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By Robert H. Merrill

In 1960, the Kennecott Copper Corp. and the U.S. Bureau of Mines entered into a joint research program to determine the changes in stress, strain, and displacement created by changes in slope angles. The field site was the Kimbley Pit, located near Ely, Nev. The supporting laboratory investigations were made at the Bureau's Denver Mining Research Center. Following a study to determine the rock stresses in the mining district, the strike and dip of geologic defects (fractures) in the rock, and the laboratory-determined strength of the rock, a new slope angle and pit configuration was selected, and instruments were installed to measure changes in stress, strain, displacement, microseismic noise rate, and sonic velocities at several points in the west wall of the pit. Between February and November 1966, the west slope was changed from an average angle of 45' to a slope angle that started at 45° on the south wall, and gradually increased to 575' at the center of the west wall, and to 61° on the north wall. The pit was deepened from about 500 ft to about 550 ft. About 2.55 million tons of waste and ore were removed by stripping and mining. Measurements made between November 1965 and September 1967 showed that the pit wall was stable before excavation began, that changes in the measured quantities occurred in the pit wall during the excavation, and that the wall returned to a condition of equilibrium after the mining was completed. The need for a method and the technology to pre-design a pit slope does not require elaboration. About 5 billion tons of material are excavated yearly by surface mining methods in the United States, and the economics, safety, and availability of lower grade metals, nonmetals and fuels could all be substantially improved if operators could preselect optimum slope angles. Mining companies and the Bureau of Mines have recognized the need for developing engineering methods to determine stable slope angles. In the industry, Kennecott Copper Corp. had been making an

Jan 1, 1969

By S. Andrew Kulin, L. Kaufman

A general treatment of the broadening of X-ray double-crystal spectrometer rocking curves due to distributions of tilts and strains has been carried out. The results of this treatment have been applied to the case where dislocations in a crystal give rise to both tilts and strains. On this basis it is found that the rocking-curve halfwidth due to imperfections in the crystal, ßC , increases with increasing Bragg angle, 8. These considerations predict linear relationships between & and tan2 , and ßCand tan 8, for Gaussian and Cauchy distributions of dislocations, respectively. The zero intercept of the ßC us 8 curve, which is a measure of the degree of imperfection of the crystal, is found to depend directly on the distribution rather than the number of dislocations. In addition, the slope-intercept ratio for the ßc - ? curve has been calculated for both the Gaussian and Cauchy distributions. Experimental determination of the - 8 curves for a series of copper single crystals prepared by the Bridgeman and Czochralski methods yields results in agreement with the calculated slope-intercept ratio for a Cauchy distribution of dislocations. CERTAIN X-ray techniques have been used to study the perfection of single crystals. Gay, Hirsch, and Kelly have considered the spreading of spots into arcs on Debye-Sherrer rings due to imperfections. The Double-Crystal Spectrometer has also been applied to this problem.2 The characteristics of Double-Crystal Spectrometer Rocking Curves have been described in the literature by various workers.2-4 By assuming a random spacing of dislocations which give rise to a Gaussian distribution of misorientations relative to some mean orientation in a crystal Gay, et al.,1 arrive at the conclu- sion that the most probable misorientation is given by where h is the spacing between dislocations, and 8 and are the relative misorientations of neighboring regions. If ßC is the halfwidth of the rocking curve (the width at one-half the peak intensity) due to dislocations in the crystal then5

Jan 1, 1960

This discussion was prepared by members of the Phelps Dodge Corpn. staff who were designated to review the minutes. It is confidential but permission has been granted to distribute copies among those who attended the Conference. The minutes of the above conference contain numerous references to the plants and practices of Phelps Dodge Corporation. As many of these references are inaccurate, the following notes are made covering such points: Construction Magnesite Brick in Reverb. Arch (1) An.18" arch, 20 feet long, was built on one furnace with 5 feet on each side of magnesite brick, the center portion .being of silica brick. The furnace was put in operation February 13, 1930 and on April 23rd after 68.8 days' operation it was replaced due to failure of the silica brick forming the center of the arch. At this time, while the silica brick had completely gone, there was 14" of magnesite left. Bridgewall Tapping (2) Tapping matte through the burner wall was first tried at the Copper Queen Smelter in April, 1925. At that time, a copper casting with 1 ½ extra heavy pipes for water cooling cast in it, and having one taphole block was used. Later, on the other (old) furnaces, similar castings, but equipped with two tapholes, were used. We found with these furnaces that we could tap through the burner wall at the beginning of a campaign, but would subsequently have to go back to side wall tapping. This was the case until sometime in 1926. The new furnaces, that went in operation at the beginning of 1927, each had water. cooled castings with two tapholes. These have been used entirely and have been quite successful. Sometime after the furnaces were put into operation sidewall tapholes were put in, simply for the removal of any accumulation of matte that might be temporarily retained at the front end by bridges or dams in the furnace formed by overcharging locally.

Jan 1, 1930

By R. V. Norris

It was stated in a former paper1 that an intensive study of anthracite costs was being made by the engineers of the United States Fuel Administration. The results of this study are now available and are offered as a supplement to the former paper. Anthracite costs as reported for the 6-months period, December, 1917, to May, 1918, inclusive, as compiled by the Federal Trade Commission, were used as a base, and charts similar to those described in the former paper were made for the standard white ash anthracite and for red ash and Lykens valley coals. Adjustments The adjustments of cost from a reported to a price-fixing basis, as described for the bituminous methods, were applied but showed only minor adjustments as necessary. The great spread in anthracite prices on the varyingsizes, which for the 6-mo. period under review ranged in average from $5.244 for nut to $2.074 for barley coal, makes the question of the percentage of sizes produced at the different collieries a vital one. The realization with the same prices for each size must be within very wide limits, when it is considered that the percentage of prepared coal reported from different collieries varied from over 80 per cent. to below 30 per cent. for fresh-mined coal. Hence, as the spread in prices for the various sizes must be predicated on some percentage, it is essential to find some method of adjustment to allow for this variation. The logical method of adjustment is to calculate actual costs to costs as of the standard percentage of sizes, so that the margin between the adjusted costs and the average realization shall be the actual margin for each colliery between its actual costs and actual realization due to its particular percentage of sizes. As a base for realization the actual percentage of sizes for fresh-mined coal for the 6-mo. period was adopted. This percentage is given in Table 1.

Jan 1, 1920

By Charles H. White

Discussion of the paper of A. L. Sweetser, Tress., xxxviii., 236. CHARLES H. WHITE, Harvard University, Cambridge, Mass. (communication to the Secretary*) :-Those interested in leaching-processes gladly welcome contributions on either laboratory-tests, such as are reported by Mr. Sweetser, or mill-tests made on larger quantities of ore; and especially valuable are such reports, even with negative results, if the methods adopted are the outgrowth of an intimate knowledge of the conditions to be met. It seems very desirable, therefore, that as complete details as possible be given in reporting on such tests, particularly with respect to the character of the ore under examination. The questions of prime importance in the sampling of ore and in the making up of charges for laboratory-tests are those respecting its grade and uniformity. In Mr. Sweetser's paper the first of these questions is answered definitely, but the answer to the second, although apparent on inspection of the results, might be more strongly emphasized with advantage, since it is only on account of the extreme uniformity of the ore that the satisfactory results obtained could have been expected. It seems advisable to call attention to the fact that for average ores, and more particularly for uneven, or 11 spotted ores, the quantities used by Mr. Sweetser are not usually regarded as sufficient to insure satisfactory results. Let us examine the method somewhat in detail. The ore was crushed to pass a 5-mesh screen, and a sample of 45 lb. was taken out for assay and tests. The diameter of the largest grain of this ore, according to Richards,1 was 2.67 mm.-assuming here, as in the following cases, that the wire of the screen was of the largest size used for the purpose, thereby giving any possible advantage in this respect in favor of the method. The table in Richards's book, p. 852, shows that 45 lb. is a sufficient quantity for a fair sample with uniform ores, but not sufficient

Jan 7, 1908

By H. A. Baumann, A. J. Rostosky

EVER since the first installation of wet-washing methods of coal preparation, the removal of the water added by the washing process has created serious technical and operating problems. The rapid development and widespread adoption of wet washing has necessitated a corresponding development of dewatering methods and equipment. The amount of water retained by coal varies with the size of the coal. The moisture that is largely on the surface of the particles is easily shaken off coarse coal; thus coarse coal can be dried comparatively easily and quickly. Fine coal presents a larger surface area than does coarse coal, so its capacity for retaining moisture is greater. In fine coal, moisture below about 5 per cent is held between the coal particles and its removal is difficult. Draining and drying of the finer sizes is a major problem at many washeries. Two common problems of coal-washing practice that are intimately connected with dewatering and drying are the clarification of washery water and the prevention of stream pollution. The first field has been studied from many angles, including the use of coagulating and settling agents followed by vacuum filters to remove the solids, and the use of sludge tanks with dewatering screens incorporated therein. In many districts, the disposal of sludges into streams has been stopped by law. The lack of space for impounding sludges has created problems of disposal in other districts. Reduction in moisture means a saving in freight costs as well as the elimination of difficulties in unloading and storing. Excess moisture in coal causes it to freeze in localities where severe climatic conditions prevail. Excess water in coals to be used in the production of coke and manufactured gas has a chilling effect on the walls of the oven or retort. Dryness of coal is important in producing good briquettes. CLASSIFICATION OF MOISTURE Moisture in coal can be classified as inherent and surface moisture. Inherent moisture is regarded as a part of the true coal substance, having been present in the original vegetal matter from which coal was

Jan 1, 1943

By George A. Colligan, John M. Lo

Electrical resistivity as a function of temperature was investigated in undercooled gallium and under-cooled Ga-In alloy of eutectic compositian to examine structural changes in undercooled melts. A continuous linear extension of the resistivity was determined for 99.999 pct pure gallium when undercooled below the stable Ga(l)-liquid, and the metastable Gafl)-liquid equilibria. In icndercooled Ga-In, a significant continuous negative departure from linear resistivity behavior was observed below the eutectic temperature. All resistivity measurements were made by the four-point probe technique. ThE temperature range for investigations into the physical properties of liquids before freezing has been greatly extended with the undercooling of massive melts of pure metals and alloys. Experimental determinations of thermal and transport properties and diffraction patterns in undercooled systems should throw further light on the structural behavior of liquids. Many measurements showing prefreezing anomalies in viscosity1'5 have been regarded677 as evidence for structural groupings in the liquid state. However, the majority of viscosity results,~)~-~~ conforming to the law 17 =A exp(~/RT), do not support theories of atomic aggregation. For the system Cd-Sb, which exhibits a metastable solid phase, Fisher and Phillipsg reported that, in undercooled alloys, the activation energies for viscous flow below the stable liquidus are much higher than those above it. Similarly, Miller et a1." observed a doubling of the activation energies for electrical conductivity in undercooled Cd-ib alloys. These effects were suggested as indicative of atomic association and clustering in the liquids at lower temperatures. With the recent establishment of metastable phases in pure gallium22 and its alloys,23 attempts were made during this investigation to undercool bulk samples and to determine the temperature dependence of electrical resistivitv. Pure gallium, when undercooled below -17°C at atmospheric pressure, can crystallize to either a stable Ga(1) phase (mp, 29.78"C) or a metastable phase identi-

Jan 1, 1967

By Robert Linton

Tats paper deals with work that has been carried on for over three years by the management and staff of the North Butte Mining Co. in an effort to standardize mining methods, to eliminate lost motion in the operating organization, and to improve the efficiency of the individual workman. Naturally, much of this work is an adaptation of ideas suggested by other mining operations or by practice in manufacturing and other productive industries. The studies, however, have developed some features that it is believed are new. Both new and adapted features are given in detail. The normal capacity of the North Butte mines is about 2000 tons per. day. Hoisting is done principally through the Granite Mountain shaft, which is 3740 ft. (1140 m.) deep and equipped with an electric hoist capable of hoisting 200 tons per hour from a depth of 4000 ft. (1219 m.). The lowest stoping level is the 3400-ft. The Granite Mountain shaft is connected with the Speculator shaft, about 850 ft. (259 m.) south, and with the Gem, about 1150 ft. (350 m.) east, which are 3000 (914.4 m.) and 2000 ft. (609.6 m.). deep, respectively. These two shafts are equipped with reversible fans for ventilation. The Speculator has a; steam hoist with a capacity of 150 tons per hour from the 3000-ft. level. The Granite Mountain shaft is concrete lined from the 1600-ft. to below the 3000-ft. level and gunited with cement from the 1600-ft. level to the surface. Sections of the Speculator shaft with a total combined length of 500 ft. (152.4 m.) are also concreted. There are approximately 36 miles (58 km.) of drifts and crosscuts in the mine, in which there are 15.2 miles (24.5 km.) of track. At normal capacity, about 75 stoping faces, 20 raises, and 30 sill development faces are working. Stoping is usually by the standard square-set method. Where the size of the oreshoot and character of. the walls permit, the rill (or cut-and-fill) system is employed.

Jan 8, 1920

By E. A. Trager, A. W. McCoy, H. R. Shidel

Sampling of the fluid from oil wells for percentages of oil, emulsified oil, and water during the last two years has brought out some interesting facts concerning oil-water emulsion. This result led to a laboratory investigation of emulsion, which substantiated the conclusions made from the field observations. The purpose of this paper is to present the information collected, the laboratory experiments, and our interpretation of the same. In order to define emulsified .oil exactly, give its synthesis and origin, and to show how and when it is formed in the wells, the work was necessarily divided into two separate lines—laboratory work and field observations. It is hoped that this study may lead to a discussion of such points so that the petroleum engineer, geologist, or technologist may be benefited by its practical bearing on oil-field management. Special credit is due Mr. Everett Carpenter, chief geologist of the Empire Gas & Fuel Co., for his assistance and cooperation in this work. Labobatory Investigations on Emulsified Oil Laboratory investigations were conducted in an attempt to learn the composition and some of the properties of emulsified oil, or B. S., as it is more commonly called, also to demonstrate, by laboratory methods, how B. 5. may be formed under conditions similar to those existing at the time a well is being pumped, and how it may be broken down. Literature bearing on this subject is widely scattered and very limited in scope. Bacon and Hamor define B. S., or bottom settlings, as "earthy matter, inert organic matter, or, in the case of Pennsylvanian petroleum, an emulsion of paraffin wax and water, which accompanies crude oil." In this discussion we will limit the term B. S. to that heavy, dark-brown emulsion, composed of a physical mixture of water, oil, and air with some included inert matter, either organic or inorganic. Possibly the first step in a description of this product should be a description of its physical properties, but since most operators are quite

Jan 1, 1921

By C. R. Austin

THE investigation described in this paper deals with the development of high-temperature alloys of the Konel series over a considerable period of time at the Research Laboratories of the Westinghouse Electric and Manufacturing Co. The word "Konel" has been coined to designate a class of alloys which are essentially nickel-cobalt-iron with other elements added, usually in small quantities. Most of these alloys contain titanium, usually about 2.5 percent. The main consideration related to strength at high temperatures but the thermionic characteristics of the alloys was also of considerable importance. It is, however, with the mechanical properties at elevated temperatures that we are herein concerned. Much attention has also been devoted to their age-hardening characteristics. There is little need to stress the utilitarian aspect of this subject, as it has been discussed frequently. Alloys exhibiting superior properties at normal or elevated temperatures rapidly find many suitable applications. There may be some question, however, as to what constitutes a "superior alloy, " or an alloy that may have good service possibilities at high temperatures. The problem is complex and various means have been employed to predict the useful life that may be expected of a material. In recent years the "creep limit" has been used as the best means of evaluation. Such tests necessarily are prolonged and hence collection of data is very slow. In a general survey of many alloys therefore the creep test is not desirable and recourse must be had to much shorter time tests in order more rapidly to select alloys which appear to show promise. For this reason the present survey has been made using the high-temperature tensile test and a bend' test developed by Howard Scott.1 In the tensile test the conditions have been standardized for all tests, so that values obtained are at least comparable, one alloy with another, although the absolute values recorded merely indicate the mechanical properties at the given temperature under a definite rate of loading.

Jan 1, 1931

By P. D. P. Hamilton

In any mining district where benefication of the ores is practiced, one can usually observe several types of flow sheets even though the various producers are milling ore from the same geological formations. It is obviously impossible to describe all the mills in the Porcupine district. However, the following description of some of them is of particular interest in that it shows where one mill has paid particular attention to corduroy blanket practice; one to gravity concentration of sulfides for intensive cyanidation of the values therein contained; one to flotation of values and cyanidation of this product only; and others where small variations of standard all-slime cyanidation practice prevail. In all cases it will be seen that cyanidation is the basic principle. Discussions involving the difference in the flow sheets and the reasons for them will always be of value to those interested in the art. The author of this article, however, believes that such discussions should originate from the operators of the mills under discussion. For this reason no comparisons or discussion will be included in the description of the mills mentioned. Should any reader of this article obtain information therefrom which will be of value to him in helping to solve his own particular problems, the author will consider that the article has satisfactorily served its purpose. All values mentioned throughout this article are based on the value of gold at $20.67 per ounce. Coniaurum Mines, Limited The ore of the Coniaurium mines consists of a mineralized schist and quartz containing about 2.5 per cent sulfides. Owing to the low grade of ore developed in the mine, extreme care has had to be exercised throughout the entire mining and milling operation in order that an operating profit might be obtained. The management, however, has been able to keep costs down, with the result that at present the outlook for the property is considerably brighter than it has been in the past. This is partly due to the fact that it is now known how to mine and treat

Jan 1, 1935

By Frederick Laist

While remodeling No. 1 section of the concentrator at the Washoe Reduction Works of the Anaconda Copper Mining Co. during the summer of 1912, for the purpose of ascertaining what additional metal recoveries could be made by improvements in concentration practice which had been developed by the staff of the Boston & Montana Reduction Department, at Great Falls, it was deemed advisable to carry on some experiments on the treatment of the regular mill tailings by roasting and leaching, to see whether the losses of metal values might not be still further decreased. In the course of these experiments about 5,000 tons of tailings were roasted, and of the resulting calcine about 200 tons were leached. The results were so satisfactory that it was decided to continue the work on a larger scale, particularly as regards the leaching end of the process, and construction work on an 80-ton roasting and leaching plant was accordingly commenced during the latter part of February, 1913, and is being pushed as rapidly as possible. 80-Ton Leaching Plant. This plant will consist of one 20-ft. roasting furnace of the Mac-Dougall type, specially adapted to the work of roasting for leaching; two leaching tanks, 32 ft. in diameter by 12 ft. deep, together with the necessary solution tanks, air lifts, precipitating launders, etc. It has been decided to precipitate the copper from solution on scrap iron, so as not to hamper the roasting and leaching departments by any weakness that might develop in the precipitating department. It is, however, the intention to experiment with a number of different methods of precipitation; for example, by electrolysis and by hydrogen sulphide. The acid required for the operation of this plant will be purchased, ad the quantity required is not sufficient to justify the installation of an acid plant. An acid plant will, however, be established in con-

Jan 1, 1914

By Richard S. McCaffery

A Research on iron blast-furnace slags has been in progress since 1922 in the metallurgical department of the University of Wisconsin, supplemented by observations, tests and critical examination of operating records at 10 furnaces. From the data thus collected information has been put into forms suitable for use by furnace operators and is given in this report and the three accompanying papers.' In the viscosity results reported, it is estimated that temperatures were measured within 5" C. and that the viscosity determination of the particular slag investigated is accurate within ± 4 per cent. In the viscosity diagrams in the various papers, the equal viscosity lines were drawn in such a manner that in all cases they are located correctly with respect to the actual experimental results, but when determining a viscosity for a particular composition from the diagrams, the viscosity may. be slightly less accurate than the experimentally determined viscosity for the same slag, for the reason that the iso-vis lines interpolated between the experimental results may be shifted slightly without violation of the experimental data. The viscosity diagrams represent our judgment of their position under the above conditions, and we believe the relative accuracy of viscosity data taken from the diagrams is more than sufficient for any use to which the data may be put. Hundreds of samples of slags synthetically produced in the laboratory and of actual furnace slags have been studied in various ways. The ranges of compositions of slags used for viscosity measurements are: silica 25 to 55 per cent.; alumina 0 to 35 per cent.; lime 25 to 55 per cent.; magnesia 0 to 20 per cent. These measurements were made at temperatures of 1250°, 1300°, 1350°, 1400°, 1450°, 1500°, 1550°, 1600" and 1650" Centigrade. Descriptions of apparatus, methods used and statements of the problems investigated have been published in the papers listed in the bibliography beginning on page 120.

Jan 1, 1932

By W. Spencer Hutchinson

The experiments described in this paper were undertaken primarily for the purpose of measuring the quality of work done in screening and sorting in American concentrating-mills for Prof. Richards' work on Ore-Dressiag. For this purpose a series of samples of screened and sorted products was obtained from four different mills, and a plan was devised for sizing these products, and for tabular and graphic representation of the results. The results, together with discussions as to their significance, so far as they serve to interpret the quality of mill-work, have already been published in the above-mentioned work, and it is not the purpose of this paper to duplicate the work, except so far as is necessary to explain and illustrate methods and appliances. The sizing of coarse products was clone by sifting on screens of sheet-metal with round punched holes. These holes ranged in size from 64 mm. to 0.5 mm. in diameter. Below the smaller limit, punched metal, suitable for this work, was not obtainable, and the sizing was carried on by sifting on woven wire-sieves. The finest woven wire-cloth obtainable gave a hole 0.073 mm. square, and the finest silk bolting-cloth pave a pentagonal hole with an average diameter of 0.069 mm. Even these very fine screens served no useful purpose in sifting certain of the very fine classifier-products, of which all, or nearly all, passed through the finest screen. A new adaptation of a method of settling and decantation with water in beakers was therefore worked out, and by this means the sizing mas carried down to grains of a diameter as small as 0.012 mm. Description of Apparatus and Manipu1ntion.—Sieves of sheet-metal with round punched holes were used as far as practicable, chiefly because, as already explained, the primary object of the experiments was to test mill-work, and the mills had all used

Jan 1, 1905

By R. J. Evans

The Combined Metals Reduction Co.'s plant is at Bauer, Utah. It was built primarily to treat ore from the Combined Metals mine at Pioche, Nevada. Shortly after its completion, the company acquired the old Honerine mine, and ores from this mine, as well as custom ores from widely scattered districts, are now being treated at the Bauer plant. Pioche Ore The Pioche ore, which comprises about 50 per cent. of the mill feed, occurs as a replacement of an impure thin-bedded limestone. The lime beddings are separated by thin layers of micaceous shale containing appreciable quantities of sericitc and graphite. The ore is an exeeptionally fine-grained, dense mixture of minerals and gangue, assaying approximately 0.03 oz. gold, 7.2 oz. silver, 7 per cent. lead, 15 per cent. zinc, 16 per cent. iron, 29 per cent. insoluble and 1.5 per cent,. lime. Of the total lead 7 per cent. is in the form of sulfate and carbonate. Approximately 10 per cent,. of the ore consists of chert nodules impregnated with barren pyrite; the nodules average 0.8 mm. in size, and ire extrcmely hard, compared with the other constituents of the ore. Microscopic examination of Pioche ore ground to 80 per cent . minus 150 nicsh shows pyrite, sphalerite and galena still attached. Early Attempts at Ore Dressing Metallurgists, both professional and those working for private companies, have been more or less familiar with the Pioche ore sinee. 1913; in fact, a total of 35 commercial and plant laboratories expenmented with the ore between 1913 and 1921. The work during this period was decidedly negative, it being seemingly impossible to effeet a satisfactory separation of the valuable minerals into marketable products. In 1913, it was recognized that fine grinding-—that is, about minus. : 200 mesh—-was required to liberate the minerals from one another, and from the gangue, but such fine grinding was not considertd practicable at this time.

Jan 1, 1928

By R. H. Richman

Coarse-grained Fe-Ni-C martensites formed at subzero temperatures were strained in compression at room temperature and the plastic deformation modes examined as a funclion of carbon content. At very low carbon levels, defovmation occurs hy wavy slip; commencing at about 0.05 pct C, mechanical twinning accounts for an increasing proportion of the plastic flow. Beyond approximately 0.4 pct C the deformation appears to be entirely by mechanical twinning. The habit planes of the operative twin modes were determined to be (113), (0131, and "(089)" in decreasing frequency of occurrence. Some of the implicaticms of Ihe o6served deformation modes upon the structure and properties of ferrous martensites are discussed. METALLURGISTS have long sought to explain the mechanical properties of ferrous martensites, particularly the extraordinary hardness and strength of the high-carbon varieties. Until recently the lack of precise knowledge of the structure of martensite crystals was the most serious barrier to understanding; however, the advent of thin-foil techniques for transmission electron microscopy has removed much of the mystery from martensite structure. Observations by Kelly and Nutting,"' and many others since, indicate that Fe-C based martensite crystals are internally twinned to a degree dependent upon the carbon content, and that a martensite plate containing 0.8 wt pct C or more consists of a stack of twin-related lamellae with each lamella on the order of l00A thick. Such a structure would be expected to influence greatly all aspects of metallurgical behavior, and many of the properties of Fe-C based martensites have been rationalized accordingly. Recent independent experiments led Winchell and cohen3 to conclude that the strength of ferrous martensite is derived primarily from solid solution hardening of the iron lattice by interstitial carbon atoms. Strength contributions from precipitation, dislocation-atmosphere interactions, clustering, short-range order, crystallite size, and so forth, were considered to be of secondary importance. In the view of Winchell and Cohen, the twinned structure of martensite serves mainly to limit the effective length of dislocation lines in the lattice, and thereby confers an added measure of hardening. Few properties of metallic crystals are as sensitive to metallurgical structure as are deformation mechanisms. It is somewhat surprising therefore, that, except for metallographic evidence of mechanical twins in strained martensite,4 no systematic examination of plastic deformation

Jan 1, 1963

By Wilbur Stout

IN Ohio, the annual output of coke made from native coals has averaged not more than 70,000 tons, or about enough to run a 200-ton blast furnace. Raw coal locally mined from the Sharon, or No. 1, bed was used for iron smelting in Mahoning furnace at Lowellville, Mahoning Co., as early as 1846. This practice was continued in the Youngstown district until the coal became scarce and the Connellsville coke more popular. At present, the Sharon bed contributes a part of the fuel for the production of ferrosilicon iron at Jackson, Jackson Co. About one-third native coal and two-thirds soft coal are used. In this state, coke, on a commercial basis, was first made in Jefferson County probably about 1861 or 1862. The coals used were the Steubenville Shaft, or Lower Freeport, and the Strip Vein or Middle Kittanning. The practice, however, was discontinued before 1880 because of the competition of superior foreign cokes. Beehive ovens, later built at Leetonia, Columbiana Co., were operated for only a few years; the coal used at this place is the Lower Kittanning or No. 5. In a desultory way, slack from the Pittsburgh coal has been coked at a few places in eastern Ohio, and attempts, either for experimental or for commercial purposes, have been made in other parts of the State. Small quantities of coke were made from the Upper. Freeport coal at Salineville, Columbiana Co., and near Nelsonville, Hocking Co.; the results were poor. Tests were made of the Middle Kittanning coal near Zanesville, Muskingum Co., at Moxahala Furnace, Perry Co., and at Washington Furnace, Lawrence Co.; the cokes were all high in sulfur and ash. Attempts were made to coke the Clarion coal at Vinton Furnace, Vinton Co.; a small battery of molding ovens and a pile of high-sulfur coke yet remain to tell the story. The Pittsburgh coal was coked on a commercial scale at Utley and at Lathrop, Athens Co.; but the operations were not successful and were soon abandoned.

Jan 9, 1919

By R. L. Klueh, W. W. Mullins

During the course of a study on the periodic precipitation (Liesegang phenomenon) of water vapor bubbles in solid silver, some interesting observations were made concerning the formation and growth of these bubbles. Under certain conditions it was found that the bubbles which formed in the single crystals led to the formation of large blisters ad cracks, and when poly crystalline specimens were used grain boundary cracking was observed. These two phenomena were examined in some detail, and it was found that an important step in the formation of the blisters and grain boundary cracks was the impingement of neighboring bubbles. DURING the study of Liesegang precipitation in solid silver,' some interesting observations were made concerning the formation and growth of water vapor bubbles. The bubbles were formed by annealing oxygen-saturated silver in hydrogen. Because of the similarity of this phenomenon to the hydrogen embrittlement of tough pitch copper,' it is known in the literature" as the hydrogen embrittlement of silver. In some respects it appears that the hydrogen embrittlement of silver is similar to the processes of hydrogen attack of steel5 and blister formation.' Hydrogen attack results from the high-temperature reaction of hydrogen with dissolved carbon or carbides in the steel to form methane (usually at grain boundaries). Blister formation occurs in steel and other materials as a result of hydrogen produced by corrosion, pickling, and so forth. Because of these similarities, some speculations were made concerning these other processes. EXPERIMENTAL PROCEDURE Water vapor bubbles were formed in 99.999+, 99.99+, and 99.9 pct Ag by annealing in a partial pressure of oxygen (pq) until saturated and then in a partial pressure of hydrogen (pH,). The hydrogen diffuses into the silver and reacts with the dissolved oxygen to form water vapor which precipitates in the form of bubbles. The specimens were then sectioned, silver-plated, mounted in cold-setting plastic, ground and polished through a preliminary polishing, and finally chemically polished.' The bubbles were visible without etching.

Jan 1, 1969

By Richard A. Arterburn

For many years hydrocyclones, commonly referred to as cyclones, have been extensively utilized in the classification of particles in comminution circuits. The practical range of classification for cyclones is 40 microns to 400 microns, with some remote applications as fine as 5 microns or as coarse as 1000 microns. Cyclones are used in both primary and secondary grinding circuits as well as regrind circuits. The information given in this chapter is intended to provide a method, at least for estimating purposes, of selecting the proper number and size of cyclones and to determine the proper level of operating variables. Generally, it is recommended that cyclone suppliers be consulted for sizing confirmation. Sow cyclone suppliers employ digital computers to aid in the sizing and selection of cyclones. CYCLONE DESCRIPTION AND BASIC OPERATION Figure 1 shows a cutaway view of a typical cyclone. During operation, the feed slurry enters the cyclone under pressure through the feed pipe into the top of the cylindrical feed chamber. This tangential entrance is accomplished by two types of design, as shown in Figure 2. Since the majority of research has been done with the involuted type, the graphs and relationships shown may not be strictly applicable to other designs. As the feed enters the chamber, a rotation of the slurry inside of the cyclone begins, causing centrifugal forces to accelerate the movement of the particles towards the outer wall. The particles migrate downward in a spiral pattern through the cylindrical section and into the conical section. At this point the smaller mass particles migrate toward the center and spiral upward and out through the vortex finder, discharging through the overflow pipe. This product, which contains the finer particles and the majority of the water, is termed the overflow and should be discharged at or near atmospheric pressure. The higher mss particles remain in a downward spiral path along the walls of the conical section and gradually exit through the apex orifice. This product is termed the underflow and also should be discharged at or near atmospheric pressure.

Jan 1, 1982

By William E. Brewster

Dating back some five years ago, various foundries made inquiries as to the probable total carbon content in a given specification and grade of pig iron. Up to that time we had no data, and except for a few occasional determinations of total carbon, no attempt had ever been made to set up routine determinations of total carbon. In fact, at that time, a cast analysis was considered sufficient, without examining at further length into possible variations of analysis between ladles of the same cast. In view of the various inquiries as to total carbon, we sampled all piles of all grades of pig iron, making complete analyses. Finding some variation in carbon content, we secured more samples and plotted the carbon content with respect to silicon in each grade of iron except for the higher phosphorus foundry iron, where it was necessary to divide the group into twenty-point phosphorus ranges. This is easily understood, owing to the falling off of carbon content with the increase in phosphorus content, silicon in each case being the same. In addition, these curves clearly demonstrate the lowering of the carbon content as the silicon is increased. In order to insure accuracy and reliable determinations the following method of sampling and analysis was employed: Two pigs selected from each end of the car, each car containing iron from one 50-ton ladle, are drilled with a 5/6-in. drill all the way through from top to bottom in the center of one section of the pig. The drillings are collected and transferred to an 80-mesh sieve; the portion passing through and the portion retained are placed in separate envelopes and labeled accordingly. The amount of sample in each envelope is weighed to obtain the proportionate amount of coarse and fine drillings, then a 1/2 factor weight of sample (1.3636 grams) made up of proportionate

Jan 1, 1936