72 1 6. Pittsburgh, Pa. Oct., '72 1 7. Boston, Mass. Feb., '73 1 8. Philadelphia, Pa. May, '73 2 9. Easton, Pa. Oct., '73 2 10. New York N. Y. Feb., '74 2 11. St. Louis, Mo. May, '74 3 12. Hazleton, Pa. Oct., '74.. 3 13. New Haven, Conn. Feb., '75 3 14. Dover, N. J. May, '75 4 15. Cleveland, Ohio Oct., '75 4 16. Washington, D. C. Feb., '76 4 17. Philadelphia, Pa. June, '76 5 18. Philadelphia, Pa. Oct., '76 5

Jan 1, 1932

Trans No. Place Date Vol. 1. Wilkes-Barre, Pa. May, '71 1 2. Bethlehem, Pa. Aug., '71 1 3. Troy, N. Nov., '71 1 4. Philadelphia, Pa. Feb., '72.. 1 5. New York, N. Y. May, '72 1 6. Pittsburgh, Pa. Oct., '72 1 7. Boston, Mass. Feb., '73 1 8. Philadelphia, Pa. May, '73 2 9. Easton, Pa. Oct., '73 2 10. New York N. Y. Feb., '74 2 11. St. Louis, Mo. May, '74 3 12. Hazleton, Pa. Oct., '74.. 3 13. New Haven, Conn. Feb., '75 3 14. Dover, N. J. May, '75 4 15. Cleveland, Ohio Oct., '75 4 16. Washington, D. C. Feb., '76 4 17. Philadelphia, Pa. June, '76 5 18. Philadelphia, Pa. Oct., '76 5

Jan 1, 1934