Search Documents

By Milton E. Wadsworth

The reduction of a metal in an aqueous solution may be defined in terms of its source of electrons. In electrolytic reduction electrons flow as a result of applied voltages. In other systems electron sources are reducing gases or metals of lower electron affinity. In the latter two cases it is increasingly evident that reduction is electro-chemical in nature in that electrons flow through one or more conducting phases even though the discharge reactions may not be rate controlling. Electrode processes impmtant in reduction and metal deposition are reviewed and extended to contact reduction by precipitation (cementation) on mme electropositive metals. Contact reduc- tion Processes are ancient in origin and represent techniques which have had successful application for hundreds of years. In spite of this, little has been understood in terms of the dynamic properties of such systems or of important side reactions. In recent years an increasing amount of laboratory work has been carried out on contact reduction and an attempt is made here to summarize the important findings. Undoubtedly the interest aroused recently may be attributed to improved reactor design which has demonstrated on a commercial basis accelerated metal recovery by taking advantage of important kinetic and chemical parameters. of gold cyanide (Au(CN)-), respectively. The most common method for the reduction of ions in solution is by electrolysis. The impure metal is made the anode and the ions are reduced at the cathode. Leach liquors containing reducible ions may similarly be reduced electrolytically using an inert anode. The latter process is referred to as electrowinning. In recent years hydrogen has been used successfully as a gaseous re-Metallurgical Society of the American Institute of Mining, Metallurgical and Petroleum Engineers (AIME). He was also a member of the Executive Board of the Metallurgical Society (AIME). In 1957 Dr. WADSWORTH was given the AlME Best Paper Award as coauthor of a paper on the roasting of sphalerite. On July 1, 1968 Dr. WADS-WORTH started an 24 month assignment in the Philippine Islands to assist in the development of graduate metallurgical engineering at the University of the Philippines. In addition to graduate teaching and curriculum development he will establish research programs to develop extraction processes for metals production in the islands. The project is supported by the Ford Foundation through Educational Projects, Incorporated.

Jan 1, 1970

By Dr. O’Neil Thomas J., Donald W. Gentry

"There is nothing so disastrous as a rational investment policy in an irrational world. " John Maynard Keynes INTRODUCTION The principles of time value of money concepts were discussed and illustrated in Chapter 3. The importance of time value of money and the role it plays in the discipline of engineering economy have been well established. Chapter 9 further illustrates that project evaluation criteria that fail to consider this aspect are seriously deficient in providing appropriate capital budgeting decisions. Obviously future project receipts and expenditures must be discounted to permit valid comparisons with current cash flows. Although the concept of discounting is widely accepted, the selection of the appropriate discount rate has been the source of considerable debate and much disagreement. This chapter, then, provides an overview of the theoretical and practical considerations for determining or calculating the appropriate rate for discounting, the corporate cost of capital. Note specifically that the assumption of constant business risk among the various investment alternatives is maintained throughout this chapter. THE INTUITIVE APPROACH The fact that interest exists suggests that all money has a cost associated with its use. The cost of this money may be the result of either explicit or implicit charges or some combination of the two. Any time an individual or a firm consumes less than the total earnings generated in a given time period that individual or corporation has either consciously or unconsciously decided to "invest" these excess funds in some type of activity. For instance, the individual or corporation may decide to invest the excess funds in tangible property, place the funds in a money market account, or simply keep these funds in cash. In any case, these funds have been invested in some type of activity which promises the investor some return on the invested capital. Remembering that the firm's primary objective should be the maximization of shareholder wealth, then intuitively it seems reasonable to suggest that the firm should not invest in any project where the anticipated return does not exceed the cost of funds (capital) committed to the project. Indeed, if the firm always invests in projects having returns in excess of the cost of capital committed to them, then the wealth of the firm (as measured by the price of common stock) should be in-

Jan 1, 1984

By A. A. Smith, J. S. Smart, A. J. Phillips

The preparation of a sufficient quantity of high-purity copper for extensive laboratory use in the study of impurity effects has been an integral part of a research program in progress at the Central Research Laboratory of the American Smelting and Refining Co. for the past several years. Initially the degree of purity desired was that necessary to meet two essential requirements for experimental accuracy: (I) no impurity was to be present in the copper in an amount capable of producing a detectable effect on the properties to be studied, (2) even the minor changes in propertties caused by the addition of extremely small amounts of individual elements were to be measured without significant interference from contaminants. In the course of the investigation, evidence has accumulated to justify the conclusion that this copper is of exceptional purity; in fact, no impurity can be detected by any chemical method or spectrographic technique known to the authors. The present paper describes the method of production of the copper and presents some of its properties. Preparation oF High-purity Copper The elimination of impurities from commercial electrolytic copper is accomplished by a three-step purification process. The first operation consists of an electrolysis through a purified CuSO4-H2SO4 electrolyte using bagged anodes, low current density (5 amp. per sq. ft.) and low tem- perature (25°C max.). The electrolyte is prepared from commercial Cu2SO4 and is purified by heating the solution and adding I gram Fe as Fe2(SO4)3 to each 100 grams dissolved copper; sufficient hydrated oxide of copper, prepared from Cu2SO4 and NaOH, is supplied to precipitate Fe(OH)3, which carries down arsenic, antimony, tin, selenium, tellurium and other minor constituents. Silver and remaining traces of lead are removed by addition of Na2S to the acidified filtrate, which is then reatered after standing hot overnight. The cold solution is made up to 30 grams per liter Cu and 150 grams per liter H2SO4 and introduced without addition agents into the electrolytic system, which consists of a double row of hard-rubber battery boxes with the inside webbing and partitions removed, arranged in seven steps of ascending height. Electrolyte is uniformly supplied to the two highest cells by glass siphons operating from a constant-level reservoir, and flows by gravity through the bank of seven cells, entering each at the bottom and leaving at the top. Glass air lifts continuously supply the returning liquid to the constant-level reservoir for recirculation. Each cell contains three 4 3/4 by 4 3/8 by 1-in anodes cast from high-quality electrolytic wirebar copper, various lots of which have been used with equal success. The anodes are pinned to sheet copper hangars shaped to prevent crawl of the electrolyte to the multiple-wired bus bars—which would result in fouling of the contacts—-and hang inside wool flannel bags stretched over open wood frames. The deposit

Jan 1, 1941

By Howard N. Evanenson

WASTE of coal, or perhaps more properly the percentage of its recovery in mining, has keenly interested me during an experience of over a half century in coal mining. In the early part of that time and until a few years after World War I the use of better mining plans and supervision resulted in an increased recovery. Since then, and particularly since controlled prices have been effective and mechanized mining has been extensively practiced, the tendency has all been toward lesser recovery in mining. The purpose of this paper is not to criticize the methods being used, which after all are the result of economic forces, nor to point to any particular mining method or to any locality where the best recovery is obtained, but to call attention to the tremendous waste of our national coal resources that is constantly continuing. A possible method of reducing this loss will be suggested, but without any illusions that it will be followed or that such attention will be paid to it. AVERAGE RECOVERY The actual average percentage of recovery of coal in the ground is not known. Many papers giving figures of recovery have been published but no definite base of measurement has been agreed upon by those presenting the data. As an example, the Pittsburgh bed can be considered-it has produced a greater output than any other and has been sold or leased at higher prices in place than any other coal bed. Over a large part of its area where high prices have prevailed it has been customary for many years to draw pillars. In these mines it has been usually assumed that the average recovery was higher than in other beds and percentages of recovery of 85 to 95 per cent are usually said to be realized. In all of these mines, however, it has been customary to leave from 4 to 8 in. in the bottom, and 10 to 16 in. in the roof; thus from 14 to 24 in. of coal has been left in the mines out of a total seam thickness of about 96 in., or 15 to 25 per cent of the total coal. No attempt is made to mine this coal for metallurgical purposes, owing to its impurities, although this unmined coal is of excellent quality and better than much of the coal used for steam purposes. If we add to this discard the coal usually lost in mining and those parts of the bed that must be left because of bad seam conditions, barrier pillars, reservations for buildings, gas and oil wells, one can easily see that the actual recovery from this bed cannot be more than 65 to 75 per cent. In addition, the rooster or rider or wild coal over the mining bed often 3 to 4 ft. of coal, high in ash and sulphur but still good steam coal-is hardly ever mined, nor is any attempt made to mine it. It is safe to say that from a broad economic viewpoint not more than 50 to 60 per cent of the fuel value of this great seam is actually being recovered.

Jan 1, 1946

By A. J. Phillips, A. A. Smith, J. S. Smart

The preparation of a sufficient quantity of high-purity copper for extensive laboratory use in the study of impurity effects has been an integral part of a research program in progress at the Central Research Laboratory of the American Smelting and Refining Co. for the past several years. Initially the degree of purity desired was that necessary to meet two essential requirements for experimental accuracy: (I) no impurity was to be present in the copper in an amount capable of producing a detectable effect on the properties to be studied, (2) even the minor changes in propertties caused by the addition of extremely small amounts of individual elements were to be measured without significant interference from contaminants. In the course of the investigation, evidence has accumulated to justify the conclusion that this copper is of exceptional purity; in fact, no impurity can be detected by any chemical method or spectrographic technique known to the authors. The present paper describes the method of production of the copper and presents some of its properties. Preparation oF High-purity Copper The elimination of impurities from commercial electrolytic copper is accomplished by a three-step purification process. The first operation consists of an electrolysis through a purified CuSO4-H2SO4 electrolyte using bagged anodes, low current density (5 amp. per sq. ft.) and low tem- perature (25°C max.). The electrolyte is prepared from commercial Cu2SO4 and is purified by heating the solution and adding I gram Fe as Fe2(SO4)3 to each 100 grams dissolved copper; sufficient hydrated oxide of copper, prepared from Cu2SO4 and NaOH, is supplied to precipitate Fe(OH)3, which carries down arsenic, antimony, tin, selenium, tellurium and other minor constituents. Silver and remaining traces of lead are removed by addition of Na2S to the acidified filtrate, which is then reatered after standing hot overnight. The cold solution is made up to 30 grams per liter Cu and 150 grams per liter H2SO4 and introduced without addition agents into the electrolytic system, which consists of a double row of hard-rubber battery boxes with the inside webbing and partitions removed, arranged in seven steps of ascending height. Electrolyte is uniformly supplied to the two highest cells by glass siphons operating from a constant-level reservoir, and flows by gravity through the bank of seven cells, entering each at the bottom and leaving at the top. Glass air lifts continuously supply the returning liquid to the constant-level reservoir for recirculation. Each cell contains three 4 3/4 by 4 3/8 by 1-in anodes cast from high-quality electrolytic wirebar copper, various lots of which have been used with equal success. The anodes are pinned to sheet copper hangars shaped to prevent crawl of the electrolyte to the multiple-wired bus bars—which would result in fouling of the contacts—-and hang inside wool flannel bags stretched over open wood frames. The deposit

Jan 1, 1941

By J. Szekely, Robert G. Lee

Calculations are presented for the prediction of the combined radiative-convective heat loss from molten steel held in a ladle, covered by initially molten slag. A mathematical formulation is given and numerical solutions are presented of the resultant partial differential equation. It is shown that a slag layer of about 2 to 3 in. thick would prevent any significant heat loss from the top metal surface for Periods of up to 1 hr. For thinner slag layers (or in the absence of the protective slag) appreciable heat losses would occur; plots given in the paper allow the quantitative evaluation of the fall in the metal temperature due to this heat loss. The ironmaking-steel processing operation involves several intermediate steps where the molten metal is held or transfered in ladles. The quantitative assessment of the heat losses occurring under these conditions is of considerable importance in process design considerations since the temperature of the metal may have to meet rigid specifications in any given part of the processing sequence. While the metal is held in a ladle, heat is lost by two mechanisms: i) by conduction into the ladle walls; ii) by radiation and convection from the top surface. Calculations relating to the conductive heat loss, are readily made and information may be found in the literature both on data pertaining to metallurgical situations and on the techniques that are available for performing additional computations.'-3 The evaluation of the combined convective-radiative heat loss is less straightforward, especially when there exists a protective slag layer covering the metal. In this latter case, as the top slag surface loses heat partial solidification may occur and the latent heat thus given up by the slag may represent an effective barrier to the heat loss from the metal. A semiquantitative assessment of the role played by the slag in preventing heat loss from the metal has been made in a previous paper,4 where it was shown that a slag layer 6 in. thick would act as a near-perfect insulator for periods of up to 2 hr. However, this first paper reported on an essentially preliminary investigation that considered only one slag layer thickness; furthermore, the boundary conditions used for the calculations were somewhat restrictive since no allowance was made for changes in the metal temperature. The purpose of this second paper is to extend the scope of the preliminary investigation previously reported by: i) considering several slag layer thicknesses; ii) allowing for variations in the metal temperature; and finally iii) by performing calculations on the net loss from the metal rather than from the slag surface. FORMULATION Consider a slag phase extending from y = 0 to y = L1, covering a metal phase that occupies the region extending from y = Ll to y = L, as illustrated in Fig. 1. Denote the slag and metal temperatures by TS and Tm and assume that initially both slag and metal are molten, having a uniform temperature Ti. At time = zero the surface represented by the y = 0 plane (top slag surface) is brought into contact with cold air, the temperature of which is given as T Thus for time > zero, heat will be transfered from the slag to the air by natural convection and thermal radiation; as a result of this heat loss the slag temperature will fall and after some time a solid phase is formed and a solidification boundary (more realistically described as a zone) will move progressively from y = 0 toward y = L1. Once the region of temperature gradients, moving ahead of the solidification zone, reach the slag-metal interface (y = L1) there will be a net transfer of heat from the metal through the slag to the cold air, across the y = 0 plane. In the formulation of moving boundary problems, involving phase changes that require computer solution, it is convenient to represent both phases by a single equation, making allowance for the latent heat of solidification by assigning an appropriate temperature dependence to the specific heat content. Thus the energy equation for the whole of the slag may be written as follows:

Jan 1, 1969

By C. C. Miller, C. A. Hutchinson, A. B. Dyes

This paper presents a simple and practical method for the estimation of effective permeability and reservoir pressures from the rise in bottom hole pressure when a well is shut in. The equations on which the method is based consider a stabilized well producing from a uniform formation when flow is shut off at the sand face followed by the unsteady state flow of a compressible liquid from regions far from a well to points closer in during the build-up period. An approximation for extending the method for two-phase flow in the reservoir is also presented. In order to compare the effects of extremes in the conditions existing at the external boundary of the reservoir, two general equations defining the build-up characteristics of the reservoir are shown: first, for a reservoir with constant pressure maintained at the radius of drainage, and second, for a reservoir with no influx over the radius of drainage. These equations define the build-up characteristics for the shut-in well as functions of time, the permeability and porosity of the drainage area, the viscosity and compressibility of the fluid flowing, the production rate prior to shut-in, and the radius of drainage of the well. When a well is shut in at the surface, rather than at the sand face, flow will continue into the well bore after shut-in due to the compressibility of the fluids in the bore. The effect upon the build-up characteristics of this flow into the well after shut-in has been studied by means of the electrical analyzer. From the basic relationships and consideration of the limitations due to continued flow after shut-in, methods have been developed for the estimation of the effective permeability of an area from the plot of the build-up characteristics of the well producing that area. Likewise, equations are presented for the estimation of the reservoir pressure from an early portion of the build-up curve. By means of the electrical analyzer, studies have been made concerning the variations in the build-up characteristics resulting from damaged or improved permeability in the immediate vicinity of the well bore. These studies show that the permeability estimated from the build-up curves represents the average effective permeability of the majority of the drainage area and that the permeability in the immediate vicinity of the well does not affect this estimate. Since the major portion of the drawdown is in the immediate vicinity of the well, variations in the permeability of this area greatly affect the productivity of the well as expressed by the productivity index. From pressure build-up and productivity index tests the ratio of the permeability of some arbitrarily specified area immediately surrounding the well and the effective permeability in the majority of the drainage area can be computed. This ratio is a measure of any damage to the well incurred in drilling or reworking and likewise any benefits as a result of remedial measures. In build-up tests and many other tests that are made on a producing well it is at least desirable, if not absolutely necessary, that the well be stabilized prior to testing. A method is presented here for relating the time required to stabilize a well and the reservoir and fluid properties. An equation is also presented for approximation of the length of shut-in time required to obtain sufficient build-up data for the estimation of permeability and reservoir pressure.

Jan 1, 1950

By J. L. Byers

THE metals of the platinum group, ruthenium, rhodium, palladium, osmium, iridium and platinum, have certain characteristic effects on the structure and appearance of the cupellation bead. These effects are dependent not only on the particular member of the group present, but also on its concentration. The extent and quality of these effects, however, have not received investigation worthy of their importance. Bannister and Patching investigated the qualitative effects of some of these metals in 1913. This paper deals with a possible rapid qualitative and quantitative delineation of the effects exerted on the cupellation bead by the various metals in the platinum group. More particularly, the paper correlates quantitatively the presence of such metals with their characteristic structural and surface effects. Inasmuch as the various members of the group are more frequently associated with gold than with silver, the gold button will receive the greater consideration. The past two decades have seen a much greater interest manifested in this particular group of metals, because of the increased industrial demand for them. There has been a constantly growing need for assays that will not only express the silver and gold values but also indicate the presence of the other metals. If these indications could be used as quantitative criteria, their value would be greatly increased, because the methods of chemical analysis for the elements of the platinum group are lengthy, involved and unreliable except when the analysis is made by an expert and experienced rare-metal analyst. The materials for the X-ray diffraction examination of the different beads were prepared by filing and screening to -200 mesh, and were then rayed in a General Electric X-ray diffraction unit. The filed samples were treated magnetically in order to eliminate any possible steel particles. Sodium chloride and C.P. gold and silver were used as standard comparators.2 (See Table 1.)

Jan 1, 1932

By S. J. Aresco, A. A. Orning

This paper presents a scheme for determination of precision for the various steps in coal sampling, sample preparation, and analysis. It also gives resulting data on 100 American coals from various coal fields covering all ranks from lignite to low-volatile bituminous. The variances of the various steps of the process are studied as a function of the coal preparation, the top size of coal as sampled, and the ash content. To understand the variables that control the overall precision of coal sampling, sample reduction and analysis, there is need for measurement of the variances due to successive steps of the process. The purpose of the present paper is to study these variances as a function of the coal preparation, the top size of the coal as sampled, and the ash content. The sampling data included are from 100 individual preparation plants from various coal fields in the United States and cover all ranks of coal from lignite to low-volatile bituminous; various types of preparation; deep, strip and auger mines; and, in many instances, mixtures from two or more coal beds or types of mines. The collection of data is continuing, but sufficient work has been done to draw some important conclusions. Since the work was done in conjunction with the Bureau of Mines sampling of coal, the time that could be spent at any one preparation plant was limited. This, in turn, limited the number of samples that could be taken for each coal. For this reason, the results from various plants were pooled in blocks. Two major blocking categories were "raw" and "washed!' The raw coals had no preparation other than crushing and/or screening. The washed coals were prepared further by some combination of wet or dry mechanical cleaning. The categories were further divided ac-

Jan 1, 1965

By C. L. Horn

All the tunneling machines discussed in this chapter are Calweld machines. I have divided the machines designed for soft-to-medium formations into three groups, correlated with typical formations and estimated compressive strengths: 1) Running sand and silt requiring blade cutters. 2) Chalk and clay, requiring drag-type cutters, for compressive strengths ranging from 0 to 2000 psi. 3) Sandstones, hard clays, and limestones, requiring disk cutters, for compressive strengths ranging from 2000 to 20,000 psi. All of the soft-to-medium-formation machines are powered by electric hydraulic or diesel hydraulic pumps. The pumps in turn drive hydraulic motors for the torque reaction. The soft-to-medium-formation machines are enclosed in a 360º shield with a cutting shoe, a center portion of the shield housing the mechanical equipment, and a tail shield. Tunnel supports are usually required, and the supports are normally erected in the tail shield. Our first machine was built in 1963, and all but five of the machines illustrated in the figures with this chapter were built in 1966 and 1967. Some of the machines manufactured will be discussed in the following sections; specific features as development has progressed will be pointed out. HARD-ROCK MACHINE The Calweld hard-rock machine has been designed along basic and accepted principles for a hard-rock mechanical borer. The features of this machine are as follows: 1) Direct electric motor drive through a planetary gear reduction ring and pinion gear, and then to the cutting head through a hollow drive shaft.

Jan 1, 1970

OF the many categories into which scientific knowledge has been arbitrarily divided, the one that has proved most applicable in our attempts to gain an insight into the details of steelmaking processes is physical chemistry. To the young metallurgist who has only recently survived a college course in this subject, the application of its principles to reactions in liquid steel should not be difficult. The many excellent textbooks of physical chemistry1 provide ample material on the principles of this subject for the student of steelmaking processes; yet with- out exception those texts draw the greater part of their illustrative material from room-temperature data. It is the purpose of this chapter to provide a concise review of the principles of physical chemistry that have proved useful in studies of steelmaking and at the same time to present some of the data that are essential to the application of physical chemistry to reactions at high temperatures. Certain parts of physical chemistry require rather formidable mathematics for their complete development but in the main those applications in which we are here interested can be ex- pressed in fairly simple terms. Many of the equations are logarithmic, and we shall use the symbols In X and log X to represent respectively the natural and the common logarithm of X. The two are related by In X = 2.303 log X. Logarithms may be read from a good slide rule with sufficient accuracy for our purposes. In a few spots a knowledge of the calculus will be, helpfu1. MATTER Innumerable forms of matter may be recognized, and these may be converted by chemical processes into others, as ore, coke, and air are converted into iron, slag, and gas. Whenever such a change occurs there is a corresponding change in

Jan 1, 1951

By C. D. King

THE advances in basic open-hearth practice which have occurred during the past decade, have been principally in the direction of the physical development of larger units, decreased fuel consumption and increased length of furnace campaigns. However gratifying the economies which have resulted from these improvements are, it must be granted that like advances in chemical, physical-chemical and metallurgical knowledge pertaining to open-hearth practice would have added greatly to these already attained reductions in costs and it is the purpose of this paper to stimulate the discussion and study of these phases of cooperation and their relation to ingot yields. DEFINITION AND CALCULATION OF LOSSES In discussing basic open-hearth yields, the writer refers to the ingot yields obtained in the process of converting pig iron, scrap and ore to steel. In this conversion some scrap is produced. The difference between the total weight of ingots and scrap produced and the original metallic charge, constitutes the loss. Basic open-hearth yields vary considerably, and while some of the divergencies can be explained metal-lurgically because of differences in slag analysis, proportion of pig charges, and character of steel made, there often remain differences that apparently do not readily disclose themselves. In reconciling yields by the usual accepted methods of calculation, one is often forced to evaluate large losses as "unaccounted for." Such a system is prone to accept these large losses as inherent to the operation and not susceptible of appraisement and correction. It is obvious that if the ingot yield, plus the scrap recovered subtracted from 100 per cent. determines the loss, a drop in yield may increase the loss or the scrap, and both contribute to a monetary loss. To be sure, one can subdivide these unaccounted for losses into various components, such as metallics in slag, free Fe in slag, spillage, etc., and thereby reduce the amount of losses considered as "unaccounted for." If empirical methods could indicate with some degree of accuracy all losses, however slight, the undetermined losses become appreciably smaller and resultant calculations will check actual

Jan 1, 1929

By G. Y. Chin, A. R. Von Neida

The relationship between crystalline texture and magnetic anisotropy in a Co-10 pct Fe alloy has been investigated by means of X-ray pole .figures, hysteresis-loop tracings, and magnetic-torque measurements. The cold-rolled texture of a sample after 95 pct thickness reduction is of the "brass-type", that is, (110)[112/. At 500°C the cold-rolled material recrystallizes fo a (311)[112] orientation, plus a minor (210) [001] component. This texture persists up to 1400°C, at which point it is replaced by a secondary recrystallization texture, (111)[112]. Magnetic-torque measurements show that the easy axis of magnetization lies along the transverse direction in the cold-rolled material and changes to the rolling direction after primary recrystallization. This change is also reflected in the hysteresis characteristic; whereas the B-H loop along the rolling direction is nonsqunre with a low value of residual induction (750 vs 19,000 gauss at saturation) for the cold-rolled material, the annealed sample exhibits a very square loop with a large value of residual induction (16,000 gauss). Theoretical torque curves calculated from the texture data on the basis of magnetocystalline anisotropy energy compare favorably with the observed torque curves. Values of the crystal anisotropy constant, K1, obbained from these comparisons are in good quantitative agreement with single-crystal data. Furthermore, both the torque amplitude of the poly crystalline material and the magnitude of K1 obtained with a single crystal show an identical temperature dependence over the range from —195° 10 +220°C. It is thus conclzided that the magnetocrystalline anisotrpopy erlergy is chiefly I-esponsible. for the nzagnetic allisotropy of textured Co-10 pct Fe alloy. THE results of a study on the relationship between crystalline texture and magnetic anisotropy of an alloy containing 90 pct Co and 10 pct Fe by weight are reported in this paper. The alloy has an fcc structure: a saturation magnetization of about 19,000 gauss,2 and a nearly zero value of magnetoatriction.3 puzey4 has reported a large crystalline anisotropy constant of -7.4 X 105 ergs per cu cm, indicating the (111) directions as the easy axes of magnetization. Hence, texture variations resulting from cold work and annealing are expected to introduce a magnetic anisotropy in the polycrystalline material through the alignment of these (111) easy directions. In the present study, magnetic anisotropy of both rolled and annealed strips was determined by magnetic-torque measurements and hysteresis-loop tracings The results were correlated quantitatively with calculations based on texture data obtained by the X-ray pole-figure technique. It is established that the magnetocrystalline anisotropy energy is indeed chiefly responsible for the magnetic anisotropy of textured Co-10 pct Fe alloy. EXPERIMENTAL Material Preparation. The alloy nominally contained 90 pct Co, 9.5 pct Fe, and 0.5 pct Mn by weight, the chief impurities being nickel, carbon, and silicon (less than 0.5 pct total). The cast ingot was rolled to strips 10 mils thick and slit to 1 in. in width. After strand annealing at 900°C in a hydrogen atmosphere, these strips were further cold-rolled to various thicknesses down to 0.25 mil (97.5 pct reduction), with samples collected for study of the development of texture and magnetic anisotropy as a function of thickness reduction. subsequently, samples cold-rolled from 100 mils to 0.25 mil were obtained, thus extending the thickness reduction range to 99.75 pct. In addition, 0.5-mil samples cold-rolled from 10 mils were selected

Jan 1, 1965

By G. D. Tranter

THE importance of ladle and teeming practice and its relationship to the yield and quality of the product has focused considerable attention on this phase of open-hearth operation. Inherently bad steel cannot be perfected by manipulation during teeming, but through the proper handling of "off heats" serious losses, which in many cases arise, can be greatly minimized. Conversely, good steel can be ruined as a result of poor workmanship in the -pit, and rendered unsuitable for application to orders where exacting surface and physical requirements are essential. Sufficient evidence has been secured to warrant careful investigation of pit operations when an abnormal number of basic defects develop in the semifinished product or are found on the inspection table. This paper is written from the standpoint .of the problems involved in making satisfactory ingots for sheet rolling. Furthermore, since rimming steel is almost universally used for manufacture of sheet, the data presented herein refer particularly to that type of steel. Going back some 30 years in open-hearth .history, the word "pit" properly described the hole into which the ladle was lowered to receive the metal tapped from the furnace. The working space was necessarily limited, and the disposal of slag and metal overflowing the ladle or lost because of breakouts presented a serious problem. The equipment, of course, was rather crude in the light of present-day development. Often no ladle stands were provided, the ladle being hung on the crane beneath the runner. When delayed taps occurred, the ladle was often held in position for many hours, thus tying up pit operations and causing delays to adjacent furnaces. Because of crowded conditions, slag thimbles were not used and a network of cast-steel triangles and old tapping bars, placed before the heat was tapped, was the means for removing slag from the hole. Furnaces of different capacities were used

Jan 1, 1935

Endurance Properties of Non-ferrous Metals Discussion of the paper of D. J. McADAM, JR., presented at the Syracuse Meeting of the Institute of Metals Division and issued, as Paper No. 1506-E, with Mining AND METALLURGY, October, 1925. H. F. MOORE,* Urbana, Ill. (written discussion).-This paper says that the results of tests of some non-ferrous metals made at the University of Illinois and in the testing laboratories at McCook Field are abnormal when they do not agree with results obtained by the author. The writer can see no inherent reason why results of the type uniformly reported in this paper are any more normal. The knee of the stress-cycle, or S-N, diagram comes at different lengths of life for different metals. For steel, the knee occurs after from 1,000,000 to about 4,000,000 cycles, and for some non-ferrous metals not before 20,000,000 of 30,000,000 cycles. The writer can see nothing inherently abnormal in finding certain nonferrous metals for which the knee of the stress-cycle curve occurs at 1,000,000,000 cycles or higher. The writer has had experience with both the cantilever and the rotating-beam testing machine, and thinks that the chance of whipping action in the cantilever machine is a more serious handicap than the action the author has pictured. Furthermore, the writer questions whether the author has shown the existence of a clearly marked and definite endurance limit for all the non-ferrous metals he has tested including the following: Nickel (CT and A-16, Fig. 1), monel (CS-5.5 and BK, Fig. 2), bronze (F and CK, Fig. 4), aluminum bronze ( BJb-1, Fig. 6).

Jan 12, 1925

By A. C. Fieldner

The usual analytical determinations made in analyzing coal are comprised in the proximate and ultimate analysis and the determination of calorific value. The proximate analysis includes determinations of moisture, volatile matter, fixed carbon and ash; the ultimate analysis determinations of carbon, hydrogen, oxygen, nitrogen, sulfur and ash. The proximate analysis is an empirical process, and the results, obtained, especially for volatile matter and fixed carbon, are materially influenced by the method of conducting the determinations. This fact must be kept in mind in comparing coal analyses from different laboratories for classification purposes. Unless these laboratories have followed the same standard methods, differences due to variations in the analytical methods may, in some cases, be large enough to change the classification of the coal. It is unfortunate that the determinations of volatile matter and fixed carbon, which are most used in coal classification, are the two that are subject to the greatest variations. Moisture comes next in the scale of variation. The ultimate analysis is less influenced by method, since it comprises the determination of definite chemical elements. However, even the ultimate analysis is subject to errors, due to inorganic carbon from carbonates and to inorganic hydrogen from water. For these reasons it is evident that the Committee on Coal Classification must be informed on the evolution of methods of coal analysis in America and the average differences in analytical results that may be expected from variations in methods. The committee should be informed also on which methods should be accepted as standard for classifying coal and on the magnitude of the differences in results that may be expected from duplicate samples of the same coal sent to different laboratories for analysis by the same standard methods. It is the purpose of the present paper to provide this information. The subject will be discussed under the following main heads:

Jan 1, 1930

By Kenneth A. Gutschick, Robert S. Boynton

Lime has become a very general loosely used term that includes virtually all finely divided types of limestone as well as burned forms of lime. However, actually (and according to Webster) lime is only calcined limestone (known as quicklime, calcium oxide, or calcia). It also embraces the secondary product of quicklime-namely hydrated lime or slaked lime (calcium hydroxide). Fundamentally there are two basic types of limestone-high calcium and high magnesium (dolomitic). The dolomitic quicklimes and hydrated limes correspond to their "high calcium" counterparts in terminology. The only difference is that the dolomitic types are a combination of the elements, calcium and magnesium, whereas high calcium limes contain only 0.5 to 2 pct of magnesium. Both types, which are made through calcination at high temperatures, are shown chemically as follows: [ ] In the above reactions either type of limestone is burned (calcined) in furnaces known as lime kilns so that the carbon dioxide content of the stone is expelled as a gas. With high calcium limestone 44 pct of the weight of the stone is lost; with dolomitic limestone the loss is 48 pct. Also, both reactions are chemically reversible since quicklime absorbs carbon dioxide readily and in so doing transforms itself back to the original carbonate form. Most quicklimes correspond in size to the original rock, which varies from one to eight inches, depending upon the type of kiln employed. Slight shrinkage occurs during calcination. A few limestones when calcined will decrepitate (break down) into a fine powdery form; however, chemically it is the same as the pebble or lump-sized material. For some uses the lumps of quicklime are pulverized and packaged. Since quicklime has great affinity for moisture and carbon dioxide, it is perishable. When it absorbs moisture and CO2 slowly from the atmosphere, the resulting product is "airslaked." Consequently, it must be stored in dry, moisture-proof areas that are free from CO2. Because of its fugitive nature, it is usually consumed within a few weeks after manufacture. A more stable form of lime is hydrated lime-commercially made by adding water to the quicklime until the lumps break down into a fine, white powder. This product, which is usually packaged, can be stored more safely; and many of its uses are interchangeable with quicklime. The irreversible chemical reactions of hydrated lime for both high calcium and dolomitic types are: [ ]

Jan 1, 1960

By William Meyer

DURING the seventeen years Fairchild has been making aerial surveys and aerial photographic equipment many changes and improvements have been made in the equipment and in the technique of using it. Aerial photographic operations have, been developed from the crude methods and apparatus of wartime to a science upon which exact engineer-ing calculations can be based. In the early days there was no airplane specially designed to meet aerial photographic requirements. One type of camera was used in all types of operations. Projects were made to conform to the products we could make and deliver. Obviously this entailed definite limitations. We now have airplanes designed specially for aerial photographic operations in accordance with the conditions encountered in actual service experience. We have cameras for every photographic purpose and our laboratory technique and equipment have been developed to the point where it is possible to make contour maps from vertical aerial photographs with contour intervals as close as five feet. Therefore, the selection of equipment now can be made in accordance with the time at our disposal, the justifiable cost, the scale desired, and the general requirements of the client. This does not mean, however, that arbitrary deliveries can be set for aerial photographic products. The factors in each project vary and each job must be considered in accordance with the factors involved. This accounts for the fact that products of aerial surveys may readily vary from $1.25 to $200 a square mile. Sherman M. Fairchild became interested in the design of a radically different aerial camera during the war. At its close he was about ready to demonstrate his design to the War Department. By 1920, he was able to bring out his first completed model. It had a number of radical features, one of which was the first dependable high-speed between-the-lens shutter, which makes for accuracy in aerial photographic work. Another feature was the use of metal throughout, eliminating the wooden parts that had been employed in wartime cameras. Another feature was the interchangeability of parts.

Jan 1, 1936

By Charles Malcolmson

IN the United States, improvements in methods of combustion have made possible the use of the smaller sizes of anthracite. This coal is now being reclaimed from the culm banks accumulated by the miners in the more prodigal years. To-day the freshly mined "slush" is segre¬gated and stored against the time when, through briquetting or some other means, it may be utilized as a commercial fuel. For many years Europe has achieved enviable results in the conservation of her small coal by briquetting. This profitable industry, first placed on a commercial basis in France in 1842, has grown to such magnitude that the "'production during 1913 exceeded 36,000,000 short tons. Briquets produced in Great Britain are used principally for steam purposes and the greater part of the output is either exported or taken by the Admiralty. On the Continent the briquets are sold in all branches of the coal trade. The French railways consume from 16 to 40 per cent. of their fuel in briquet form, and in Germany the lignite briquets, (of which 20,000,000 tons were produced in 1913) are as familiar and interesting to the American traveler as the huge porcelain stoves in which they are burned. It is not strange, therefore, that aggressive and far-sighted promotors heralded this means of converting our ubiquitous coal waste into revenue. To the superficial student the briquetting process seemed so simple as to make success easy. It is for this reason principally that the path of the industry in America has been strewn with failures. The rash enthusiast had either not worked out the mechanical problems involved or had failed to grasp the local conditions under which the briquets must be made and sold. It may be confidently stated, however, that in spite of the many vicissitudes through which it has passed, coal briquetting is now an accomplished economic success in this country, and as a profitable adjunct to the coal industry it has come to stay.

Jan 2, 1915

By J. G. Kura, P. D. Frost, L. W. Eastwood

THE preparation and general properties of mag-nesium-lithium base alloys have been described in earlier papers.l,2 Lithium forms solid solutions with pure magnesium, lowers its density, and improves its formability. The improved isotropy and better formability of these alloys are attributed to the ability of lithium to convert the magnesium lattice from hexagonal (alpha phase) to body-centered-cubic (beta phase) structure. This change in structure is, however, accompanied by some limitations which make it difficult, if not impossible, to combine all the desirable properties in one composition. The most important of these limitations is the tendency of the high-strength alloys to overage and lose strength at temperatures of the order of 150" to 200°F. A study was made of the effects of composition and heat treatment on the age-hardening character- istics of the alloys. Stability of properties was determined by hardness and tensile testing of the alloys at room temperature after various periods of exposure to elevated temperatures. Precipitation-hardening curves are shown for a number of magnesium-lithium-zinc alloys in fig. 1. All the alloys were solution treated at 700°F and aged at 200°F. The reason for using the drastic quench, toluene at —105°F, was to reduce the rapid precipitation of the hardening phase which takes place at room temperature. Previous experience has shown that a hardness of about 90 to 95 Rockwell E is indicative of yield strengths in the range of 35,000 to 40,000 psi. This is a desirable strength level for magnesium-lithium base alloys with specific gravities of the order of 1.65 or less. Although the hardness level is raised by large zinc contents, a corresponding decrease in the cold formability of these alloys occurs with increasing zinc content. Thus, alloys containing up to about 10 pct Zn and having hardnesses of the order of 70 to 75 Rockwell E may be bent over a very small radius; however, the 16 pct Zn alloy requires a bend radius of more than 8T. The hardness relationships illustrated in fig. 1 can be explained by the partial phase diagram shown in fig. 2. The zinc-free magnesium-lithium alloys are entirely body-centered cubic (100 pct beta phase) when the lithium content of the binary is 10.3 pct or higher. When zinc is added, the boundary between the beta and alpha-plus-beta fields at 700°F is shifted to lower lithium contents. In addition, the solubility of zinc at 700°F decreases with decreasing lithium content. It is shown by a study of micro-structures that the limit of solid solubility of zinc in the 6Mg/Li* matrix at 700°F is about 19 pct. It is believed from a study of the aging curves that this limit is about 2 to 3 pct at 200°F. Zinc additions up to about 2 pct increase the hardness and strength of the magnesium-lithium base solid solution. Larger additions of zinc effect further hardness increases, but precipitate during prolonged aging at 200°F. Alloys having 14 to 20 pct Zn precipitation harden extremely rapidly during quenching, even when quenched in toluene at —105° F. The high-zinc alloys overage at 200°F but are capable of maintaining reasonably high hardnesses for periods up to 1000 hr.

Jan 1, 1951