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By A. V. Deforest, A. R. Anderson, C. W. MacGregor

One of the important problems in the design of structures and machine parts subjected to rapidly applied loads is the determination of the strength and ductility of the material itself under such conditions. Realizing the need for such information, many investigations have been carried out in the past. Most of these experiments have been made using notched bars in bending either with the Charpy or Izod impact machine. While yielding valuable data on the effect of notches in the energy absorption of the material and on the condition of heat-treatment, this form of test does not lend itself readily to a determination of fundamental stress-strain relations. For this and other reasons, more attention has been paid recently to the tension impact test. An extensive literature already exists on the subject and only those investigations somewhat closely related to the present paper will be mentioned. For a more comprehensive list, the reader is referred to the Symposium on Impact Testing or the A.S.T.M., 1938. The work of Mann.' Clark and Dätwyler,2 Nádai and Manjoine,3 Winlock and Leiter,4 Deutler,5 and Brinkmann,6 however, deserves especial mention. In most of the previous investigations on the subject, when stress-strain relations were obtained under high rates of loading, it was the common procedure to determine stresses based on the original area and strains based on the original length. Stress-strain curves were then constructed by P ALo plotting So = 1/A0 as a function of eo = ALr,o where A0 and Lo refer to the original area and gauge length, respectively. At present there exists a paucity of true stress-strain data in the impact problem. The advantages of constructing so-called true stress-strain curves in order to reveal the fundamental physical properties of the material for the tension test were discussed recently by one of the authors.7'8 It was shown that the true physical behavior of the material in the tension test could best be represented P by plotting the average true stress, S = 2P as a function of 6 = q' = log —r = log ~9L where P, A, L, S, E, and q' refer to the load, the cross-sectioned area while the load P is applied, the actual gauge length, the average true stress, the true strain and true reduction in area, respectively. Besides giving a better physical representation, the method has certain additional advantages in that the stress-strain curve so plotted becomes a straight line from the maximin load point to fracture. The construction of such curves, however, while comparatively simple for the slow rates of loading customarily used in the tension test, presents certain dificulties for tensile impact in that both loads and diameters are required throughout the test. The continuous measurement of test-piece diameters to fracture under impact conditions offers considerable experimental difficulty. In order to make it feasible to

Jan 1, 1942

By B. M. Larsen

AN increasing amount of attention is being paid to the possible influence of oxygen, in its several modes of occurrence in steel, upon some of the properties of the metal; but clearly investigations along this line are unlikely to yield significant results until reliable analytical methods for the determination of oxygen have been established. This paper presents some results of a study of an improved Ledebur method which aims to determine a definite portion of the oxygen; namely; that which presumably is present in diffusible form in the. solid steel. These more or less preliminary results lead us to believe that the oxygen so determined has some significance; their presentation at this time will, we hope, stimulate discussion of, and focus attention upon, certain rather neglected aspects of the problem. Considerable work has been done recently on methods of estimating inclusions and on the vacuum-fusion method for oxygen; with results which, however, are of somewhat questionable significance. It is commonly presumed that the oxygen in a bath of molten steel in a furnace is present partly as a separate phase consisting of suspended particles of oxides or silicates, partly in solution, the latter being probably the more important. Upon addition of a deoxidizer, it reacts with some portion of the dissolved oxygen and precipitates it as additional suspended particles; the oxygen not eliminated in this way, which remains in solution in the liquid steel, may be regarded indifferently either as dissolved oxygen or as dissolved FeO. When the steel is poured, some of the suspended particles rise towards the surface, others remain trapped in the finished steel, appearing as nonmetallic inclusions in amounts corresponding usually to about 0.01 to 0.03 per cent 02 in the steel. The dissolved oxygen may be partly eliminated as CO during freezing of the ingot, but some remains either in solution or in part precipitated as FeO. Finished steels may thus contain oxygen in several forms: (1) In non-metallic inclusions which are probably quite insoluble in the solid metal at all temperatures; (2) in solid solution which at lower temperatures may be supersaturated; (3) in FeO particles produced by relief of this

Jan 1, 1932

By D. J. McAdam

RESULTS of investigation of corrosion-fatigue of metals at the U. S. Naval Engineering Experiment Station, Annapolis, Md., have been discussed by the writer in several recent papers. 1.2.3.1 In those papers references were also given to the work of Haigh in 1917.6 Efforts by the writer to obtain a better understanding of some of the factors involved in corrosion-fatigue finally developed into a general investigation of the effect of time and cyclic stress on corrosion of metals. Corrosion fatigue of Metals As described in the papers mentioned, rotating-cantilever specimens, conically tapered so that the stress is nearly uniform over a length of about 1.5 in., were subjected to simultaneous fatigue and corrosion. For this purpose a stream of water was diagonally applied so as to sweep the specimen from the outer to the inner fillet and surround the stressed surface with water. The fresh water is a calcium carbonate water, the composition of which has been given in a previous paper .2 The salt water is Severn River water having a saline content about one-third that of sea water. The previous papers described results obtained with ferrous and nonferrous metals. The experiments showed that even slight corrosion simultaneous with fatigue may cause failure at nominal stresses far

Jan 1, 1928

By Kefton H. Teague

Olivine is a mineral containing a mixture of forsterite (Mg2SiO4) and fayalite (Fe2SiO4) in solid solution. The name olivine was first applied by Werner in 1790 (Hunter, 1941) because of the olive-green color of the mineral. Olivine is the principal component of the rock dunite, which itself is a member of the peridotite group of ultrabasic igneous rocks. European uses of the terms olivine and dunite are somewhat different than US usage. In commerce European usage defines dunite as containing 36 to 42% MgO, 36 to 39% SiO2, and loss on ignition of approximately 10%. The term olivine is used to designate a material containing approximately 85% forsterite, with a chemical composition of approximately 45 to 50% MgO, 40 to 43% Si02, 5 to 8% Fe2O3, and ignition loss of 1 to 2%. Uses The principal use of olivine is in various applications involving hot metal. Table I shows the best estimate available as to the amount of olivine consumed by various industries worldwide. In excess of 90% of world uses involve hot metal, with approximately 75% as a slag conditioner in the blast furnace production of pig iron. Approximately 15% of total olivine usage is as a special foundry sand in mold making for the brass, aluminum, magnesium, and manganese steel foundries (Schaller, 1957, 1958; Snyder, 1957; Anon., 1977). Olivine was first used as an industrial mineral in the early 1930s for a refractory material (Anon., 1970; Hunter, 1941). As a refractory raw material, olivine was first introduced in the United States as hand-cobbed, selected, shaped blocks of crude olivine. This use met with limited success. More recently, finely ground olivine blended with MgO and pressed into bricks, which are then fired, has found use in glass tank furnaces and open-hearth furnaces. Ramming or gunning mixes for basic furnace linings also utilize olivine. Olivine has been used in ladle linings with varied success. In Europe, substantial tonnages of olivine are utilized in refractory brick for night storage heaters (Anon., 1970). A limited amount of olivine has been used in the past as a fertilizer (magnesium source) and has been fused with rock phosphate to produce a magnesium phosphate as a plant food. The relatively high magnesia (MgO) content of olivine also attracted attention to it as a potential source of both magnesium compounds and as a source of metallic magnesium (Bengston, 1956; Hunter, 1941). The use of olivine as an additive in the blast furnace is relatively new, having been developed during the last decade. J.W. Currier (Private communication, 1978) describes the behavior of olivine in the blast furnace as follows: Olivine (Mg, Fe2, Si04) represents a unique lime-free source of magnesia and silica as a blast furnace charge component. Only recently has the magnesia-rich mineral been recognized as beneficial cleaner for blast furnace experiencing the problems associated with excessive alkali inputs and/or low coke stabilities. For some time it has been known that alkalis tend to recycle and accumulate in blast furnaces. In the high temperature region of volatilization the elemental alkalis formed attack the integrity of the graphite structure. This alkali attachment is enhanced when poor coke is used. Higher in the furnace stack volatile alkali cyanides condense and oxidize. This accumulation of alkalis tends to agglomerate the burden materials, producing scabs and scaffolds. These structures not only inhibit the downward movement of burden materials and the upward passage of wind, but also lead to erratic furnace performance

Jan 1, 1983

By Colin C. Harris, Nathaniel Arbiter

Flotation kinetics is the study of the variation in amount of froth overflow product with flotation time, and the quantitative identification of all rate con- trolling variables. With such variables maintained constant, the algebraic relationship between the proportion of mineral floated and flotation time is a flotation rate equation. This contains the constant values of all the rate determining variables implicit in one or more rate constants which must be evaluated from experimental data. The form or forms of the rate equation can either be deduced from assumptions or facts established about the mechanism of the process, or, more commonly, determined empirically or by analogy with similar rate process. (See pages 215 to 226 and 221 to 227.) The earliest contributions on flotation kinetics came in the mid- ~hirties,3,27,31 with interest increasing in recent years. The broad goal of these studies is the reduction of the total variability of a flotation system to equations among the individual variables, or among a lesser number of dimensionless groups.43 Although the goal is at present far from realization, its attainment will yield many practical advantages. The quantitative under- standing of the process should lead to improvements both in metallurgical results and flotation cell design. It should lead also to reliable methods for scaling-up* of laboratory and pilot plant results to full size operation without the need for expensive and time-consuming mill tests on commercial machines. Finally, this type of quantitative information will assist the automation of flotation cells and circuits as well as control by computers. Replacing present methods of control based on post-mortem assays and trial and error manipulation of reagents and cell variables by control through exact process analysis should yield higher levels of grade and recovery through reduction of the large variability of present-day results. CHEMICAL ENGINEERING ANALOGIES: Industrial flotation always involves flow of pulp through a series of stirred tanks. It thus bears formal analogy to many chemical engineering processes, the kinetics of which in flow systems has been studied for many de- cades.38,39,73,83 It will be useful to review these studies before considering flotation kinetics. Batch and continuous treatment are classified in Fig. 1. Batch operation in the true sense is unknown in flotation where concentrate is always removed

Jan 1, 1962

By Ta M. Li

A change in priorities is perhaps the underlying motivation reflective of governments and people in the Far East nations of the Philippines, Indonesia, and Thailand. While in the past, local governments have focused on feeding their large populations, the transfer of modern technology to the agricultural sector has produced record crop years in many of these countries. As a result, agricultural objectives have become more manageable, prompting governments to shift priorities to the accelerated development of their industrial sectors. The basis of any sustained industrialization program, the governments concede, is the development of a vibrant minerals sector.

Jan 8, 1979

By Walter Knox, J. D. Doherty

In 1939, at the request of The Koppers Coal Co., the Koppers Company Research Department established a Stoker Coal Research Laboratory for the purpose of investigating the performance characteristics of stoker coals in typical domestic stoker installations. Fig. I is a photograph of part of the laboratory. The material in this paper deals with the performance of one coal in a single installation. Presented are details of this installation, mechanical characteristics of the stoker, and the results of a series of burning tests in which adjustments were systematically varied. Summary OF Results Mechanical Features I. A large percentage of the total power consumed by a domestic stoker is used to operate the gear case and motor (Fig. 3)- 2' The power consumption per ton goes down with an increase in rate of coal feed (Fig. 4). 3. The automatic air control in this stoker showed advantages Over manual control (Fig. 5). 4. The air velocities through the retort tuyeres were uniform around the retort, but varied between the rows of tuyere rings (Fig. 6). 5. The crushing of coal with this stoker increased the minus 1/8-in. from 11.6 per cent to 19.2 per cent. There was a notice- able segregation of coal as it came from the retort and distribution was not uniform (Fig. 7). 6. Measurement of air entering over the fire showed that with some not uncommon adjustments this air may be a high percentage of the total air supply (Fig. 8 and Table I). Operating Performance ,. The, removal of clinker was shown to give a period of poor performance characterized by a high fuel-bed resistance and a high percentage of carbon monoxide in the flue gases (Fig. 10). 2. The fuel-bed resistance was found to decrease as the fuel bed built up in a normal response to heat demand after a prolonged hold-fire (Fig. II). 3. with intermittent operation the burning rate was almost always less than the feeding rate, The wide variations were found to be greatly dependent upon the length of the On period and the length of the intervals between On periods (Fig. 12). 4. The maintenance of a low furnace draft led to less fly ash, better clinker formation, and higher efficiencies (Fig. 13). 5. The maintenance of a minimum of excess air also resulted in less fly ash, better clinker formation, and higher efficiencies (Fig. 14). Guiding Principles In planning the investigations, the following principles were adopted: Selection of Heating Equipment.—Use only standard models of stokers and furnaces representative of those widely used.

Jan 1, 1942

By Ellis Clark

These mines were discovered by George W. Lufkin in August, 1878, and mere worked almost continuously until August, 1893, a period of fifteen years. The ore-deposits in this locality lie close to the surface, and it is strange that the Spaniards, passing and repassing for nearly three centuries, did not discover them, particularly as they are marked in several places by large outcrops of black flint and iron, which would lead even a casual observer in a mineral country to suspect the proximity of ore. The mines are situated about six miles from the old Santa Fé trail, and fifteen miles from the Rio Grande. The geology is such that a skilled geologist passing over the ground would probably suspect, from the peculiar arrangement of the formations, the occurrence of ore. The formations in the district, eruptive and sedimentary, are nine in number; and of these eight are to be found within one thousand feet of the Grande workings. Topography. The topography of the Lake Valley district shows its dependence on the surface and underground geology, in a degree more marked

Jan 1, 1895

By W. E. Remmers

THE defect in wire known as "cuppiness" has appeared and disappeared from time to time but the exact cause of its appearance or disappearance has not heretofore been known definitely. This defect is not limited to one particular metal or alloy but seems to be found at various times in all of the materials that are drawn into wire. Fig. 1 represents cuppy wire made from an aluminum alloy and Fig. 2 illustrates copper wire that became cuppy in drawing without intermediate annealing. Differences of opinion are current among the manufacturers of wire and in the literature (see Bibliography at end of paper), some believing that the source of the trouble lies in segregation in the wire rod, oxides, sulfides, slag, etc. ; others that the trouble is created by an unevenness of stress distribution across the section of the wire, usually thought of as being a function of die shape. All of the published work is of a qualitative rather than a quantitative nature. With this in mind, it was deemed advisable to make a study of the shape of the wire-drawing dies, as well as the oxygen content of the copper to be drawn.

Jan 1, 1929

By Arvid E. Anderson, Frank K. Cameron

The weathering of copper-bearing ores with the formation of a water soluble salt and the recovery of the metal by leaching and evaporation precipitation, are processes long known, which have at various times been utilized with more or less commercial success. It is commonly assumed that the technical principles involved are quite simple and their application to a particular case easy. If these assumptions were true, we should anticipate a long list of uniformly successful leaching operations, which is not the case, and one would be justified only in calling attention to either unique or peculiar features of any particular enterprise; whereas, in studying the problems of the recovery of copper from the old workings of the Ohio Copper Co. in the light of available descriptions of other leaching projects, we have not found the assistance hoped for to guide us to the best realization of our own opportunities. It appears desirable, therefore, to call attention to these guiding principles, together with a description of the facts of accomplishment. Justification lies in the fact that the leaching and precipitation of copper at the Ohio Copper Co. mine in Bingham Canyon, Utah, is successful. This success is measured by an average monthly production of 600,000 Ib. of copper at an average operating cost, to date, of 3.85 cents per pound at the company's workings, or 6.32 cents per pound after paying shipping and smelter charges on the finished copper. A total of 17,076,099 Ib. of copper had been shipped to the smelter from the beginning of operations, August, 1922, to April 1, 1925. Probably these results can be bettered with further study and realization of the factors involved and the devising of methods of meeting them. It is now an assured fact that the enterprise will have a relatively long and continuously profitable life and that its reserves will not, nor cannot, be suddenly and permanently depleted. It is particularly noteworthy that the financing of this enterprise has been entirely from its earnings. The few hundred dollars necessary for the first experiments and production were obtained by the sale of used, and no longer needed, equipment from a mill belonging to the company; and all funds since required throughout the growth and development of the enterprise, from a few troughs with a production of 2000

Jan 1, 1926

By H. F. Perkins

CORROSION as an economic problem is growing rapidly in importance not only because it entails a replacement of corroded parts, but because it interrupts operation and causes hazards of damage and injury through failures of corroded equipment. Nearly all industries are interested in this subject. This paper, although concerned chiefly with corrosion in a complete refinery on the sea coast, discusses many of the various forms of chemical decomposition, most of which are found universally. The petroleum refining industry, in the writer's opinion, is seriously afflicted by the presence of sulfur in nearly all crude oils. The complete removal of sulfur from the crude as it enters the plant would cause such a relief from corrosion that this problem for research would seem to demand active attention. Next to sulfur the most troublesome corrosive agent is the saltmagnesium chloride, found in the water associated with crude oil, the last traces of which are so difficult to remove. This is much less serious, however, than the sulfur. Refinery corrosion can be tabulated as follows: Causes of Corrosion and Places of Attack in a Refinery CORROSIVE ELEMENTS MATERIAL ATTACKED Air: Pipe lines Moisture Tanks COQ Electric fixtures Sulfur gases Still towers Chlorine Condenser boxes Other acid-forming gases Ladders and stairs Smoke stacks Steel buildings All exposed metal Water: Alkali Fire lines and hydrants Acids Cooling water lines Air Condensers and coolers Other dissolved gases Boiler tubes Steam Steam pipes and valves Turbines Oil Sulfur compounds Tanks, pipes, valves, stills, towers MgCl2-Hydrochloric acid Condensers Air contamination

Jan 12, 1926

By K. B. Nowels

DATA pertaining to pipe line flow for both oil and gas in horizontal or nearly horizontal pipe lines are both extensive and accurate. However, the pipe formulas used to determine pressure drop for flowing fluids (liquids and gases) under conditions of horizontal flow are not generally applicable to flowing oil and/or gas wells. Except for some data developed in small laboratory set-ups, little or no information was available concerning flow of oil and gas mixtures in vertical pipes, when, in 1928, the writer began the development of a method for computing pressure drops in the pipes of flowing oil wells, as outlined in this paper. In 1929, Uren1 presented some findings based on laboratory experi-ments in which vertical flow of an extremely heavy crude oil was induced, by means of air through a 2-in. pipe, 41 ft. long. The friction factors derived from the experiment, as well as values for pressure drop per 100 ft., are not believed satisfactory in their application to flow problems in the field. Versluys,2 also in 1929, presented a mathematical development of the theory of flowing oil wells, and in 1930, gave a companion paper3 setting forth some principles governing the choice of length and diameter of tubing in oil wells. Since the termination of the study presented in this paper, another. valuable contribution pertaining to the problem, of vertical flow was made by Moore and Wilde.4 Although there were writings dealing with the principles of the gas-lift, a search of the available literature indicated that specific information on vertical flow was lacking with the exception of the writings of Uren, Versluys, and Moore and Wilde, to which reference .has been made.

Jan 1, 1932

By C. D. Woodward

PRIOR to 1902, the tramming of ore from the stopes to the shafts, in the Butte mines, was done by man or animal power, but the demand for greater tonnage and the need for more improved methods of tramming the ore resulted in the installation of the compressed-air locomotive in a few mines. Some of the objections to animal haulage were thus eliminated, but this locomotive was not generally adopted, because of a number of limiting features. It was superseded by the electric trolley locomotive. The early types of trolley locomotives were not satisfactory from the operators' viewpoint, but. they had many economic advantages and after a number of years, in which improvements were constantly made in the structural details, they became dependable, so that the use of animals for main haulage in the mines has now been practically abandoned. There are now operating in the Butte mines of the Anaconda Copper Mining Co., 185 trolley-type locomotives, each of which has replaced five animals. It cost approximately $50 per month to keep an animal, hence the economies that resulted fully justified the expenditure of electrifying the haulage systems. The haulage conditions in this company's mines possess many advantageous features that are not always found in underground mine workings. The drifts and crosscuts are large, approximately 5 ft. (1.5 m.) wide by 7 ½ ft. (2.3 m.) high, and have curves of no less than 18 ft. (5.5 m.) radius; and the roadbeds of the drifts and crosscuts have a uniform grade of 0.5 per cent. toward the shaft, or in favor of the loaded trips. In all of the mines, except two, the track gage is 18 in. (45.7 cm.), and all tracks in the main haulageways are laid with 25-lb. (11.3 kg.) rails. The stations at the shaft are large, well-lighted, clean, and cool. The average temperature of ,the stations, however, is about 75° F. (24° C.), while the average temperature in some of the drifts and crosscuts is 92° F. (33.5° C.).

Jan 2, 1922

By W. A. McCurdy, Joseph W. Leonard, T. S. Spicer

INTRODUCTION by T. S. SPICER and J. W. LEONARD Utilization from the Point of View of the Preparation Engineer The primary interest of the coal preparation engineer is process design, operation and maintenance to produce an acceptable product which can be sold at the greatest possible profit. Because of these interests, attention to the unrealized potential value of coal is sometimes neglected. The intent of this chapter is to alert the preparation engineer to the utilization possibilities of both the product as well as the tributary coal flow moving through a modem coal preparation plant. Thus, coal types having similar physical and chemical properties tend to become increasingly concentrated according to size, specific gravity and/or surface chemistry. The basis for any unrealized potential derives directly from the coal seam itself. Most coal seams consist of coal bands and benches which are sandwiched together in layers and which visually differ from each other by physical structure and luster. Randomly selected physical and chemical ranges determined from layers of coal in a single seam can easily differ by as much as 30 Hardgrove grindability indices, 0.30 of a specific gravity point, 20 percentage points of ash, six percentage points of dry volatile matter and 10 percentage points of total sulfur.1 Similar wide variations exist for calorific value, seam moisture content and undoubtedly, for any other conceivable coal test parameter which might be used for comparative purposes. These observations all point to the long established fact that coal is not homogeneous as conventional analytical methods might tend to indicate, but is instead a highly heterogeneous substance. Since both coal and associated minerals entering a preparation plant are highly heterogeneous, a multi-product potential beginning with certain flotation cell concentrates and ending with plant refuse may exist. Such a potential can and to a limited extent is achieved at some operations by drawing off coal according to size, specific gravity and/or surface chemistry differences from individual coal cleaning processes. Such a sequenced draw-off of coal of increasing ash increments could yield, among other products, chemical coals, coking coals, coals for pulverized fuel boilers, cyclone boiler coals, coals for cement plants, limestone, shale and clay. Basis for Coal Evaluation and Selection Despite the known heterogeneity of coal, the final analysis of a given coal product is customarily reported as a single average since coal has always

Jan 1, 1968

By J. D. Sullivan, Chester R. Austin, J. L. Nunes

IT is characteristic of most shales and surface clays that a bloated or vesicular structure is produced by burning to a sufficiently high temperature, usually about 150° to 200°F. above the normal maturing temperature used when making common brick of the same material. Early investigators recognized this phenomenon and attributed it to the release or expansion of gases in the vitrified piece. A number of theories have been advanced on the source of the gas and the manner in which it is evolved, but none has been proved conclusively. T. E. Jackson's1 early theory that bloating was due to oxygen evolved by ferric oxide in forming ferrous oxide or ferrous silicate was never substantiated by experimental work. Orton2 thought this explanation unlikely, as all clays should bloat at the same temperature, if this were true. He found that clays bloated in the range from 1100° to 1706°C. Also, many clays containing iron did not bloat at all, but fused. Orton also stated that oxidation of different types of carbon was a source of gas during burning of clay and should be expelled before vitrification was reached. Seger3 said that he had dissolved ferric oxide in glass with no apparent evolution of gas. Matson4 came to the conclusion that ferruginous days burned to vitrification in an oxidizing atmosphere did not have the iron in a ferrous form. According to Wilson,5 the gases responsible for bloating might be entrapped air, water, sulphur dioxide, carbon monoxide, oxygen, and hydrocarbons. He spoke of two causes of bloating: (I) incomplete oxidation below the vitrification temperature, and (2) decomposition of compounds and evolution of gases above the vitrification temperature. Decomposition of calcium sulphate was given as the principal example of the latter. Bleininger and Montgomery6 stated that if any time after vitrification began the evolution of gaseous matter was rapid enough so that pressure was produced within the day, bloating was certain to take place, and a vesicular structure would be formed. They recognized that initial structure, rate of heating, kiln atmosphere, and many other variables affected bloating. Work done by the Emergency Fleet Corporation7 indicated that clays suitable for bloating should be high in compounds of metallic oxides, carbon, sulphur, sodium, potassium, or other equivalent compounds, which either act as a flux or give off gas. Orton2 studied the chemical status of the carbon, iron, and sulphur in clays during various stages of burning, using a Huron shale that bloated under oxidizing or reducing conditions. He found that only a small amount of sulphur remained in the clay and that it could be detected only by dissolving the clay in hydrofluoric acid; i.e., the sulphur was in combination with the silica. Orton said that this sulphur was probably in the form of iron sulphide dissolved in the glass phase. The iron sulphide was not broken down by heating, but was dissolved in the basic slag or glass,

Jan 1, 1942

By G. C. Kuczynski

Two particles in mutual contact form a system which is not in thermo-dynamical equilibrium, because its total surface free energy is not a minimum. If such a system is left for a certain period of time, the bonding of the two particles will take place in order to decrease the total surface area, even though the temperature is lower than the melting point. This phenomenon of bonding of two or more particles with the application of heat only and at temperatures below melting point of any component of the system will be called sintering, although the powder metallurgists use this term in a broader sense, including the presence of molten phase and pressure. It is the objective of this paper to study this process and the mechanisms involved in it. This problem is of utmost importance to powder metallurgy and powder ceramics, and its technological aspects have been studied for a good many years. However, the powder metallurgical operations are too complex and include too many superimposing mechanisms, and too many variables for a direct study. It was therefore advisable for the purposes of this study to reduce the variables to a minimum. In this work the radius of the interface formed during bonding in a simple system composed of a spherical particle and a large block of the same metal was studied as a function of time and temperature. It is believed that the mechanism involved in this simple process is fundamental to any sintering operations. Previous Work J. Frenkell was the first to make a serious attempt to develop a theory of sintering. He assumed that the process consists of a slow deformation of crystalline particles under the influence of surface tension which reduces to a viscous flow where the coefficient of viscosity n is related to the self-diffusion coefficient D by the following equation 1 _ D D6kT [1] where 6 is interatomic distance, k the Boltzman constant, and T the absolute temperature. This type of viscous flow of a crystalline substance is essentially different from the ordinary plastic flow. The latter is a specific propcrty of crystals and cannot take place in amorphous bodies. According to Frenkel this viscous type of flow is due to the diffusion of the holes or vacancies arising in the lattice. He was able to derive an equation relating the growth of the interface between two spherical crystalline particles or between a particle and a semi-infinite crystal (Fig 1) to time t at constant temperature. This relationship can be written as follows: x²— 3/2 a/nt [2] where x is the radius of the interface assumed to be circular, a the original radius of the sphere and a surface tension of the material. The other assump- x tions are that x is less than 0.3 and that during the period 1 the original radius, a, of the metallic particle did not change appreciably. A. J. Shaler and J. Wulff2 followed closely the ideas of Frenkel in their theory of sintering of a mass of metallic powder. Neither Frenkel nor Shaler has validated his theoretical speculations with conclusive experimental data. Two measurements reported by

Jan 1, 1950

By Frederick V. Lawrence

Broadly speaking, fluxes are substances which promote wetting and spreading or enhance the fluidity and manipulative properties of materials in joining, fusion, and smelting operations. The term most strictly applies to substances used to facilitate soldering and brazing; but in the discussion which follows, the definition will be broadened to include materials used to promote fusion in arc welding and which aid in the smelting of metals. In each of the mentioned applications, the role of flux and its composition differs considerably; consequently in the discussion which follows, fluxes for soldering, brazing, arc welding, and smelting applications will be considered separately. Soldering and Brazing Fluxes Soldering and brazing are metallurgical joining processes in which a joint is formed using a filler metal of composition dissimilar to that of the joined pieces (base metal). The filler metal, solder or braze, is always of lower melting point than the base metal. Consequently with these processes there is no melting of the base metal. The successful creation of the joint depends upon the wetting and spreading of the solder or braze around or throughout the joint. The obstacles to successful soldering and brazing are contaminants and tenacious oxide films on the metal surfaces. In soldering and brazing applications, fluxes act to remove and exclude oxides and other impurities from the joint. There is no fundamental difference between soldering and brazing. The distinction between them is made on the basis of temperature. Join¬ing operations carried out at temperatures above 500° C are termed brazing. The higher temperatures used in brazing enable higher melting point filler metals to be used which generally provide stronger joints. The higher temperatures of brazing processes also require fluxes which are substantially different from those used in soldering and which will remain effective at high temperatures. Consequently the distinction between soldering and brazing can also be made on the basis of the fluxes used. Soldering Fluxes Even thoroughly cleaned and degreased, metal surfaces still possess tenacious oxide films which inhibit the soldering of the joint. The reactions which form these films increase with increasing temperature. The rate of oxide formation and the tenacity of the film varies with each base metal. Aluminum, magnesium, stainless and high alloy steels, aluminum and silicon bronzes when exposed to air form hard, adherent oxide films. These metals require highly active and corrosive fluxes to remove and prevent the reformation of oxide films during soldering. Metals such as copper and silver, on the contrary, form less tenacious films at a lower rate, so that mild fluxes are sufficient. Soldering fluxes must dissolve or remove oxides and promote rapid wetting and spreading of the solder. The flux removes the oxide film by loosening and floating it off into the main body of the flux. The oxide removal is often aided by the alloying of the solder with the base metal, undermining small areas of the oxide and detaching them (Anon., 1959). The flux coagulates and suspends the oxide particles which have been loosened. The flux forms a protective cover preventing the reformation of the oxides, permitting the spreading of the solder and the formation of a metallic bond between the solder and the base metal. Constituents of Soldering Fluxes: Soldering fluxes are commonly grouped into three categories: highly corrosive, intermediate, and noncorrosive. One normally would choose to use the least corrosive flux which will give satisfactory performance.

Jan 1, 1983

By Enid C. Plante, K. L. Sutherland

Practical metallurgists are unanimous in stating that oxidation of mined sulphide ore adversely affects separation of the constituent minerals under standard conditions in a mill. Frequently, the need for different operating conditions in different plants is attributed either to the extent of oxidation of the mineral surfaces or to soluble oxidation products. The investigations reported in this paper study the effect of oxidation on the primary condition necessary for flotation, namely that of adhesion of an air bubble to the mineral. Oxidation Products from Sulphide Minerals Weinig and Carpenter1 state that the relative rates of oxidation of some sulphides are: FeAsS > FeS2 > CuFeS2 > ZnS > pbS > Cu2S tetrahedrite. They also show that oxidation is rapid at first, that sulphides of older geological formations are more stable than those of recent date and that mixtures oxidize more rapidly than any one mineral alone. The more rapid oxidation of mixtures is borne out by results of Gottschalk and Buehler2 who found that PbS, zns and cus oxidized 8 to 20 times faster when pyrite was present. Only the end products of oxidation, sulphate and the metal ion, were identified in these tests. Gaudin3 states that pyrite is difficult to oxidize in the laboratory. Winchell,4 using pyrite containing 99.98 pct FeS2, allowed water to percolate through a bed of this ground material. At the end of three months only a trace of Fe+++and So4- were found. Even after 10 months the quantity of Fe+++ was only 7.8 mg and SO4- 25.5 mg per liter of percolating solution. Car-michael5 confirmed Winchell's results on the slow oxidation of pyrite. Usually the oxidation of sulphide surfaces has been conducted in neutral or acid solutions, the acidity being caused by the oxidation products. Eliseev and Nagir-nyaka found that treatment of pure pyrite with lime gave only thiosulphate, but considered it possible that thiosulphate was obtained not by oxidation of the pyrite itself, but by oxidation of calcium disul-hide. Mitrofanov7 suggested that thiO~ulphate was formed by the following process: oxidation of sulphide to give free sulphur (this frequently happens with marcasite), followed by oxidation of the sulphur to sulphur dioxide and reaction of the sulphur dioxide with sulphur to give thiosulphate. Ferric and ferrous hydroxides, or similar compounds, were shown to be present. The Research Staff of the United Verde Copper Co.8 treated ore from the United verde mine with lime. Thiosulphate, sulphide and polysulphides were stated to be

Jan 1, 1949

By G. Sachs, E. J. Ripling, S. I. Liu, J. J. Lynch

INTRODUCTION IT is a generally recognized fact that by repeated straining the fracture stress of any metal is reduced to a fraction of its value for static loading. The value of this fatigue strength depends upon numerous factors, such as the states of strain and stress, the limits of straining, and the number of repeated loadings or cycles.1.2,3 The large volume of previous investigations on fatigue is concerned entirely with large numbers of cycles, say exceeding 10,000 and up to 500,000,000. However, the changes in fracture stress and fracture strain occurring after a small number of repeated loadings have attracted little attention. Apparently the only investigations of this type are those byLudwik4,5 who subjected steel, copper, and aluminum wire to repeated torsion. In each test the torque was varied between identical values in the two directions of twisting. Depending upon the magnitude of the moment, fracturing occurred after a number of cycles, up to 10,000. The results of Ludwik's tests will be discussed later. Such torsion does not readily yield actual stress and strain values. Therefore a preliminary study was made of the effects of repeated direct stresses, i.e., tension and compression stresses. Because such tests are considerably more tedious than torsion tests, only high stresses were used in the preliminary work, yielding definite changes in metal properties after a few cycles. It was found in these tests that strains could be controlled more readily than stresses in the investigated range of high stresses. This paper, consequently, deals with the effects of strains of equal magnitude but different signs. The metal was first prestrained in static tension by a certain amount then subjected to compression to yield the same amount of strain after unloading. This loading cycle was repeated up to a maximum of seven times. After various cycles, the stress-strain curves in tension were determined, up to fracturing. As material for these tests, the aluminum alloy 24ST was selected because it exhibits a comparatively small amount of necking in the tension test. This alloy permitted, therefore, rather accurate control at the high strains necessary to produce fatigue failures at less than seven cycles. The results of this investigation appear sufficiently interesting to extend the method of attack in various directions. It is intended to study (I) the effects of cycles up to a few thousands and (2) the effects of strains of different magnitude in tension and compression, respectively. The latter variable also includes as a special case the effects of compressive strains.

Jan 1, 1948

By J. W. Gwinn

The underground haulage system in the lead-silver mine of the Bunker Hill & Sullivan Co., situated at Kellogg, Idaho, is the most extensive in the Cæur d'Alene district, comprising about 35,000 ft. of tunnels, drifts, and cross-cuts where locomotive haulage is in use. The main working adit, known as the Kellogg tunnel, or No. 9 level, is approximately 10,000 ft. in length from the portal to the main working shaft and extends beyond that about 4,500 ft. in each of two directions. On this level all the ore from the Bunker Hill & Sullivan, the Sierra Nevada Consolidated, and the Caledonia Mining Co.—an average of 44,600 tons per month—is hauled to the mills, located a short distance from the portal. Electric haulage by the trolley-type locomotive is used on this level throughout. Below this level, at intervals of 200 ft. vertically, are five other working levels (Nos. 10, 11,12,13, and 14); and the total haulage distance on these, where electric locomotives are used, is approximately 16,000 ft. No. 13 level has but recently been opened up and is not working to capacity; while the main cross-cut on No. 14 has just been started, and has no need of power haulage. Of the four levels equipped with electric haulage, Nos. 11, 12, and 13 have storage-battery locomotives, and No. 10 still retains the trolley type by reason of the small and decreasing tonnage produced. Electrification - Electric power is supplied to the mine by two systems. The first, the 500-volt direct-current trolley, was installed in 1897 when the Kellogg tunnel was started, and was extended as the tunnel advanced, being used for the first trainload of ore on Nov. 17, 1902, since which time it has been in continuous service. The current for the line is transmitted on a No. 00 copper wire as far as the shaft, and on No. 0 wire throughout the remaining length of the Kellogg tunnel; and until the installation of the storage-battery locomotives it was used in practically all the cross-cuts and drifts on each level below No. 9.

Jan 1, 1915